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Novagold Resources Inc – ‘40FR12G’ on 10/29/03 – ‘EX-99.19’

On:  Wednesday, 10/29/03, at 5:20pm ET   ·   Accession #:  1062993-3-1067   ·   File #:  0-50443

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  As Of                Filer                Filing    For·On·As Docs:Size              Issuer               Agent

10/29/03  Novagold Resources Inc            40FR12G               91:19M                                    Newsfile Corp/FA

Registration of Securities of a Canadian Issuer — SEA’34 §12(g)   —   Form 40-F
Filing Table of Contents

Document/Exhibit                   Description                      Pages   Size 

 1: 40FR12G     Registration Statement Pursuant to Section 12 of    HTML     79K 
                          the Securities Exchange Act of 1934                    
 2: EX-99.1     Revised Initial Annual Information Form of the      HTML    228K 
                          Registrant Dated July 22, 2003                         
11: EX-99.10    Quarterly Report of the Registrant for the Six      HTML    137K 
                          Months Ended May 31, 2002                              
12: EX-99.11    Quarterly Report of the Registrant for the Three    HTML    139K 
                          Months Ended February 28, 2002                         
13: EX-99.12    Management Information Circular of the Registrant   HTML    101K 
                          Dated April 14, 2003                                   
14: EX-99.13    Form of Proxy for Use in Connection With the May    HTML     35K 
                          28, 2003 Annual and Special Meeting                    
15: EX-99.14    Management Information Circular of the Registrant   HTML     98K 
                          Dated April 15, 2002                                   
16: EX-99.15    Form of Proxy for Use in Connection With the May    HTML     33K 
                          22, 2002 Annual and Special Meeting                    
17: EX-99.16    Final Short Form Prospectus Dated September 25,     HTML    159K 
                          2003                                                   
18: EX-99.17    Underwriting Agreement Dated September 15, 2003     HTML    221K 
19: EX-99.18    Warrant Indenture Dated October 1, 2003             HTML    309K 
20: EX-99.19    Report Dated August 13, 2003, as Amended September  HTML    619K 
                          22, 2003                                               
 3: EX-99.2     Annual Information Form of the Registrant Dated     HTML    422K 
                          April 17, 2002                                         
21: EX-99.20    Qualifying Certificate of Ken Kuchling Dated        HTML     30K 
                          September 22, 2003                                     
22: EX-99.21    Report Dated April 16, 2002                         HTML    161K 
23: EX-99.22    Qualifying Certificate of Curtis J. Freeman Dated   HTML     29K 
                          April 16, 2002                                         
24: EX-99.23    Report Dated April 15, 2002                         HTML     88K 
25: EX-99.24    Qualifying Certificate of Curtis J. Freeman Dated   HTML     29K 
                          April 15, 2002                                         
26: EX-99.25    Report Dated April 1, 2002                          HTML    109K 
27: EX-99.26    Qualifying Certificate of Curtis J. Freeman Dated   HTML     29K 
                          April 1, 2002                                          
28: EX-99.27    Report Dated March 2002                             HTML    600K 
29: EX-99.28    Qualifying Certificate of Stephen B. Hodgson Dated  HTML     27K 
                          March 8, 2002                                          
30: EX-99.29    Qualifying Certificate of Stephen Juras Dated       HTML     28K 
                          March 8, 2002                                          
 4: EX-99.3     Annual Report of the Registrant for the Year Ended  HTML    281K 
                          November 30, 2002                                      
31: EX-99.30    Report Dated February 2002                          HTML    264K 
32: EX-99.30.A  Appendix A of Report Dated February 2002            HTML     26K 
33: EX-99.30.B  Appendix B of Report Dated February 2002            HTML   1.29M 
34: EX-99.30.C  Appendix C of Report Dated February 2002            HTML     60K 
35: EX-99.30.D  Appendix D of Report Dated February 2002            HTML     25K 
36: EX-99.31    Qualifying Certificate of Stephen Juras Dated       HTML     28K 
                          February 25, 2002                                      
37: EX-99.32    Material Change Report of the Registrant Dated      HTML     33K 
                          October 24, 2003                                       
38: EX-99.33    Press Release Dated October 23, 2003                HTML     31K 
39: EX-99.34    Material Change Report of the Registrant Dated      HTML     36K 
                          October 16, 2003                                       
40: EX-99.35    Material Change Report of the Registrant Dated      HTML     48K 
                          October 10, 2003                                       
41: EX-99.36    Material Change Report of the Registrant Dated      HTML     36K 
                          October 2, 2003                                        
42: EX-99.37    Material Change Report of the Registrant Dated      HTML     36K 
                          September 12, 2003                                     
43: EX-99.38    Material Change Report of the Registrant Dated      HTML     46K 
                          August 13, 2003                                        
44: EX-99.39    Material Change Report of the Registrant Dated      HTML     45K 
                          August 11, 2003                                        
 5: EX-99.4     Annual Report of the Registrant for the Year Ended  HTML    294K 
                          November 30, 2001                                      
45: EX-99.40    Material Change Report of the Registrant Dated      HTML     63K 
                          August 7, 2003                                         
46: EX-99.41    Material Change Report of the Registrant Dated      HTML     39K 
                          July 30, 2003                                          
47: EX-99.42    Material Change Report of the Registrant Dated      HTML     42K 
                          July 3, 2003                                           
48: EX-99.43    Material Change Report of the Registrant Dated      HTML     59K 
                          June 4, 2003                                           
49: EX-99.44    Material Change Report of the Registrant Dated May  HTML     34K 
                          1, 2003                                                
50: EX-99.45    Material Change Report of the Registrant Dated      HTML     46K 
                          April 28, 2003                                         
51: EX-99.46    Material Change Report of the Registrant Dated      HTML     47K 
                          April 9, 2003                                          
52: EX-99.47    Press Release of the Registrant Dated February 25,  HTML     30K 
                          2003                                                   
53: EX-99.48    Material Change Report of the Registrant Dated      HTML     44K 
                          February 11, 2003                                      
54: EX-99.49    Material Change Report of the Registrant Dated      HTML     70K 
                          February 6, 2003                                       
 6: EX-99.5     U.S. Gaap Reconciliation Which Includes Audited     HTML    279K 
                          Comparative Financial Statements                       
55: EX-99.50    Material Change Report of the Registrant Dated      HTML     92K 
                          January 30, 2003                                       
56: EX-99.51    Material Change Report of the Registrant Dated      HTML     37K 
                          December 30, 2002                                      
57: EX-99.52    Material Change Report of the Registrant Dated      HTML     72K 
                          December 11, 2002                                      
58: EX-99.53    Material Change Report of the Registrant Dated      HTML     55K 
                          November 26, 2002                                      
59: EX-99.54    Material Change Report of the Registrant Dated      HTML     65K 
                          November 14, 2002                                      
60: EX-99.55    Material Change Report of the Registrant Dated      HTML     70K 
                          October 3, 2002                                        
61: EX-99.56    Material Change Report of the Registrant Dated      HTML     37K 
                          September 19, 2002                                     
62: EX-99.57    Material Change Report of the Registrant Dated      HTML     54K 
                          September 13, 2002                                     
63: EX-99.58    Press Release of the Registrant Dated September 9,  HTML     28K 
                          2002                                                   
64: EX-99.59    Material Change Report of the Registrant Dated      HTML     35K 
                          September 5, 2002                                      
 7: EX-99.6     U.S. Gaap Supplement to Management Discussion and   HTML     42K 
                          Analysis                                               
65: EX-99.60    Material Change Report of the Registrant Dated      HTML     80K 
                          September 4, 2002                                      
66: EX-99.61    Material Change Report of the Registrant Dated      HTML     51K 
                          August 7, 2002                                         
67: EX-99.62    Material Change Report of the Registrant Dated      HTML     74K 
                          July 16, 2002                                          
68: EX-99.63    Material Change Report of the Registrant Dated      HTML     57K 
                          June 5, 2002                                           
69: EX-99.64    Material Change Report of the Registrant Dated May  HTML     76K 
                          22, 2002                                               
70: EX-99.65    Material Change Report of the Registrant Dated      HTML     43K 
                          April 30, 2002                                         
71: EX-99.66    Material Change Report of the Registrant Dated      HTML     42K 
                          April 18, 2002                                         
72: EX-99.67    Material Change Report of the Registrant Dated      HTML     37K 
                          March 26, 2002                                         
73: EX-99.68    Material Change Report of the Registrant Dated      HTML     78K 
                          March 15, 2002                                         
74: EX-99.69    Material Change Report of the Registrant Dated      HTML     56K 
                          February 18, 2002                                      
 8: EX-99.7     Quarterly Report of the Registrant for the Six      HTML    110K 
                          Months Ended May 31, 2003                              
75: EX-99.70    Material Change Report of the Registrant Dated      HTML     55K 
                          January 24, 2002                                       
76: EX-99.71    Consent of Pricewaterhousecoopers LLP               HTML     25K 
77: EX-99.72    Consent of Ken Kuchling                             HTML     26K 
78: EX-99.73    Consent of Curtis J. Freeman                        HTML     28K 
79: EX-99.74    Consent of Stephen B. Hodgson                       HTML     27K 
80: EX-99.75    Consent of Stephen Juras                            HTML     28K 
81: EX-99.76    Consent of Phillip St. George                       HTML     29K 
82: EX-99.77    Consent of Harry Parker                             HTML     27K 
83: EX-99.78    Consent of Norm Johnson                             HTML     27K 
84: EX-99.79    Consent of Norwest Corporation                      HTML     27K 
 9: EX-99.8     Quarterly Report of the Registrant for the Three    HTML    121K 
                          Months Ended February 28, 2003                         
85: EX-99.80    Consent of Avalon Development Corporation           HTML     27K 
86: EX-99.81    Consent of Amec E&C Services Limited                HTML     30K 
87: EX-99.82    Consent of Kennecott Exploration Company            HTML     26K 
88: EX-99.83    Consent of Newmont Mining Corporation               HTML     27K 
89: EX-99.84    Consent of Placer Dome Inc.                         HTML     27K 
90: EX-99.85    Consent of Mark Jutras                              HTML     27K 
91: EX-99.86    Consent of Robert Prevost                           HTML     27K 
10: EX-99.9     Quarterly Report of the Registrant for the Nine     HTML    148K 
                          Months Ended August 31, 2002                           


EX-99.19   —   Report Dated August 13, 2003, as Amended September 22, 2003
Exhibit Table of Contents

Page (sequential)   (alphabetic) Top
 
11st Page  –  Filing Submission
"1-1
"2-1
"2-2
"Figure 2.1
"Figure 2.2
"Figure 2.3
"3-1
"3-2
"Table 3.1
"3-4
"3-5
"Table 3.2
"3-7
"Table 3.3
"3-8
"Table 3.4
"Table 3.5
"4-1
"4-2
"4-3
"4-4
"Figure 4.1
"Figure 4.2
"Figure 4.3
"Figure 4.4
"4-10
"4-11
"4-12
"4-15
"4-16
"5-1
"5-3
"5-4
"6-1
"Figure 6.1
"Figure 6.2
"6-4
"Table 6.1
"Table 6.2
"Table 6.5
"Figure 6.3
"Figure 6.4
"Table 6.3
"Table 6.4
"7-1
"Figure 7.1
"7-3
"Table 7.1
"7-4
"Table 7.2
"8-1
"Figure 8.1
"Table 8.1
"Figure 8.2
"9-1
"Table 9.1
"9-3
"9-4
"10-1
"10-2
"Table 10.1
"Table 10.2
"10-4
"Table 10.3
"Table 10.4
"Table 10.5
"10-6
"Figure 10.1
"10-8
"10-9
"11-1
"11-2
"11-4
"Figure 11.1
"Table 11.1
"11-5
"Table 11.2
"11-6
"12-1
"Table 12.1
"Table 12.2
"Table 12.3
"12-3
"Table 12.4
"Table 12.5
"12-7
"Table 12.6
"12-8
"Table 12.7
"Table 12.8
"Table 12.9
"Table 12.10
"12-12
"Table 12.11
"Table 12.12
"12-16
"Table 12.13
"12-17
"Table 12.14
"13-1
"13-2
"13-3
"Table 13.1
"Figure 13.1
"13-6
"Figure 13.2
"13-8
"Table 13.2
"13-10
"Table 13.3
"Table 13.4
"Figure 13.3
"Figure 13.4
"13-14
"Figure 13.5
"Figure 13.6
"13-17
"Figure 13.7
"13-19
"14-1
"15-1
"15-2
"15-3
"Figure 15.1

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 <! 
  Filed by Automated Filing Services Inc. (604) 609-0244 - NovaGold Resources Inc. - Technical Report - August 13, 2003  


 

PRELIMINARY ECONOMIC STUDY

 

ROCK CREEK PROJECT
NOME, ALASKA

 

 

 

Submitted to:
NOVAGOLD RESOURCES INC.

 

August 13, 2003
Amended September 22, 2003

Prepared by:

Ken Kuchling, P.Eng. - Qualified Person:
Norwest Corporation

Suite 400, 205 – 9th Ave SE
Calgary, Alberta T2G 0R3
Tel:      (403) 237-7763
Fax:      (403) 263-4086
Email    calgary@norwestcorp.com

www.norwestcorp.com




 

 

TABLE OF CONTENTS

EXECUTIVE SUMMARY 1  
1 INTRODUCTION 1-1  
  1.1 TERMS OF REFERENCE 1-1  
  1.2 SOURCES OF INFORMATION 1-1  
  1.3 DISCLAIMER 1-1  
2 BACKGROUND 2-1  
  2.1 PROPERTY LOCATION AND DESCRIPTION 2-1  
  2.2 CLIMATE 2-1  
  2.3 LOCAL INFRASTRUCTURE 2-2  
3 GEOLOGICAL SETTING 3-1  
  3.1 REGIONAL GEOLOGY 3-1  
  3.2 LOCAL GEOLOGY 3-1  
  3.3 MINERAL RESOURCE 3-2  
    3.3.1 Computer Modelling (supplied by NovaGold) 3-2  
    3.3.2 Geological Model Review 3-4  
    3.3.3 Cut-Off Grades 3-5  
    3.3.4 Recovery Criteria 3-7  
    3.3.5 Block Model Summary 3-8  
4 GEOTECHNICAL & PERMAFROST 4-1  
  4.1 GEOTECHNICAL DESIGN CRITERIA 4-1  
    4.1.1 Pit Slopes 4-1  
    4.1.2 Development Rock Dumps 4-2  
    4.1.3 Tailings Impoundment 4-2  
    4.1.4 Infrastructure 4-3  
    4.1.5 Recommendations 4-3  
    4.1.6 Geotechnical Assessment 4-4  
  4.2 PERMAFROST 4-10  
    4.2.1 Introduction 4-10  
    4.2.2 Engineering Properties of Permafrost 4-11  
    4.2.3 Engineering Design in Permafrost 4-12  
    4.2.4 Impact of Global Warming 4-15  
  4.3 GROUNDWATER ISSUES 4-15  
  4.4 SITE WATER MANAGEMENT 4-16  
5 ENVIRONMENTAL & PERMITTING 5-1  
  5.1 INTRODUCTION 5-1  
  5.2 PERMITTING REQUIREMENTS 5-1  
  5.3 BASELINE STUDIES 5-3  
  5.4 POTENTIAL ISSUES 5-4  
  5.5 PERMITTING SCHEDULE AND BUDGET 5-4  
6 OPTIMIZATION STUDIES 6-1  
  6.1 PLANT SITE LOCATION 6-1  
  6.2 TAILINGS DISPOSAL OPTIONS 6-1  
  6.3 PIT OPTIMIZATION (LERCHS-GROSSMANN) 6-4  
7 MINE PLAN 7-1  
  7.1 PIT DESIGN 7-1  
 
NOVAGOLD RESOURCES INC. 03-2291
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OCK CREEK - PRELIMINARY ECONOMIC STUDY
TOC 1



 

  7.2 PRODUCTION SCHEDULE AND SEQUENCE 7-3  
  7.3 GRADE CONTROL 7-4  
  7.4 RECLAMATION AND CLOSURE 7-4  
8 WASTE DISPOSAL STRATEGY 8-1  
  8.1 ORGANIC SOILS 8-1  
  8.2 DEVELOPMENT ROCK MANAGEMENT 8-1  
  8.3 TAILINGS MANAGEMENT 8-1  
9 MINE EQUIPMENT 9-1  
  9.1 MINING EQUIPMENT 9-1  
    9.1.1 Loading Equipment 9-1  
    9.1.2 Hauling Equipment 9-3  
    9.1.3 Drilling and Blasting Equipment 9-3  
    9.1.4 Mine Support & Miscellaneous Equipment 9-3  
    9.1.5 AN Storage 9-4  
    9.1.6 Explosive Magazine 9-4  
10 MILLING 10-1  
  10.1 METALLURGICAL STUDIES 10-1  
    10.1.1 Early Studies 10-1  
    10.1.2 1999 Program 10-2  
    10.1.3 2001 Program 10-4  
    10.1.4 2003 Program 10-4  
    10.1.5 Discussion of Metallurgy 10-6  
  10.2 FUTURE METALLURGICAL WORK 10-8  
  10.3 PLANT DESIGN CONSIDERATIONS 10-8  
  10.4 BASIS OF CAPITAL AND OPERATING COST ESTIMATES 10-9  
11 PROJECT INFRASTRUCTURE AND SERVICES 11-1  
  11.1 ADMINISTRATION OFFICES 11-1  
  11.2 MINE MAINTENANCE SHOP & WAREHOUSES 11-1  
  11.3 MINE DRY 11-2  
  11.4 ELECTRIC POWER SUPPLY 11-2  
  11.5 FUEL STORAGE 11-4  
  11.6 SEWAGE COLLECTION AND TREATMENT 11-5  
  11.7 PROCESS WATER SUPPLY 11-6  
  11.8 POTABLE WATER 11-6  
12 COST ESTIMATE 12-1  
  12.1 CAPITAL COSTS 12-1  
    12.1.1 Basis of Estimate 12-1  
    12.1.2 Project Development Costs 12-3  
  12.2 SUSTAINING CAPITAL COSTS 12-7  
  12.3 OPERATING COSTS 12-7  
    12.3.1 Manpower 12-8  
    12.3.2 Mining Operating Cost 12-12  
    12.3.3 Mill Operating Cost 12-16  
    12.3.4 Reclamation Operating 12-16  
    12.3.5 Administration Operating 12-16  
    12.3.6 Summary of Operating Costs 12-17  
13 FINANCIAL ANALYSIS 13-1  
 
NOVAGOLD RESOURCES INC. 03-2291
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OCK CREEK - PRELIMINARY ECONOMIC STUDY
TOC 2



 

  13.1 FINANCIAL ASSUMPTIONS 13-1  
    13.1.1 Projections 13-1  
    13.1.2 Net Smelter Returns 13-1  
    13.1.3 Royalties and State Mining Tax 13-1  
    13.1.4 Income Taxes 13-2  
    13.1.5 Working Capital 13-3  
    13.1.6 Salvage Value 13-3  
  13.2 MILLING OPTION COMPARISON 13-3  
  13.3 OPERATING COST PROFILE 13-6  
  13.4 CAPITAL COST PROFILE 13-8  
  13.5 PROJECT ECONOMICS 13-10  
  13.6 FINANCIAL SENSITIVITIES 13-14  
  13.7 RISKS & OPPORTUNITIES 13-17  
  13.8 UPSIDE TARGET SCENARIO 13-19  
14 RECOMMENDATIONS FOR FEASIBILITY DESIGN STUDY 14-1  
15 PROJECT EXECUTION SCHEDULE 15-1  
  15.1 OUTSIDE AGENCY RESPONSIBILITIES 15-2  
  15.2 CONSTRUCTION LOGISTICS 15-3  
 
NOVAGOLD RESOURCES INC. 03-2291
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OCK CREEK - PRELIMINARY ECONOMIC STUDY
TOC 3



 

List of Figures

Figure 2.1 Project Location Map
Figure 2.2 Monthly Air Temperatures
Figure 2.3 Monthly Precipitation and Snowfall
   
Figure 4.1 Pit Wall RQD Data
Figure 4.2 Overburden & Regolith Core Photo
Figure 4.3 Topography in Waste Dump Areas
Figure 4.4 Conceptual Tailings Dam Design
   
Figure 6.1 Plant Site Location
Figure 6.2 Tailings Location Options
Figure 6.3 Optimized Pit Resource Sensitivity to Gold Price
Figure 6.4 Lerchs-Grossman Optimized Pit
   
Figure 7.1 Final Pit Design
   
Figure 8.1 Project Site Layout
Figure 8.2 Tailings Storage Curve
   
Figure 10.1 Gravity Recovery versus Grind Size
   
Figure 11.1 Nome Power Grid - single line diagram
   
Figure 13.1 IRR vs Milling Case
Figure 13.2 Operating Costs
Figure 13.3 Net Cash Flow After Tax
Figure 13.4 Unit Cost and Margins
Figure 13.5 NPV Sensitivity
Figure 13.6 IRR Sensitivity
Figure 13.7 Cumulative Impact on IRR
   
Figure 15.1 Project Execution Schedule
 
NOVAGOLD RESOURCES INC. 03-2291
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OCK CREEK - PRELIMINARY ECONOMIC STUDY
TOC4



 

List of Tables

Table 3.1 Rock Creek Block Model Summary
Table 3.2 Cut-Off Grade Sensitivity
Table 3.3 Breakeven Cut-Off Grade Matrix
Table 3.4 Recovery Criteria
Table 3.5 Block Model Resource
Table 6.1 Tailing Facility Options
Table 6.2 Pit Optimization Input Parameters
Table 6.3 Pit Optimization Summary (Variable Cut-off Grade)
Table 6.4 Pit Optimization Summary (Cut-off Grade = 1.0 g/t)
Table 6.5 Optimized Pit Resource
Table 7.1 Pit Resource (Cut-off = 1.0 g/t)
Table 7.2 Production Schedule Summary
Table 7.3 Production Schedule Detail
Table 8.1 Development Rock Material Balance
Table 9.1 Loading Equipment Productivity
Table 10.1 Gravity Concentration Results
Table 10.2 Pilot Gravity & Cyanidation of Mids/Tails Results
Table 10.3 Direct Agitated Cyanidation Testing
Table 10.4 Knelson Gravity Concentration Results
Table 10.5 Falcon Gravity Concentration Results (2003)
Table 11.1 Diesel Generated Electric Power Cost (Fuel only component)
Table 11.2 Diesel Fuel Cost Breakdown
Table 12.1 Indirect Cost Basis
Table 12.2 EPCM Cost Basis
Table 12.3 Allowance Basis
Table 12.4 Mining Equipment Capital Cost
Table 12.5 Annual Equipment Requirements
Table 12.6 Summary of Initial Capital Cost
Table 12.7 Manpower Summary
Table 12.8 Staffing Detail
Table 12.9 Mine Operations Detail
Table 12.10 Mill Operation Detail
Table 12.11 Unit Mining Activity Calculations
Table 12.12 Mining Cost Summary
Table 12.13 Mill Operating Cost Summary
Table 12.14 Operating Cost Summary
Table 13.1 Milling Options
Table 13.2 Capital Summary
Table 13.3 Project Economics
Table 13.4 Cash Flow Summary
 
NOVAGOLD RESOURCES INC. 03-2291
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OCK CREEK - PRELIMINARY ECONOMIC STUDY
TOC 5



 

EXECUTIVE SUMMARY

 

NovaGold Resources Inc. (NovaGold) commissioned Norwest Corporation (Norwest) to complete a preliminary economic study, mine plan, and cost estimate for the Rock Creek Project. In addition a summary of key environmental issues and permitting requirements has also been compiled. All amounts in this study are in US dollars unless otherwise stated.

Various sources of information have been used to develop the preliminary mine plan and cost estimate. NovaGold provided the geological data base, interpretation and the geological block model for Rock Creek. Information related to milling and ore recoveries are based on independent test work previously supervised by NovaGold. A single gold recovery test was completed by Norwest to verify and build on the previous test work. Capital and operating costs and equipment productivities are based on data from Norwest's in-house database and costing information from published sources.

This Preliminary Assessment includes the use of inferred resources that are considered too speculative geologically to have economic considerations applied to them that would enable them to be categorized as mineral reserves. Inferred gold resources will require further exploration to upgrade them to the higher Measured and Indicated categories.

The Rock Creek Project is located on the Seward Peninsula along the west coast of Alaska, north of Norton Sound. The project area lies about 13 km north of Nome and is accessed via state maintained roads. The city of Nome (population 4,000) is situated on the Bering Sea coast and serves as the logistical and administrative center for this portion of western Alaska. Nome has daily commercial jet service from Anchorage and has large container barge service from June through October. The project occurs partly on patented mining claims owned 100% by the Alaska Gold Co., a wholly owned subsidiary of NovaGold, and partly on land owned by Bering Straits Native Corporation (BSNC) subsurface, and Sitnausuak (the Nome native village corporation) surface, under agreement with NovaGold. The project site is characterized by cool summers and cold winters. Summer temperatures range from +80C to +15‹ C and winter temperatures average around -15°C.

Mineralization at Rock Creek is confined to quartz veining in two ore types; tension veins and the Albion shear zone. The resource in the tension veins accounts for approximately 90% of the tonnes and 85% of the gold content in the deposit. The two ore types each have different process recovery criteria, for reasons that may be related to sulfo-salt and silicification. Previous studies indicate that the Albion vein quartz poses the greater challenge in gravity recovery circuits, with recovery values in the range of 30%-40% for coarse grind material. The tension vein material yields higher gravity recoveries in the range of >80%. Cyanidation of the whole rock yields


 
NOVAGOLD RESOURCES INC. 03-2291
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fairly high recoveries for both ore types. Finer grinding tends to liberate gold and helps to improve recovery. For design purposes, it is assumed that gold recoveries for the two ore types will be 94% for the Albion zone and 97% for the tension vein material, based on a process relying on a fairly fine grind (80% < 140 microns).

The site specific geotechnical data gathered to date is limited to core descriptions, rock quality (RQD) and hardness information from 2002 and 2003 drilling. It is recommended that an inter-ramp pit wall angle of 45 degrees be adopted in scoping pit design. Although there is no geotechnical information available on the proposed plant site, foundations for buildings and the various items of plant are not expected to require any special methods of construction other than those dictated by permafrost conditions. As the project proceeds through the various levels of design, site specific information will need to be gathered to support the basic mine, development rock and tailings facility designs, as well as the design of building and plant foundations. To the best of our knowledge no thermistors or permafrost measurements have been made at the Rock Creek site although published data indicates that permafrost depths in Nome could be approximately 60 to 100 metres.

The proposed mine development and operation would have to obtain numerous State and Federal permits before development and operation can proceed. Various Federal and State agencies have jurisdiction over many of the activities proposed for the Rock Creek Project. The types of permits and approvals, which may be needed in order to construct and operate the Rock Creek Project, will depend on the design and operation of the mine. It is assumed that all project activities will occur on private or state lands and that no federal land use or rights-of-way will be required. Environmental baseline studies are required in support of the preparation of a number of State and Federal permit applications. Baseline information and data regarding surface water, groundwater, tailings and development rock geochemistry soils, vegetation, fish and wildlife, and threatened/endangered species, are necessary for design of the mining (i.e., extraction) and reclamation plan that must be prepared in support of the Plan of Operations application and the Solid Waste permit application. It is estimated that baseline collection will take about a year and then permitting will take an additional year.

Prior to developing the pit designs and mine layouts, the Lerchs-Grossmann (LG) pit optimizer was used to evaluate pit geometries based on variable economic factors. Each variation in revenue or cost would result in a different optimized pit. The Rock Creek pit has been design based on a gold price of US$325/oz and a mining cut-off grade of 1.0 g/t.


 
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A Base Case mine development plan and schedule consisting of a three-phase pit based on the LG pit was generated. A Target Upside Potential mine schedule based on NovaGold’s reasonable expectation that the block model resource and block model gold quantity at Rock Creek will increase due to additional resource drilling within the main orebody, drilling along untested extensions of the ore zone, and improved geological modeling and interpretation techniques resulting in improved ore grades was also generated. The estimated pit resource for the Base Case and the Target Upside Scenario is summarized below and is as yet unclassified into resource categories.
           
      Base Case Upside Target
Scenario
 
           
    Waste tonnes (millions) 47.5 74.8  
    Ore tonnes (millions) 10.7 16.8  
    Ore Grade (g/t) 2.03 2.03  
    Waste:Ore ratio 4.44 4.44  
    Gold, In-Pit Resource (oz) 699,500 1,100,200  
    Gold, recovered (oz) 674,000 1,059,100  
     
 

Production scheduling and mining equipment requirements are based on an assumed project operating schedule of two12-hour shifts per day for 360 days per year. Based on the waste to ore ratio, it is expected that the mining operations will either focus on ore or waste mining operations on a shift-by-shift basis. Probably four out of every five shifts would be assigned to waste stripping operations. During the periods when the fleet is stripping waste, the mill will continue to operate by reclaiming ore from the live ore stockpile. Ore mining rates are about 1.78 million tonnes per year while waste volumes range from 6.8 to 9.2 million tonnes per year.

The two main waste products to be generated by the Rock Creek mining operation will be the development rock stripped from the mine and tailings generated by the milling process. Each of these products will be disposed of separately. Development rock stripped from the pit will be placed into development rock dumps located adjacent to the pit. If pit sequencing permits, some development rock may be backfilled into mined out portions of the pit. Some development rock will also be used periodically for raising the tailings dam. A single tailings facility will be used to store the mill tailings, provide seasonal water inventory for processing needs and act as a storm water runoff buffer. The tailings will be retained behind a centerline constructed containment dam, which will be constructed using mined development rock and will incorporate a graded filter zone to mitigate piping issues.


 
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The major mining equipment fleet for the Rock Creek Project will be modest in size. Two 12 m3 front end loaders, equivalent to a CAT 992G loader, will provide sufficient capacity for the ore and waste loading operations. A truck fleet of four 100 tonne trucks will provide operating capacity. Other mining equipment will consist of a blasthole drill, dozers, graders, and various pickup trucks and service vehicles.

A screening of nine different milling options was undertaken, including gravity, flotation, and cyanidation processes. The preferred case consists of a gravity circuit supplemented with a flotation circuit. One benefit of flotation is that it may be possible to achieve good recoveries at a relatively coarser grind than required for gravity alone. The gravity circuit would produce a relatively clean gold stream that would be refined into gold dore bars at site. The flotation circuit would produce a concentrate that would be shipped off-site for further processing, such as cyanidation. Approximately 82% of the recovered gold would report to the dore circuit. Further test work is required to support these assumptions.

The plant site area includes space for the primary crusher and crushed ore stockpile, the mill, a maintenance shop, an administration and mine dry building, and fuel storage. Ammonium nitrate and explosives will be stored remotely away from the main plant area. It is estimated that a total power requirement of 4 to 5 MW will be needed and will be supplied from the Nome Joint Utility System or on-site generation. The estimated cost for electric power from the local utility is US$0.12/kwh however further discussion with the power utility is required to confirm the responsibility for installing the step-up transformer and upgrading the power line to 25 kv from 125 kv.

The total project manpower is estimated at about 123 personnel, as summarized in the table below.


Admin Staff 17
Mining - Operating & Maintenance 63
Milling - Operating & Maintenance 43
Total Personnel 123

 
Project operating costs will average about $22 million per year in full production. On a unit cost basis, the operating cost should average about $12.08 per tonne milled, as shown below:

 
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Mining (per tonne mined) $0.82       
Mining ($/t milled) $4.44       
Milling ($/t milled) $6.40       
Admin ($/t milled) $0.95       
Reclamation ($/t milled) $0.29       
Total $12.08       

  Project initial capital costs will be about $39.2 million, sub-divided as follows:

Infrastructure/Development
Milling and Processing
Mining Equipment
EPCM, Indirects, Owners Costs, and Allowances

Total
US$5.0 M       
US$16.5 M       
US$11.4 M       
US$6.3M       
US$39.2 M       

  Based on the preliminary mine plan and cost estimates, the after tax economics of the project are:



Net Present Value (at 5% discount rate)

Internal Rate of Return (IRR) - after tax
Base Case Target Upside
Scenario
US$16.0 M

16.2%
US$34.5 M

21.7%

 

The project economics are most sensitive to the gold price and the grade of the ore body. Each 10% change affects the NPV by about 77% and the IRR by approximately 51% in the Base Case.

Operating cost changes of 10% will impact the NPV by 45% and the IRR by 31%. The capital cost affects the NPV by about 22% and the IRR by 20% for each 10% increase or decrease in the capital cost in the Base Case.

NovaGold geological staff indicate that there is a reasonable expectation that the in-pit resource and block model gold quantity at Rock Creek will increase due to additional resource drilling within the main orebody, drilling along untested extensions of the ore zone, and improved


 
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geological modeling and interpretation techniques resulting in improve ore grades. NovaGold estimates that this could result in the addition of 300,000 to 400,000 ounces of block model gold to the pit resource. The incremental capital expenditure for additional ore production from Rock Creek would likely be minimal. If one assumes a total block model gold resource of approximately 1.1 million ounces, three more years of mine production at average ore grades and average strip ratios would occur. A summary of the base case project economics and the upside target scenario is provided below.

  Base Case Plan Upside Target
Scenario
Mine Life 6 years 9.5 years
Total Ore 10.7Mt 16.8Mt
Gold, block model 699,500 oz 1,100,200 oz
Gold, recovered 673,900 oz 1,059,100 oz
     
IRR 16.2% 21.7%
NPV (0%) $30.3M $61.8M
NPV (5%) $16.0M $34.5M

 
The Project Execution Schedule to undertake continued development, construction and commissioning of the Rock Creek Project is described below. The project start-up date is January 2006. The key factor impacting on the development schedule are the open water seasons allowing barge access to bring major equipment and components to site. From today to the start of production there are three barge seasons remaining; summer 2003, 2004, and 2005. Effectively only the 2005 summer provides the most realistic shipping target.

Preliminary Economic Study completed: 3rd quarter 2003
Feasibility contractor selected: 4th quarter 2003
Feasibility study completed: 2nd quarter 2004
NovaGold project decision: 2nd quarter 2004
Permitting, initiate process: 3rd quarter 2004
Permitting, completion: 3rd quarter 2005
Detailed engineering commences: 3rd quarter 2004
Major equipment and components ordered: 4th quarter 2004
Equipment delivery to site: 3rd quarter 2005
Construction start: 3rd quarter 2005
Rock Creek Project commission and production: 1st quarter 2006

 
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1 INTRODUCTION
   
1.1

TERMS OF REFERENCE

NovaGold Resources Inc. (NovaGold) commissioned Norwest Corporation (Norwest) to complete a preliminary economic study, mine plan, and cost estimate for the Rock Creek Project. In addition a summary of key environmental issues and permitting requirements has also been compiled.

   
1.2

SOURCES OF INFORMATION

Various sources of information have been used to develop the preliminary mine plan and cost estimate.

NovaGold provided the geological data base, interpretation and the geological block model for Rock Creek. Norwest validated the model based on data provided and assumptions used. There was no testing of assumptions or verification of the database.

Information related to milling and ore recoveries are based on independent test work previously supervised by NovaGold. A single gold recovery test was completed by Norwest to verify and build on the previous test work.

Operating and capital costs and equipment productivities are based on data from Norwest's in-house database and costing information from published sources.

   
1.3

DISCLAIMER

This report has been prepared for NovaGold Resources Inc. (NovaGold) by Norwest Corporation (Norwest). The findings, conclusions, and cost estimates provided are based on information developed by Norwest available at the time of preparation and data supplied by outside sources, including NovaGold staff and consultants working for NovaGold.

This preliminary economic study has been prepared under the supervision of Ken Kuchling, P.Eng. the designated Qualified Person. He completed a site visit to the Rock Creek property on April 22 and 23, 2003. Ken Kuchling has not verified any of the existing geological data nor has he verified the assumptions used to build the block model. Data verification and block modelling was undertaken by NovaGold technical staff. The completed block model was provided to Norwest by NovaGold for use in this study.


 
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2  BACKGROUND
   
2.1

PROPERTY LOCATION AND DESCRIPTION

The Rock Creek Project is located on the Seward Peninsula along the west coast of Alaska, north of Norton Sound. The project area lies about 13 km north of Nome and is accessed via state maintained roads, as shown in Figure 2.1.

The terrain is fairly hilly with broad and narrow valleys. Nome is located at sea level while the Rock Creek plant site would be at an elevation of about 70 masl (metres above sea level). The mine area is higher, at elevations ranging from 100 masl to 150 masl.

Vegetation at the site consists mainly of low shrubs and grasses. Forested areas and trees are non-existent in the mine area.

The project occurs partly on patented mining claims owned 100% by the Alaska Gold Co., a wholly owned subsidiary of NovaGold and partly on land controlled by the Bering Straits Native Corporation (BSNC). Bering Straits Native Corp also owns local mineral rights with surface rights owned by Sitnausuak (the Nome native village corporation). The Bering Straits agreement involves escalating annual payments up to production, a 2.5 % Net Smelter Return (NSR) royalty and a 5% Net Profits Interest from production from BSNC lands. The known resource at Rock Creek lies within land owned approximately 66% by Alaska Gold Co. with the remainder within Bering Straits Native Corporation lands.

The nearest area to the Rock Creek prospect that is closed to mineral entry is the Bering Land Bridge National Preserve which is over 60 miles northeast of the Rock Creek prospect at its closest point. There currently are no unusual social, political or environmental encumbrances to exploration, development or production on the prospect.

   
2.2

CLIMATE

The project site is characterized by cool summers and cold winters. Summer temperatures range from +80C to +150C and winter temperatures average around -150C. Figure 2.2 provides a chart of seasonal air temperatures, showing mean, and average high and low temperatures.

The project site is also characterized by relatively low annual precipitation averaging less than 42 cm with the majority of precipitation falling as rain in summer. Figure 2.3


 
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provides a chart of seasonal precipitation and snowfall. Monthly winter snowfall totals range from 12 cm to 28 cm on average.
   
2.3

LOCAL INFRASTRUCTURE

The city of Nome (population 4,000) is situated on the Bering Sea coast and serves as the logistical and administrative center for this portion of western Alaska. Nome has daily commercial jet service from Anchorage and has large container barge service from June through October.

The Rock Creek prospect is currently road accessible via the Glacier Creek Road and the state maintained Teller-Nome Highway, an all-weather paved and gravel road. There are plans by the State of Alaska to construct the Glacier Creek Road By-Pass, shown in Figure 2.1, and engineering for the road is underway. Construction of the 5 km long by-pass road will simplify the road access distance to site.

The city of Nome has provided electricity to past mining operations and has offered that service for future operations if necessary. Current generating capacity is about 12 MW with three diesel generators and there are discussions that local capacity will be increased up to 15 to 20 MW with newer more energy efficient diesel generators. The current local power consumption is in the range of 4 to 6 MW.

No camp facilities are required at the Rock Creek Project due to its close proximity to Nome, which is well serviced with accommodations.


 
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FIGURE 2.2
MONTHLY AIR TEMPERATURES

FIGURE 2.3
MONTHLY PRECIPITATION AND SNOWFALL

 
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3  GEOLOGICAL SETTING
   
3.1

REGIONAL GEOLOGY

The Seward Peninsula is composed of a complex series of metamorphic rock packages that have been affected by large-scale accretionary, rotational and transcurrent events.

The project area is underlain by metamorphic rocks belonging to the Nome Group, which is thought to constitute a coherent litho-stratigraphic succession consisting of four major units;


  1. a basal, complexly deformed pelitic schist,
  2. a “mixed unit” of mafic, pelitic, and calc schists and marble,
  3. a mafic dominated schist package,
  4. an impure marble package.
   
 
The protoliths of these rocks are thought to be Cambrian to Devonian sediments consisting of shales, siltstones, sandstones, marls and limestones, deposited in a shallow water continental platform. Nome Group rocks have undergone at least two periods of metamorphism and accompanying deformation.
   
3.2

LOCAL GEOLOGY

The Rock Creek mineralized zone covers an area of approximately 500 by 1500 meters. The main rock types are schist and quartz muscovite schist (QMS). QMS is dominant but other types of schist are common in the deposit area. The overall bedding is relatively flat with a slight dip to the northwest. Post-mineralization faulting is evident in the core with gouge zone typically 1 to 3 meters wide. Mineralization at Rock Creek is confined to quartz veining. The gold grains are coarse and visible gold is very common.

Exploration efforts have determined that gold was deposited at Rock Creek in replacement bodies, in tension veins, and in the Albion shear zone. Only the latter two, i.e. tension veins and Albion shear veins, were modeled. The resource in the tension veins accounts for approximately 90% of the tonnes and 85% of the gold content in the deposit.

   
  1.
Replacement Bodies: This style of mineralization is dominated by albite, quartz, dolomite and arsenopyrite plus minor galena. Metallurgical studies indicate the gold in replacement bodies occurs as fine free gold in association with sulfides (Vance and others, 1993). There is probably more than one episode of replacement mineralization at Rock Creek. The first occurred in a ductile environment and is

 
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characterized by large crystals of albite, quartz, arsenopyrite, and dolomite. These bodies occur in low angle structures and small fold noses. They can appear to be sillform but thin section analyses show trains of carbon and resistant minerals from the original rocks passing through the crystals. Tension fracture veins cut these bodies at several locations, although occasionally the replacement bodies appear to become thicker when approaching a tension vein. Although replacement bodies are small and irregular they probably add to mineral inventories if they occur within a tension vein zone and are not modeled geologically.
     
  2.
  
Tension Veins: Tension veins at Rock Creek normally are less than 10 centimetres wide with strike and dip lengths of 10 to 100 meters. The quartz veins appear to have more strike length than vertical extent. The tension vein zones contain multiple veins and can be hundreds of meters wide. Much of the native gold seen in the veins is in fractures, especially in arsenopyrite. There usually is anomalous gold associated with tension veins but the best grades appear close to mineralized shears. Tension veins can merge with flat-lying veins at Rock Creek however the extent of this type of vein geometry is unknown.
     
  3.
Shear Veins: The shear-hosted veins such the Albion shear often are wider than individual tension veins and exhibit episodic fracturing and quartz banding and quartz-cemented breccias. The quartz veins within shear zones range from few centimeters wide to over 3 meters wide. The quartz is more apheric and less fractured than quartz in tension veins. Fine-grained pyrite and lead sulfosalts give the veins a bluish color. These veins are usually highly anomalous in gold, and the presence of these veins is considered to be important from an economic evaluation standpoint. Shear veins appear to be the youngest mineralizing event at Rock Creek.
     
3.3

MINERAL RESOURCE

The block model used for resources estimation and pit planning has been developed by NovaGold. Prior to using the geological model for mine planning, Norwest completed a cursory check of the model, the results of which are described in Section 3.3.2.

     
  3.3.1

Computer Modelling (supplied by NovaGold)

The resource model for the Rock Creek Gold Deposit is built on the geological model as defined in drill holes. Geological continuity is established and defined by the stratigraphic model. Structural and mineralization controls are somewhat poorly captured and are not modeled and the current mineralized shell is purely based on assay.


 
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Two geologic domains were modeled to restrain the interpolation at Rock Creek.

(a) The Albion zone is a persistent sub-vertical shear zone that extends beyond the resource area along strike and at depth. It is composed of several tightly spaced quartz veins within a corridor of strong shearing and alteration. It consists of the Albion east and Albion west structures.

(b) The tension vein quartz muscovite schist (QMS) zone is relatively flat lying and appears to be controlled in part by host stratigraphy. The tension vein zone contains multiple narrow sub-vertical sheeted tension veins that have the same strike and dip as the Albion zone shear. Cross-section work shows that only the tension veins host significant tonnage of sheeted veins so the tension veins are used to limit kriging of the sheeted veins both laterally and vertically. Below the tension veins, the rocks are poorly mineralized and have been excluded from the resource model.

The Rock Creek model (project area) was built using geological information and assay data collected from 99 core holes (10,092 m), 118 RC holes (8,807 m). A total of 10,222 assays were used in the model. The modeling philosophy is based on using geologic features and distribution of gold mineralization in controlling the extent of interpolation. The Albion veins and the QMS were modeled as 3D solids.

The block model used for the scoping study resource estimation of the Rock Creek Gold Deposit was constructed using MineSight® mine modeling software and SaGe® for variography. The final model consists of 4,080,000 blocks, and all blocks have dimensions of 5m x 5m x 5 m. Sub-blocking was used along the Albion contacts. The model was formulated in local mine grid coordinate system which is rotated 50 degrees from true north.


 
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TABLE 3.1
ROCK CREEK BLOCK MODEL SUMMARY

Direction Model Limit (m)
(mine grid)
Number
of Blocks
Block Size
(m)
East (X) 0 to 1000 200 5
North (Y) -200 to 1500 340 5
Elevation (Z) -100 to 200 170 5

   

The items modeled include topography, felsic intrusive, faults, specific gravity, mineralized zone and in situ gold grade. A current surface topographic model that covers the resource model was used to code that portion of the blocks below the surface. A field in the block model named TOPO was initialized and coded with the percentage of each block below the surface. Overburden is minimal in the Rock Creek resource area and therefore no overburden has been coded in the model.

Mineralized shells were developed for both the Albion zone and the tension vein zone. All blocks, or portion of blocks, within the Albion zone are considered mineralized. For the tension vein zones, acceptable mineralized envelops were defined by utilizing Probability Assisted Constrained Kriging or PACK (Pan, 1995). NovaGold's project geologists and a geologist representing AMEC selected the shell used to outline the populations. Finally, it was assured that the nominated boundaries did not violate current geologic understanding of mineralization controls.

Ellipsoidal search was used to limit the interpolation and to avoid smearing grade laterally. Rotations and ranges are based on geological information and geostatistical analysis.

     
  3.3.2

Geological Model Review

The geological model development for pit designs and subsequent economic evaluations is outside Norwest’s scope of work. However, prior to proceeding with this type of assessment several preliminary checks are completed to ensure the model is reasonable. This level of checking does not test or validate assumptions, rather the methodology generating the model is reviewed and the model is compared to the base data and assumptions.


 
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    In completing this review Norwest completed the following:
       
    Check of the Minesite® run procedures
    Complete Exploratory Data Analysis (EDA) variography
    Check of model for level of smoothing
    Check of resource calculations
    Review of pit design and drill hole distributions
       
   

The results of some of these initial analyses were discussed with NovaGold staff and the model was subsequently updated. Further enhancements will be incorporated at the feasibility stage.

Norwest is satisfied that the model used for the pit evaluations is acceptable based on the assumptions used.

     
  3.3.3

Cut-Off Grades

Two types of cut-off grade can be applied when defining the mineable resource, the breakeven cut-off grade and the mill cut-off grade.

The breakeven cut-off grade is used to define the pit limit and determine whether a block of ore should be mined or not. This cut-off is most applicable to ore blocks adjacent to the pit walls, which do not need to be mined if sub-economic. The mill cut-off grade applies mainly to internal marginal waste blocks that must be stripped as part of the mine plan. In this case the decision is whether to haul the block to the development rock dump or to the mill.

The breakeven cut-off grade must support all costs, including mining, milling and administration. The calculated breakeven cut-off grade based on a gold price of US$325/oz is estimated at about 0.85 g/t. The sensitivity of the cut-off grade to various economic factors has also been evaluated, including milling cost, gold price, dilution, and ore losses. Table 3.2 summarizes the basis and sensitivity of the breakeven cut-off grade for gold price. Table 3.3 provides a matrix comparing the cut-off grade against variable gold prices and Milling, Administration, and Sustaining capital (MAS) costs. For the purposes of this scoping level study, a conservative breakeven cut-off grade of 1.0 g/t is recommended.

The mill cut-off grade must support all costs, except for the mining cost since the block of ore must be mined anyway. This assumes the cost of mining and hauling waste is the same as that for mining and hauling ore. The mill cut-off grade based


 
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on a gold price of US$325/oz is estimated to be about 0.1 g/t less than the breakeven grade. Table 3.2 also summarizes the sensitivity of the mill cut-off grade for gold price. The mill cut-off grade is not used in the production schedule for this study but may be considered at the feasibility stage for stockpiling of marginal ore.

When describing the in-pit resource in Section 7.1, the breakeven cut-off grade of 1.0 g/t will be used to summarize the resource. However the additional resource contained between the actual breakeven cut-off (0.85 g/t) and 1.0 g/t cut-off grade will also be quantified.

TABLE 3.2
CUT-OFF GRADE SENSITIVITY

  Variable Gold Price
Gold Price US$/oz $275 $300 $325 $350 3$75
Payable metal % 100% 100% 100% 100% 100%
Refining US$/oz $2.00 $2.00 2.00 $2.00 $2.00
Net Payable US$/oz $273.00 $298.00 323.00 $348.00 $373.00
Net Payable $/gm $8.78 $9.58 10.39 $11.19 $11.99
Mining Cost $/t mined $0.92 $0.92 0.92 $0.92 $0.92
Milling Cost (M) $/t milled $6.30 $6.30 6.30 $6.30 $6.30
G&A Cost (A) $/t milled $1.00 $1.00 1.00 $1.00 $1.00
Recl, Sust capex (S) $/t milled $0.20 $0.20 0.20 $0.20 $0.20
Total cost $/t milled $8.42 $8.42 8.42 $8.42 $8.42
Mill Recovery % 95% 95% 95% 95% 95%
Breakeven Cut-off Grade g/t 1.01 0.92 0.85 0.79 0.74
Mill cut-off g/t 0.90 0.82 0.76 0.71 0.66

   
The cut-off grades described above are based on preliminary costs completed prior to developing the actual scoping study mine plan. Final project costs may differ somewhat from these initial estimates.

 
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TABLE 3.3
BREAKEVEN CUT-OFF GRADE MATRIX

Gold Price and 95% Recovery
MAS Cost $200 $225 $250 $275 $300 $325 $350 $375 $400
$5.00 0.97 0.86 0.78 0.70 0.65 0.60 0.55 0.52 0.48
$5.50 1.05 0.93 0.84 0.76 0.70 0.65 0.60 0.56 0.53
$6.00 1.13 1.01 0.91 0.82 0.76 0.70 0.65 0.60 0.57
$6.50 1.21 1.08 0.97 0.88 0.81 0.75 0.69 0.65 0.61
$7.00 1.30 1.15 1.04 0.94 0.86 0.80 0.74 0.69 0.65
$7.50 1.38 1.23 1.10 1.01 0.92 0.85 0.79 0.74 0.69
$8.00 1.46 1.30 1.17 1.06 0.97 0.90 0.83 0.78 0.73
$8.50 1.54 1.37 1.23 1.12 1.03 0.95 0.88 0.82 0.77
$9.00 1.62 1.44 1.30 1.18 1.08 1.00 0.93 0.87 0.81
$9.50 1.71 1.52 1.36 1.24 1.14 1.05 0.97 0.91 0.85
$9.60 1.72 1.53 1.38 1.25 1.15 1.06 0.98 0.92 0.86
$10.00 1.79 1.59 1.43 1.30 1.19 1.10 1.02 0.95 0.89

  3.3.4

Recovery Criteria

The two ore types that have been modelled each have different process recovery criteria. The reasons for the differences in recovery are uncertain at this time but may be related to sulfosalt and silicification.

Section 10.1 describes the results of metallurgical test work. These studies indicate that the Albion vein quartz poses the greater challenge in gravity recovery circuits, with values in the range of 30%-40%. The tension vein material yields higher gravity recoveries in the range of 80%. Cyanidation of the whole rock yields fairly high recoveries for both ore types.

The application of recovery criteria to the tension vein ore is fairly straightforward. Processing the Albion zone represents a more complex situation. The Albion zone can consist of a mix of both Albion quartz and tension vein type quartz. Geological logging records percent volume of Albion quartz (within a 5-level classification scheme) and percent volume of tension vein quartz and three other vein types, along with a count of tension veins. However logging records those values in natural geological intervals, not in the 2-metre long sample intervals. Thus identifying the percent volume of quartz types in a particular sample can be a tedious and non-automated effort if logged geological breaks don't coincide with sample breaks. Both Albion and tension vein quartz can be intimately


 
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intermingled and a very small volume of one type can exist in an intercept dominated by the other.

One could expect that theoretically each Albion zone mining block might potentially have a different recovery depending on the ratios of Albion quartz to tension vein quartz contained within. Furthermore the results of the metallurgical lab testing will also depend on the respective quartz ratios. Since all of the metallurgical samples tested likely consisted of mixed quartz there maybe no real data on the recovery rate for "pure" Albion quartz.

For design purposes, Table 3.4 summarizes the recovery factor used in pit optimization and project design, assuming a gravity and flotation milling operation.

TABLE 3.4
RECOVERY CRITERIA (GRAVITY & FLOTATION PROCESS)

  Albion Zone
Ore
Tension Vein
Ore
Gravity / Flotation Milling 94% 97%

  3.3.5

Block Model Summary

NovaGold staff has prepared the geological block model for the purposes of this study. Table 3.5 summarizes the unclassified block model resource estimate. Further documentation on the block model resource and capping methodologies are described in NovaGold internal documents.


 
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TABLE 3.5
BLOCK MODEL RESOURCE
Resource Tabulation, May 22, 2003 (by NovaGold)

From 0 to 25 Meters From The Nearest Drillhole

Cut-off Tonne Capped Uncapped
Au g/t Ounce Au g/t Ounce
0.00 14,510,987        0.95 441,184        0.98 457,106       
0.25 13,290,612        1.01 433,626        1.05 449,555       
0.50 9,216,847        1.30 385,407        1.35 401,347       
0.75 6,404,476        1.60 329,776        1.68 345,699       
1.00 4,598,103        1.89 279,486        2.00 295,416       
1.25 3,352,986        2.18 234,757        2.33 250,688       
1.50 2,512,318        2.45 197,777        2.64 213,576       
1.75 1,921,972        2.70 167,113        2.96 182,609       
2.00 1,464,847        2.96 139,621        3.28 154,614       

From 0 to 50 Meters From The Nearest Drillhole

Cut-off Tonne Capped Uncapped
Au g/t Ounce Au g/t Ounce
0.00 33,319,227        0.87 933,936        0.91 976,324       
0.25 29,319,386        0.97 909,938        1.01 952,270       
0.50 18,894,174        1.29 786,534        1.36 828,880       
0.75 12,698,901        1.63 664,700        1.73 707,063       
1.00 9,233,591        1.91 568,278        2.06 610,656       
1.25 6,847,779        2.19 482,712        2.39 525,099       
1.50 5,161,276        2.46 408,429        2.72 450,663       
1.75 3,901,017        2.73 342,913        3.06 384,187       
2.00 2,990,852        3.00 288,145        3.41 327,860       

From 0 to 75 Meters From The Nearest Drillhole

Cut-off Tonne Capped Uncapped
Au g/t Ounce Au g/t Ounce
0.00 45,040,143        0.83 1,197,333        0.87 1,260,521       
0.25 38,238,353        0.94 1,157,462        0.99 1,220,626       
0.50 23,441,382        1.30 983,067        1.39 1,046,174       
0.75 15,499,608        1.66 827,284        1.79 890,422       
1.00 11,342,022        1.95 711,681        2.12 774,825       
1.25 8,599,945        2.22 613,349        2.45 676,511       
1.50 6,511,346        2.49 521,242        2.79 584,245       
1.75 4,916,522        2.77 438,482        3.17 500,392       
2.00 3,814,502        3.03 372,169        3.52 431,832       

 
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4  GEOTECHNICAL & PERMAFROST
     
4.1 GEOTECHNICAL DESIGN CRITERIA
     
  4.1.1

Pit Slopes

At this stage of study, the site specific geotechnical data gathered to date is limited to core descriptions, rock quality (RQD) and hardness information from 2002 and 2003 drilling. No geotechnical testing of core to determine shear strength has been carried out and no instrumentation has been installed to confirm permafrost or groundwater conditions. It is expected that this and other information will be gathered during subsequent stages of the study and specific recommendations for further geotechnical work are included in Section 14.

In the design of pit slopes in the relatively competent rock at Rock Creek, failure through the intact rock will generally not be a realistic failure mode. Wall stability will be governed by the presence of discontinuities such as jointing, shearing and faulting, which can act as planes of weakness along which movement can occur. Depending on the orientation and frequency, these discontinuities can also act as release surfaces that allow blocks of rock to separate from the rock mass.

A review of the available RQD data from 45 and 55 degree inclined drill holes around the potential pit limits, clearly shows a marked variation in fracture/joint frequency with hole depth. Figure 4.1 shows the RQD data from holes RKDC02-108,109 and 110 outside the north end of the ore body, RKDC02-106,107 and 111 at the south end and RKDC02-110, RKDC03-117,121 and 123 near the centre of the ore body. All holes were drilled in the same direction (140 degrees azimuth) at dips of 45 or 55 degrees from horizontal.

There is a wide spread of relatively low RQD values (average 10 to 25) through what is assumed to be the main ore bearing shear and tension zones (Albion shear and QMS) down to an inclined hole depth of 24 metres (80 feet) or a vertical depth of about 17 metres, with a marked increase in RQD (average 50 to 70) at an inclined hole depth of 24 to 34 metres (80 to 110 feet) or a vertical depth of about 17 to 24 metres. A second zone of lower RQD (average 15 to 30) is indicated between 34 and 40 metres inclined depth (110 and 130 feet), again increasing below. The zone of increased RQD may represent what may be expected in the pit walls outside the main ore zone, although none of the holes with RQD data appear to be deep enough to have penetrated beyond the QMS. This would explain the second zone of lower RQD below an inclined depth of 34 metres.


 
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Based on the indications of the limited data available on hardness (R3+) and assuming pit walls formed in hard rock with an RQD of around 60, stable inter-ramp pit wall angles in the range of 40 to 50 degrees should be possible for wall heights in the range of 150 metres. If subsequent investigations into the proposed pit walls indicate higher average RQD values of over 75, then wall angles in the upper end of the range may be adopted. However, if lower RQD values of less than 40 are found to be more representative, then wall angles at the lower end of the range will be required. The latter may well apply at shallow depth within the QMS. Along the final pit walls it is possible to use double or triple benching to increase pit wall angles.

At this scoping level of study it is recommended that an inter-ramp pit wall angle of 45 degrees be adopted in design but that sensitivity to the range of 40 to 50 degrees should also be evaluated.

These slope angles assume that any excess pore water pressures below permafrost in the walls are can be controlled or naturally dissipate (see Permafrost Section 4.2). Further investigation work is required to confirm this assumption.

     
  4.1.2

Development Rock Dumps

Development rock dumps located on the valley side slopes to the east and west of the pit will be constructed on regolith that is typically residual soil comprised of silty sand and gravel gradations. No geotechnical investigations have been conducted in these materials to date, but from inspection of core photos (see RKDC02-107 in Figure 4.2), these surficial materials are expected to provide a competent foundation for the dumps.

With maximum side hill slopes at the potential dump sites of about 10 degrees (Figure 4.3), stable rock dump slopes of 2.5H:1V should be possible and stable soil dumps slopes of 3.5H:1V may be assumed in design. Mixed rock and soil dump slopes should be stable at 3H:1V. It is assumed that only nominal thicknesses of organics will be present and that dumps can be developed in lifts of about 5 metres, without the need for any pre-stripping. However, these assumptions need to be confirmed at the next stage of design.

     
  4.1.3

Tailings Impoundment

No geotechnical information is available on the proposed site for the tailings impoundment at the lower end of the Rock Creek valley and no information has yet been gathered on suitable construction materials for the impoundment dykes.


 
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However, from inspection of some typical regolith core, it is likely that impoundment starter dykes may be constructed on natural grade stripped of organics without the need for any special measures to enhance bearing capacity or limit seepage. It will be very important to investigate the full perimeter of the proposed impoundment to confirm that a uniform thickness of low permeability regolith and or alluvium is present and to determine a suitable depth of key to be constructed under the main dykes. It will also be important to determine the depth to bedrock and the nature of jointing and fracturing in the bedrock surface.

At this level of study it is assumed that sufficient fine-grained material will be available from the regolith excavated during the initial pit opening to construct a tailings dyke with an impermeable core or upstream blanket over a shell of development rock. If sufficient fine-grained materials are available, then a homogeneous dyke section may be considered. Figure 4.4 is a conceptual dam cross-section.

Under the ponded surface of the facility it is likely that thawing of the permafrost foundation will occur. For this reason, it will be important to thoroughly investigate permafrost in the area, including ground temperatures and any excess ice conditions.

     
  4.1.4

Infrastructure

Although there is no geotechnical information available on the proposed plant site, foundations for buildings and the various items of plant are not expected to require any special methods of construction other than those dictated by permafrost conditions. For suitable types of foundation in areas of permafrost refer to Section 4.2.3

     
  4.1.5

Recommendations

Fieldwork
Other than some RQD and rock hardness data generated from the 2002 and 2003 core drilling, no geotechnical site investigation has yet been carried out at the proposed mine site. As the project proceeds through the various levels of design, site specific information will need to be gathered to support the basic mine, development rock dump and tailings facility designs, as well as the design of building and plant foundations. The following summarizes recommendations for gathering geotechnical information considered necessary for the next level of study.


 
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Core holes
At least six core holes should be drilled in the proposed walls around the pit. These may be vertical or inclined into the pit at the planned wall angle to maximize representative data over the ultimate height of the wall. All cores should be oriented, photographed, and logged in detail including RQD and hardness, as well as detailed joint descriptions. Representative samples of both weathered and intact rock core should be selected for point load and unconfined compression testing as well as grain size, moisture content and Atterberg Limits as appropriate. At least four core holes should be drilled under the proposed tailings dyke alignment and two at the proposed plant site to characterize the weathered rock strata and at least the upper 10 metres of intact rock. These holes should also be logged and sampled.

Test Pits
A program of test pitting should be planned to investigate typical near surface conditions under the proposed development rock dumps and tailings pond. Ten to twelve pits, excavated by backhoe with a permafrost bucket if necessary, should be dug to at least four metres below ground level. Pits should be inspected and logged by a geotechnical engineer to characterize the materials and any excess ice conditions. Disturbed samples of various horizons should be collected for classification testing in the laboratory.

Instrumentation
At least two piezometers and two thermistor strings should be installed to determine groundwater pressures and depth of permafrost in the pit area and a further two piezometers and thermistor strings in the tailings area. For ease of installation these should be placed in vertical holes.

       
  4.1.6

Geotechnical Assessment

With the benefit of the additional site-specific data recommended above, the next level of study should include the following:

       
    A summary report of all geotechnical data gathered to date
    Structure analysis and rock mass characterization
    Stability analyses of pit wall configurations
    Preliminary catch bench design
    Preliminary assessment of potential pit inflows and dewatering requirements
    Assessment of foundation conditions at potential development rock dump sites
    Stability analysis of development rock dumps

 
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    Assessment of foundation conditions at proposed tailings impoundment site
    Review of available construction materials
    Stability analysis and preliminary design of tailings dykes
    Preliminary seepage analysis of tailings impoundment
    Preliminary design of surface water diversion structures, retention ponds etc.
    Assessment of proposed plant site and preliminary foundation designs
    Preliminary consideration of closure planning, capping and reclamation

 
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     FIGURE 4.1
PIT WALL RQD DATA

Rock Creek - RQD Data

All holes drilled 140 deg. az., with 45 or 55 deg dip from horiz.

 
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FIGURE 4.2
OVERBURDEN & REGOLITH CORE PHOTO

 
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FIGURE 4.3
TOPOGRAPHY IN WASTE DUMP AREAS

 
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FIGURE 4.4
CONCEPTUAL TAILINGS DAM DESIGN

 
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4.2 PERMAFROST
     
  4.2.1

Introduction

Permafrost is defined as a state of ground that remains at or below a temperature of 0ºC for a minimum period of at least two consecutive years. No distinction is made in the permafrost definition for whether the frozen ground is rock or soil as long as the temperature regime meets the criteria. Permafrost develops in areas where the heat loss from the ground during winter exceeds the combined energy gain during the summer and energy radiated by the local geothermal gradient (i.e. heat radiating upwards from depth). Consequently the thickness of permafrost and actual permafrost temperatures can vary locally, depending on air temperatures, the geothermal gradient, snow cover, exposure direction and topographic conditions. Typically north facing exposures will have greater permafrost depths than south facing exposures. Areas overlain with organic soils and snow cover (i.e. insulators) will have less permafrost than barren rock exposures.

Average annual air temperatures in Nome are about -3°C; seasonal fluctuations are shown in Figure 2.2. Published map data indicates that the Rock Creek Project site is located near the regional boundary between continuous and discontinuous permafrost.

To the best of our knowledge no thermistors or permafrost measurements have been made at the Rock Creek site. Published data indicates that permafrost depths in Nome could be approximately 100 metres. The lack of trees and the presence of frozen placer gravel deposits are indicators of the permafrost regime. In the NovaGold report, "1999 Rock Creek Drilling & Metallurgical Program", mention is made on Page 15 that some of the drilling at Rock Creek is shallow and "cuts off at the approximate depth of the water table" at about 60 metres. This could be an indicator of the permafrost depth in the mine area rather than the actual water table.

Within the permafrost horizon, a near surface zone of the permafrost will thaw and re-freeze on an annual basis as seasonal air temperatures change. This upper zone is termed the “active layer” and could be about 2 to 3 metres at the Rock Creek site. In hard rock, the active layer will be deeper due to a higher thermal conductivity, and could be closer to 5 to 6 metres.

At a certain depth below the ground surface, the seasonal air temperature fluctuations become attenuated and the ground mass essentially remains at a


 
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constant temperature throughout the year. This depth is termed the “depth to zero annual amplitude”. Typical geothermal gradients can be about +1ºC per 100-metre depth, but have not been measured at the site.
     
  4.2.2

Engineering Properties of Permafrost

A key characteristic that impacts on the engineering properties of permafrost is that the majority of the groundwater or soil moisture is frozen. At temperatures a few degrees below freezing, a small amount of water remains unfrozen, bound to silt and clay sized particles in the soil matrix. The presence of ice in the pores of a soil or in rock fractures tends to increase the shear strength of the material. It also reduces the hydraulic permeability such that the material can be considered essentially impervious. The presence of interstitial ice, particularly ice lenses, causes the soil to exhibit creep behaviour under loading. This means the material may very slowly strain or deform under loading. The colder the permafrost temperature is, the higher the developed shear strength will be and the lower the creep rate.

Permafrost soils can be classified as “thaw-stable” or “thaw-unstable”. Thaw unstable permafrost loses a significant amount of strength upon thawing, due to presence of ice and excess water when melted. Thaw stable permafrost is generally ice-poor and will not settle appreciably when thawed. Generally more conventional type foundation designs can be used in thaw-stable permafrost while special permafrost protection measures may be required in the thaw-unstable areas. Where bedrock is near surface, it maybe best to seat the footing in the bedrock and excavate frozen soils entirely.

Where engineering designs purposely utilize the favorable strength or permeability characteristics of the permafrost, it is imperative that the cold temperature regime is maintained in order to retain the desired engineering properties. Construction activity can warm or degrade the permafrost in several ways. For example, the removal of layers of insulating organic vegetation, the ponding of water due to changes in topography, or heat radiating from buildings can change the temperature regime at the ground surface and cause permafrost thaw. This thawing could then lead to a reduction in shear strength or even melting of in-situ ice lenses, creating ground settlement or seepage pathways. Settlement could then lead to additional water ponding, progressively degrading the permafrost. The thawing of ice filled rock fissures can create potential water seepage pathways if incorporated in a water retaining structure. The convective heat transfer caused by the flowing water may result in even further permafrost degradation in such cases.


 
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  4.2.3

Engineering Design in Permafrost

Properly engineered designs can make use of the favourable characteristics of permafrost as long as the designs incorporate permafrost protection measures.

The engineering design process for the facilities and structures required at the Rock Creek Project should take into consideration the presence of permafrost at the site. In some cases, the existence of permafrost at the site is an advantage in that it has favourable geotechnical and hydrogeological properties that will be incorporated into operation and closure planning. In other instances, the permafrost may be a disadvantage in that it can be ice-rich and create settlement problems if thawed and possibly creep while still frozen. There is no single way to deal with permafrost and therefore each of the facilities may address the permafrost in a different manner, depending on:

       
    the depth and temperature of the permafrost at the specific location
    the type of permafrost (ice-rich or ice-poor soil, or bedrock)
    the type, size, and use of structure being constructed
    the sensitivity of the structure to settlement deformation
    the long-term closure plan intended for the structure
     
   

The following discussions are intended to be general in nature, describing the overall design concepts. The approaches described here may differ slightly from the actual designs implemented; however the overall permafrost protection objectives will be the same. Where applicable, references to the relevant engineering design reports are provided for more detailed descriptions and analyses.

Roads and Embankments
The site access roads and haul roads will be constructed over both exposed bedrock and frozen overburden. The bedrock will provide a competent sub-grade foundation whether frozen or thawed; therefore no special permafrost protection measures will be utilized in these areas. Most other areas at site only contain a 1 to 2 metre thick layer of overburden and therefore thaw induced settlement should be limited.

Site Drainage System
The site drainage system will consist of a network of interceptor ditches and settling ponds. For the most part the drainage plan should avoid the excavation of ditches in thick overburden. Thawing of the soil can lead to erosion and slumping


 
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of the side slopes. Consequently, constructed embankments and roads can be used to provide runoff containment, where practical. However in order to maintain proper drainage grades, it may be necessary in some locations to excavate ditches. Armouring of ditch slopes in thicker overburden can help to maintain stable slopes and will prevent till erosion during periods of water flow.

In some places it may be necessary to excavate ditches in bedrock and no permafrost protection measures will be required.

Where settling ponds are required small, low-height dams may be needed. These dams should be of sufficient height and dimension that they eventually freeze. The freeboard should be at least 2 metres above the maximum water level to be contained so that active layer in the dam remains above water level.

Small Structure Foundations
Small structures with low ground bearing pressures may be constructed with several alternate footing designs. The selection of the specific foundation design will depend on the depth to bedrock at a given site and the type of overburden cover present.

Where bedrock is at or near surface, shallow piles socketed directly into the bedrock can be used to support the structure. Alternatively a shallow thickness of frozen overburden can be sub-excavated down to bedrock and then replaced with compacted fill. Slabs on grade can also be utilized where the bedrock is at surface.

Where bedrock is at depth other means of support are available, although we expect such measures will likely not be needed at the Rock Creek site. Adfreeze piles frozen into the permafrost are one option available. These piles are placed to depths sufficient to resist annual freeze/thaw uplift forces and also support the loads imposed by the structure. A layer of compacted fill can be used to support the structure on top of the frozen soil and insulate the sub-grade. It may also be possible to elevate the structure off the ground surface, providing a ventilated pad beneath such that cold winter air can assist in keeping the ground frozen. Consideration will also be given to the incorporation of synthetic insulation in the sub-base design to reduce the thickness of fill required

Large Structure Foundations
Large structures, such as the mill building, truck shop, and diesel storage tanks should be constructed directly on bedrock foundations. Localized pockets of overburden and regolith (i.e. weathered bedrock) should be excavated down to


 
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competent rock, where necessary. Heat radiating from these structures will likely cause some thaw of the underlying permafrost however since it will consist of bedrock, no settlement is expected. No special permafrost protection measures are envisioned in these areas.

Development Rock Dumps
The development rock dumps will be used for disposal of rock stripped during the mining operations. The rock piles will be constructed over various permafrost foundation conditions including bedrock, with some overburden veneer. No special construction methods are envisioned for the development rock dumps, although if geotechnical investigations indicate the presence of ice-rich soils, then some provisions may be required to minimize slumping of dump slopes on thawing soils.

Over time the rock dumps will freeze into permafrost, however the large pore spaces will probably still allow seepage through the dumps. Studies have shown that sulphide oxidation and acid generation rates may be reduced in frozen rock but would likely still continue.

Tailings Management Facility
The tailings storage facility will consist of sub-aqueous deposit tailings covered by a water cap. The downstream slopes of the tailings dams will likely freeze over time.

Some thaw of the permafrost bedrock underlying the interior of the pond is likely. The thaw zone should neither extend laterally through the dams or downwards very deep. Therefore some rock fractures at depth should remain frozen and ice-filled, maintaining an impermeable barrier to subsurface seepage.

Upon project closure, once the ponded tailings water is discharged the exposed tailings beaches should begin to freeze, eventually re-establishing the permafrost regime within the impoundment. Most likely the entire tailings area will become permafrost except for the upper 2 to 3 metre active layer.

Mining
The Rock Creek pit areas are expected to be contained within permafrost rock. The geotechnical design of the pit slopes may incorporate rock shear strength increases due to permafrost. As well seepage and pore pressure conditions should take into the presence of permafrost. Where present in the pit wall, the permafrost zones are considered impermeable rock.


 
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At depth however the pits may enter the zone of thawed rock. Pore pressures with excessive heads may be encountered, resulting in some slope stability problems. Water quality of the seepage water should be also be evaluated. Seepage faces can lead to safety hazards associated with pressure build-up behind frozen faces and the formation of icings along the bench faces. The Usibelli coal mine in Alaska historically had some concern with smaller scale pit wall blow-outs caused by pore pressure build up behind frozen walls.
     
  4.2.4

Impact of Global Warming

Global climate change is difficult to quantify in both magnitude and rate, particularly if projections must be made far into the future (i.e. beyond 50 years). The Rock Creek Project has an expected life of about 6 to 9 years at which point most of the infrastructure will be dismantled and removed from site. A few of the structures however, will remain at the site and may have incorporated freeze-back into their closure planning. Specifically, the tailings and development rock dumps are the two main facilities that fall into this category.

Where critical, the thermal performance of permanent structures should be evaluated against selected global warming scenarios. Some researchers have proposed global warming temperature increases in the range of 0.3ºC to 0.5ºC per decade. This equates to a temperature increases of 3 to 5ºC over the next 100 years. Given that the annual air temperature at the Rock Creek site is about -3ºC, this level of warming would be sufficient to thaw all of the permafrost.

It is important to note that air temperature change is not the only parameter that enters into the ground surface energy balance. The presence of snow cover will act as an insulating layer and therefore it also plays a significant role. If a corresponding reduction in snow cover took place with climate change, a lessened ground-warming trend could still be seen in spite of climate warming.

     
4.3

GROUNDWATER ISSUES

No hydrogeological field investigations have been undertaken at the Rock Creek site, however it is envisioned that groundwater flow will likely occur in two horizons. They are:

     
  i.
The seasonal active layer will permit groundwater flow during summer conditions. However due to its limited thickness, flow quantities will be small particularly in the

 
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finer grained soils. Sand and gravel layers may allow more seepage and therefore must be mitigated where sub-surface seepage is a concern.
     
  ii.
The more significant hydrogeological unit may be the thawed rock beneath the permafrost zone. The well-fractured shear vein could potentially be a significant source of water. The depth of permafrost could be in the range of 60 to 100 metres, and therefore mining may not encounter groundwater until later in the mine life.
     
 
The need for perimeter pumping wells is unknown at this time and has not been included in the cost estimate. It is assumed that some surface runoff and pit wall seepage may occur at depth in the pit. This water would be collected in in-pit sumps and pumped to the tailings facility.
     
4.4

SITE WATER MANAGEMENT

Surface runoff from undisturbed areas outside the project footprint will be collected in interceptor ditches and diverted around the project area. Settling or polishing ponds will be used to remove suspended solids prior to discharge to the environment. Figure 6.1 shows the location of the interceptor ditches. These settling ponds will also be equipped with submersible pumps, which can be used to withdraw water for process use.

Surface runoff from disturbed areas inside the project footprint will be collected in mine water ditches and storm water sumps. These sumps will be equipped with submersible pumps, which will discharge to the tailings management facility.

Runoff from the downstream slope of the tailings dam and dam seepage will be collected in a toe ditch and return to the tailings pond. No hydrological studies or detailed sizing of ditches and sumps has been undertaken at this time.

The pits are expected to be largely contained within permafrost and therefore minimal groundwater seepage is anticipated until mining at depth. However surface runoff and pit seepage will be pumped out of the pit into the mine water ditches for re-use in the process. Diesel power pumps will be used for mine dewatering.


 
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5

ENVIRONMENTAL & PERMITTING

The proposed mine development and operation would have to obtain numerous State and Federal permits before development and operation can proceed. Detail descriptions of the environmental and permitting aspects of the project are included in Appendix F including:

     
  Location, climate and environmental settings in the project area
  Project description
  Required permits
  Potential environmental issues
  Recommended baseline investigations
  Permitting schedule and budget summary
     
5.1
  

INTRODUCTION

The Rock Creek gold prospect is bounded on the north and west by Mount Brynteson, to the east by the Snake River and on the south by Glacier Creek. The elevations on the project area vary from between 30 masl (100 ft above sea level) and 200 masl (650 feet). The property is located within the Bering Straits Resource Area Coastal Management District.

Vegetation in the Nome area consists mostly of tundra mat, mosses, and lichens.

Temperatures generally remain well below freezing from the middle of November to the latter part of April, with January usually the coldest month of the year. Temperatures usually begin to rise near the end of February and continue to rise until they reach a maximum in July. Precipitation reaches its maximum during the late summer months and drops to a minimum in April and May. Snow begins to fall in September, but usually does not accumulate on the ground until the first part of November.

   
5.2

PERMITTING REQUIREMENTS

Various Federal and State agencies have jurisdiction over many of the activities proposed for the Rock Creek Project. The types of permits and approvals, which may be needed in order to construct and operate the Rock Creek Project, will depend on the design and operation of the mine. For example, the Rock Creek Project intends to be a zero discharge facility and therefore will likely not require a NPDES permit however a description of the permit is included in the listing for completeness purposes.

The major permits, approvals or plans that may be required for the Rock Creek Project are listed below and described individually in Appendix F. Included with those descriptions

 
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are the scope and types of studies required and typical time frames for project permit review and approval. It is assumed that all project activities will occur on private or state lands and that no federal land use or rights-of-way will be required.

Federal Authority

       
   
Environmental Protection Agency
       
   
Clean Water Act Section 402 National Pollutant Discharge Elimination System (NPDES) Permit For Wastewater Discharge
   
Spill Prevention, Control, And Countermeasure (SPCC) Plan
   
Storm Water Construction And Operation Permit
   
Pollution Prevention Plan
       
    Corps of Engineer
   
Clean Water Act Section 404 Wetlands Permit
       
   
U.S. Fish and Wildlife Service
   
Threaten and Endangered Species Clearance
     
    Federal Communications Commission (FCC)
   
Radio and microwave station authorizations
     
    Treasury Department (Department of Alcohol, Tobacco and Firearms)
   
license for transport and use of explosives.
     
    U.S. Mine Safety and Health Administration
   
obtain an MSHA number, safety training, record keeping and reporting
       
  State Authority
       
    Department of Natural Resource
   
Plan of operations approval
   
Temporary water use permit
   
Upland mining lease
   
Rights of way
   
Mill site Lease
   
Material sale
   
Certificate of Approval to Construct Dam
   
Certificate of Approval to Operate a Dam

 
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    Permit to Appropriate Water
    Material Sale Permit
    Burn Permit
    Cultural Resources Authorization
    Mining License
       
    Department of Environmental Conservation
    Air quality control permit to operate
   
Clean Water Act Section 401 certificate of reasonable assurance for EPA NPDES and COE Section 404 permits
    Wastewater disposal permit
    Solid waste disposal permit
    Plan reviews for public water supply and domestic sewage systems
    Storm Water Discharge Pollution Prevention Plan
    Oil discharge prevention and contingency plan
    Approval to Operate a Public Water System
       
    Department of Fish and Game
    Fish Passage Authorization
    Title 16 Fish habitat permit
   
5.3

BASELINE STUDIES

Environmental baseline studies are required in support of the preparation of a number State and Federal permit applications. Baseline information and data regarding surface water, groundwater, tailings and development rock geochemistry soils, vegetation, fish and wildlife, and threatened/endangered species, are necessary for design of the mining (i.e., extraction) and reclamation plan that must be prepared in support of the Plan of Operations application and the Solid Waste permit application.

The baseline studies listed below are considered adequate to support an EA. If an EIS is required, additional and/or more rigorous studies will be required. These requirements would be identified in the EIS scoping exercise. NovaGold should consult early in the permitting process with the various agencies responsible for permitting the operation in order to develop acceptable baseline study plans.

   
    Surface water
    Groundwater
    Tailings and Development Rock Geochemistry

 
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    Soils
    Vegetation
    Air Quality
    Wildlife
    Archaeological Resources
    Threatened and Endangered Species
       
5.4

POTENTIAL ISSUES

There may be other potential project related issues that may have to be addressed. While although there is no regulatory requirement for detailed analysis within the ADNR or ADEC, permitting for these issue they would have to be analyzed if an EIS is required.

These issues could include:

     
 
Socio-Economics: Impacts and benefits to the local community and infrastructure (employment, schools, hospitals, housing and power supply).
  Transportation: Issues related to road safety and road maintenance and improvements.
  Safety: Issues related to storage and transport of materials and fuel from Nome to the Rock Creek site.
  Noise impacts on local residents
  Visual impacts
   
5.5

PERMITTING SCHEDULE AND BUDGET

A preliminary scheduling Gantt chart for the permitting of the Rock Creek Project is included in the project Execution Schedule in Section 15 and described in detail in Appendix F. This schedule requires that all necessary engineering designs and specifications needed for permitting will be available on a timely basis. During permitting, it is not uncommon for schedules to require revision because of delays in the Federal and State agencies review of environmental baseline data and permit applications. Projects which are controversial and meet with public opposition can require significantly more time than the two year timeframe presented in this schedule.


    Baseline data collection: 3rd Q of 2003 to 3rd Q of 2004
  Permit application: 3rd Q of 2004 to 3rd Q of 2005
  Project construction start: 3rd Q of 2005
     
 
The estimated budget cost for each aspect of the anticipated permitting for the Rock Creek Project is about $1.1 million, as detailed in Appendix F.

 
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6

OPTIMIZATION STUDIES

Prior to finalizing the project design, several optimization studies were undertaken to assist in siting the various facilities. The main goal of the siting optimization is to keep all of the project facilities within the same watershed disturbance area if possible. This would simply water management and minimize the potential for off-site disturbances.

The siting optimization was undertaken using a Lerchs-Grossmann optimized pit shell developed using the 1999 geological block model since the 2003 model was unavailable at that time. It was expected that the 2003 model would not drastically change the pit limits or impact on the conclusions of these evaluations.

     
6.1

PLANT SITE LOCATION

Based on the location of the Rock Creek ore body and the local topography, only one location was considered for siting the mill and administration facilities. The selected area is to the east of the Glacier Creek Road and north of Rock Creek as shown in Figure 6.1. This site has gently sloping terrain (about 1-2%) and shallow overburden cover overlying the bedrock. Condemnation drilling should be done to confirm that no mineable ore deposits are beneath the plant area.

     
6.2
  

TAILINGS DISPOSAL OPTIONS

A series of potential tailings facility sites were evaluated to determine the potential tailings storage space available and the efficiencies that different configuration could provide. The different concepts are shown in Figure 6.2.

The initial guidelines used in developing these designs are:

     
  Be confined within the watershed draining towards Rock Creek
  Have capacity for 3.9 million m3 (5 Mt ore at 1.3 t/m3 tailings dry density)
  Centerline dam construction
  4:1 downstream dam slope
  2:1 upstream dam slope
  2% slope on tailings deposit beaches
  10 metre final dam crest width
  Maintain 100m setback of dam toe from roads, mill, pit crest, and other infrastructure.
   
 
Table 6.1 summaries the storage capacity and dimensions for each of the tailings facilities shown in Figure 6.2.

 
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TABLE 6.1
TAILINGS FACILITY OPTIONS

 
Option 1

Option 2
Option 3
Option 4
Tailings Volume (1000 x m3) 3,900 3,900 3,900 3,900
Dam Volume (1000 x m3) 3,764 3,663 3,701 3,793
Dam Crest Elevation (masl) 62 71 69 69
Tailings Footprint Area (ha) 119 68 64 72
Max Dam Height (m) 32 38 37 37
Average Tailings Depth (m) 6 10 11 9
Water Volume (1000 x m3) 2,617 1,443 1,329 1,837

 

Option 4 was selected as the preferred tailings facility option based on trade-of between footprint area and dam height, with footprint minimization be the driving factor. Option 4 results in a southward shift, providing more space for the mill site yet still locating the pond over the Rock Creek gulley. The Rock Creek channel provides a convenient place to locate the process water reclaim barge.

Once the final mine plan was developed, the tailings tonnage increased to about 11 million tonnes (8.5 million m3). The selected location was able to accommodate the added tailings by raising the dams higher by about 7 metres.

Condemnation drilling must be done to discount the presence of any mineable ore zones in this area and to evaluate the extent of the mineralization to the southwest of the main ore zone.

   
6.3

PIT OPTIMIZATION (LERCHS-GROSSMANN)

Prior to developing the pit designs and mine layouts, the Lerchs-Grossmann (LG) pit optimizer was used to evaluate pit geometries based on variable economic factors. Each variation in revenue or cost would result in a different optimized pit. As a first pass, the pit optimizer is run using estimated mining and processing costs. Once the final mine plan and production schedule is developed, more refined estimated costs can be developed. If these costs were significantly different than those used in the pit optimization, the Lerchs-Grossmann would be repeated with the new costs. As will be discussed later, the final operating costs were not significantly than those used herein and the optimization work was not repeated. The "capped" block grades were used for optimization purposes.


 
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Table 6.2 summarizes the parameters (under the column labelled Base Case Values) used in the pit optimizations. In addition a series of sensitivities were run on different factors, also summarized in Table 6.2.

The base case pit slope assumed that the west wall would have an overall slope of 30° due to the likely presence of haul roads along this wall of the pit. The east pit wall was varied from 45° to 50°. The case of a 40° pit slope around the entire pit was also modeled to simulate a spiral ramp.

TABLE 6.2
PIT OPTIMIZATION INPUT PARAMETERS

  Base Case
Values
Sensitivity
Mining Cost, $ per tonne mined $0.92 not varied
Milling Cost, $ per tonne milled (M) $6.30 Base MAS = $7.50/t but
varied from $5.00 to $10.00/t
Admin, $ per tonne milled (A) $1.00
Sust capital & reclam, $ per tonne milled (S) $0.20
Gold Prices, US$/oz $325 $200 to $400
Tension vein recovery, % 97% varied from 70% to 95%
Albion vein recovery, % 94% varied from 60% to 95%
Pit Slopes* (overall), degrees 30° and 45° 40° and 50°
* west wall flatter than east wall.

 
The pit optimizer indicates that depending on the gold price, either two separate mining areas or one contiguous mining area is apparent (Figure 6.4). The results for the entire optimized pit resource and sensitivities are shown in Table 6.3 at variable cut-off grades. The resources at a fixed cut-off grade of 1.0 g/t are shown in Table 6.4. The Base Case optimized pit resource calculation is presented in Table 6.5. About 78% of the in-pit gold is contained in and recovered from the tension vein ore.

 
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TABLE 6.5
PIT OPTIMIZED RESOURCE

Gold Price ($/oz) $325 $325
Cut-off Grade (g/t) 0.85 1.00
Ore tonnes 12,851,000 10,997,000
Ore Grade (g/t) capped 1.869 2.029
Waste tonnes 45,056,000 46,910,000
Waste:Ore ratio 3.51:1 4.27:1
Gold, block model ounces 772,300 717,500
Gold, recovered ounces 744,400 691,200
Average recovery 96% 96%

 

Gold Price Sensitivity
Table 6.3 and Figure 6.3 describe the impact of the recoverable gold resource with a variable gold price. A significant sudden increase in recoverable gold occurs at a price of $275/oz, which is the point at which the two smaller pits coalesce into one larger pit. Beyond that point, the recovered gold increases gradually, at a rate of about 10,000 ounces for each $25 increase in the gold price.

Milling & Admin & Sustaining Capital (MAS costs) cost sensitivity
The combined milling, administration, and sustaining costs (MAS costs) were varied above and below the base case cost of $7.50/t. The results, shown in Table 6.3, indicate that the recoverable gold would increase by about 28,000 ounces for each $0.50/t decrease in MAS cost. The milling cost sensitivity is therefore not as great as the gold price impact.

Albion vein recovery sensitivity
The Albion vein recovery was varied downward from base case value of 94%. The results, shown in Table 6.3, indicate that the recoverable gold would decrease by about 12,000 ounces for each 5% decrease in recovery. The relatively low impact of the Albion ore recovery is due to the fact that only 22% of the block model gold is contained in the Albion zone.

Tension vein recovery sensitivity
The Tension vein recovery was varied downward from base case value of 97%. The results, shown in Table 6.3, indicate that the recoverable gold would decrease by about 50,000 ounces for each 5% decrease in recovery. The large impact of the Tension ore recovery is due to the fact that most of the block model gold is contained in the Tension zone.


 
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Pit slope sensitivity
The overall pit slopes were varied from base case value of 40° which is predicated on a 45° inter-ramp angle and the inclusion of haul roads. The results, shown in Table 6.3, indicate that the recoverable gold would increase by less than 10,000 ounces by steepening the pit slope from 45° to 50°.

Conclusion
Based on the foregoing sensitivities, it was decided to develop a pit design using the $325/oz LG pit as the starting point. This shell encompasses the single large pit configuration and should yield between 650,000 and 700,000 ounces of recoverable gold. The incorporation of ramps into the pit design will alter the ore and waste volumes somewhat from those estimated in the optimized pit shell.

FIGURE 6.3
OPTIMIZED PIT RESOURCE SENSITIVITY TO GOLD PRICE
(Cut-Off Grade =1.0 g/t)

 
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TABLE 6.3
PIT OPTIMIZATION SUMMARY (Variable Cut-off Grade)
LG Shell Sensitivity Resource Total Waste Strip Ratio Ore - Total of all Zones Gold Insitu Gold Recovered
    Cut-off Grade Material       AUK AUKU Tension Albion Total Tension Albion Total
    g / t k-tonnes k-tonnes t.wst / t.ore k-tonnes g / t g / t '000 oz '000 oz '000 oz '000 oz '000 oz '000 oz
                             
BaseCase   0.85 57,907 45,056 3.51 12,851 1.869 2.037 614.4 157.9 772.3 596.0 148.5 744.4
                             
Au Price Sensitivity                        
   Au-200 200 1.39 14,492 10,718 2.84 3,774 2.610 3.026 225.4 91.3 316.7 218.6 85.8 304.5
   Au-225 225 1.23 19,057 14,038 2.80 5,020 2.394 2.733 282.0 104.4 386.4 273.5 98.1 371.7
   Au-250 250 1.11 25,592 19,008 2.89 6,584 2.210 2.479 345.5 122.3 467.8 335.1 115.0 450.1
   Au-275 275 1.00 49,396 39,515 4.00 9,881 2.082 2.297 508.5 153.0 661.5 493.3 143.8 637.1
   Au-300 300 0.92 54,359 42,870 3.73 11,489 1.962 2.146 569.1 155.5 724.6 552.0 146.2 698.2
   Au-325 325 0.85 57,907 45,056 3.51 12,851 1.869 2.037 614.4 157.9 772.3 596.0 148.5 744.4
   Au-350 350 0.79 60,653 46,454 3.27 14,199 1.783 1.936 654.8 159.2 813.9 635.1 149.6 784.7
   Au-375 375 0.73 61,976 46,603 3.03 15,374 1.710 1.851 685.9 159.4 845.3 665.3 149.8 815.1
   Au-400 400 0.69 63,065 46,521 2.81 16,543 1.643 1.775 714.5 159.6 874.1 693.0 150.1 843.1
MAS Cost Sensitivity                        
   Mill-5.00 5.00 0.59 63,810 44,692 2.34 19,118 1.510 1.623 768.1 159.8 927.9 745.1 150.2 895.3
   Mill-5.50 5.50 0.64 62,322 44,901 2.58 17,421 1.591 1.715 731.4 159.5 891.0 709.5 150.0 859.5
   Mill-6.00 6.00 0.70 61,383 45,303 2.82 16,080 1.664 1.798 700.8 159.3 860.1 679.8 149.7 829.5
   Mill-6.50 6.50 0.75 60,385 45,474 3.05 14,911 1.733 1.878 671.8 158.7 830.6 651.7 149.2 800.9
   Mill-7.00 7.00 0.80 59,105 45,265 3.27 13,840 1.800 1.956 642.7 158.3 801.0 623.4 148.8 772.2
   Mill-7.50 7.50 0.85 57,906 45,056 3.51 12,850 1.869 2.038 614.4 157.9 772.3 595.9 148.5 744.4
   Mill-8.00 8.00 0.90 56,842 44,797 3.72 12,045 1.930 2.109 590.3 157.0 747.3 572.6 147.6 720.2
    Mill-8.50 8.50 0.95 55,165 43,979 3.93 11,185 1.996 2.189 561.5 156.4 717.8 544.6 147.0 691.6
   Mill-9.00 9.00 1.00 53,178 42,738 4.09 10,439 2.055 2.258 535.0 154.6 689.6 519.0 145.3 664.3
   Mill-9.50 9.50 1.05 51,014 41,349 4.28 9,665 2.120 2.340 505.6 153.3 658.9 490.5 144.1 634.6
   Mill-10.00 10.00 1.10 49,709 40,657 4.49 9,052 2.181 2.415 482.2 152.5 634.7 467.7 143.4 611.1
Mill Recovery (Albion) Sensitivity                        
   Rec-Alb-60 60 0.91 56,373 44,522 3.76 11,851 1.939 2.114 586.5 152.5 739.0 568.9 91.5 660.4
   Rec-Alb-65 65 0.90 56,741 44,727 3.72 12,013 1.929 2.105 591.3 153.7 745.0 573.6 99.9 673.5
   Rec-Alb-70 70 0.89 56,893 44,726 3.68 12,167 1.918 2.092 595.6 154.5 750.1 577.7 108.2 685.9
   Rec-Alb-75 75 0.88 56,915 44,612 3.63 12,303 1.906 2.079 599.4 154.7 754.1 581.4 116.1 697.4
   Rec-Alb-80 80 0.87 57,059 44,606 3.58 12,453 1.895 2.066 603.3 155.5 758.8 585.2 124.4 709.6
   Rec-Alb-85 85 0.86 57,103 44,509 3.53 12,594 1.884 2.053 607.1 155.8 762.9 588.9 132.4 721.3
   Rec-Alb-90 90 0.85 57,630 44,898 3.53 12,732 1.877 2.047 611.3 157.0 768.3 592.9 141.3 734.3
   Rec-Alb-95 95 0.84 57,906 45,031 3.50 12,875 1.867 2.035 615.0 157.9 773.0 596.6 150.0 746.6
Mill Recovery (Tension) Sensitivity                        
   Rec-Ten-70 70 1.07 39,020 31,410 4.13 7,610 2.253 2.528 401.7 149.5 551.2 281.2 140.5 421.7
   Rec-Ten-75 75 1.02 46,608 37,384 4.05 9,224 2.129 2.358 478.3 152.9 631.3 358.8 143.8 502.5
   Rec-Ten-80 80 0.98 49,333 39,285 3.91 10,048 2.063 2.275 512.8 153.8 666.6 410.3 144.5 554.8
   Rec-Ten-85 85 0.94 51,736 40,906 3.78 10,830 2.004 2.200 542.9 154.9 697.8 461.5 145.6 607.1
   Rec-Ten-90 90 0.90 55,013 43,228 3.67 11,785 1.940 2.123 578.1 156.9 735.0 520.3 147.5 667.8
   Rec-Ten-95 95 0.86 57,184 44,635 3.56 12,549 1.890 2.062 604.5 157.8 762.4 574.3 148.4 722.7
Wall Slope Sensitivity                        
   Slope-45d 30d W / 45d E 0.85 57,906 45,056 3.51 12,850 1.869 2.038 614.4 157.9 772.3 595.9 148.5 744.4
   Slope-50d 30d W / 50d E 0.85 53,391 40,419 3.12 12,972 1.869 2.036 621.1 158.3 779.4 602.5 148.8 751.3
   Slope-40d 40 0.85 59,560 46,774 3.66 12,786 1.872 2.041 611.4 157.9 769.4 593.1 148.4 741.5

 
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TABLE 6.4
PIT OPTIMIZATION SUMMARY (Cut-off Grade = 1.0 g/t)
LG Shell Sensitivity Resource Total Waste Strip Ratio Ore - Total of all Zones Gold Insitu Gold Recovered
    Cut-off Grade Material       AUK AUKU Tension Albion Total Tension Albion Total
    g / t k-tonnes k-tonnes t.wst / t.ore k-tonnes g / t g / t '000 oz '000 oz '000 oz '000 oz '000 oz '000 oz
                             
BaseCase 325 0.85 57,907 45,056 3.51 12,851 1.869 2.037 614.4 157.9 772.3 596.0 148.5 744.4
Au Price Sensitivity                      
   Au-200 200 1.00 14,492 9,302 1.79 5,190 2.219 2.522 276.1 94.2 370.3 267.9 88.5 356.4
   Au-225 225 1.00 19,057 12,924 2.11 6,133 2.161 2.438 319.7 106.5 426.2 310.1 100.1 410.2
   Au-250 250 1.00 25,592 18,340 2.53 7,252 2.103 2.348 367.5 122.9 490.4 356.5 115.5 472.0
   Au-275 275 1.00 49,396 39,491 3.99 9,905 2.080 2.294 509.3 153.0 662.3 494.0 143.8 637.8
   Au-300 300 1.00 54,359 43,727 4.11 10,632 2.043 2.242 543.1 155.1 698.2 526.8 145.8 672.6
   Au-325 325 1.00 57,907 46,910 4.27 10,997 2.029 2.226 560.2 157.3 717.5 543.4 147.9 691.2
   Au-350 350 1.00 60,653 49,389 4.38 11,263 2.017 2.209 571.8 158.5 730.3 554.6 149.0 703.6
   Au-375 375 1.00 61,976 50,614 4.45 11,362 2.012 2.203 576.3 158.7 735.0 559.0 149.2 708.2
   Au-400 400 1.00 63,065 51,623 4.51 11,441 2.008 2.197 579.6 158.9 738.5 562.2 149.4 711.6

 
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7

MINE PLAN

The mine design criteria are summarized in the Project Design Criteria (PDC) document, included in Appendix A. This includes data on pit slope criteria, dilution factors, rock density, road dimensions, and other data used for design or costing purposes. Haul road widths of 22 metres and inter-ramp wall angles of 45‹were used for pit design.

   
7.1

PIT DESIGN

As shown by the Lerchs-Grossmann pit optimizer described in Section 6.3, a single large pit area was identified. However in order to smooth waste stripping volumes, a three-phased pit development approach is required. The intent is that the initial mining area would be contained within a low strip ratio area to provide early access to ore. Waste stripping from subsequent phases would be advance only to allow smoothing of waste volumes and avoid extremely high stripping volumes in isolated years.

Figure 7.1 describes the ultimate pit layout. The deepest mining elevation is -30m in the south area while the maximum crest elevation is about +150m. Pit wall heights range around 125 metres in height.

The mineral resource shown in Table 7.1 is unclassified and therefore should be considered as an inferred resource.


 
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TABLE 7.1
INFERRED PIT RESOURCE (CUT-OFF = 1.0 g/t)

      Phase 1 Phase 2 Phase 3 Total
         
Cut-off Grade (g/t) 1.0 1.0 1.0 1.0
Ore tonnes 1,822,000 6,995,000 1,878,000 10,695,000
Ore Grade (g/t) capped 2.189 1.763 2.892 2.034
Waste tonnes 1,366,000 29,113,000 17,006,000 47,485,000
Waste:Ore ratio 0.75:1 4.16:1 9.05:1 4.44:1
Gold, block model (ounces x 1000) 128.2 396.6 174.6 699.5
Gold, recovered (ounces x 1000) 123.8 382.6 167.6 674.0
Added ounces at milling cut-off
Cut-off Grade Increment (g/t) 0.85 to 1.0
Gold, block model (ounces x 1000) 52.5
Gold, recovered (ounces x 1000) 50.9

7.2

PRODUCTION SCHEDULE AND SEQUENCE

Production scheduling and mining equipment requirements are based on an assumed project operating schedule of two12-hour shifts per day for 360 days per year.

Based on the 4.4:1 waste to ore ratio, it is expected that the mining operations will either focus on ore or waste mining operations on a shift-by-shift basis. Probably four out of every five shifts would be assigned to waste stripping operations. During the periods when the fleet is stripping waste, the mill will continue to operate by reclaiming ore from the live ore stockpile. Ore mining rates are about 1.78 million tonnes per year while waste volumes range from 6.8 to 9.2 million tonnes per year. Total material in any given year peaks at 10.98 million tonnes in Year 3. At the feasibility stage, more detailed planning should be able to further smooth waste volumes.

Table 7.2 summarizes the mine production schedule while Table 7.3 provides more detail as to when each phase is mined.

Phase 1 is largely mined in Year 1 and part way through Year 2. Some waste is mined from all three phases in Year 1 mainly to provide construction material for haul roads, laydown pads, and the tailings dam starter dyke. Although not shown on the production schedule, most likely some of this waste would actually be mined in Year -1 (2003) to prepare certain facilities prior to start-up.


 
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Phase 2 is the largest phase, containing about 65% of the total ore resource, and essentially develops the entire south end of the mine. This phase is partly mined in each year throughout the six-year mine life.

Phase 3 is partly stripped in Year 1 to provide construction material and then is not mined again until the end of Year 3. This pit is gradually mined over four years in order to smooth out its relatively high waste volumes. Since mining in Phases 2 and 3 occurs concurrently there is limited opportunity for pit backfilling. The lowermost benches from Phase 3 can be placed into the bottom of the mined out Phase 2 pit. Detailed planning at the feasibility stage may allow more pit backfilling, which is beneficial from both operating cost and environmental perspectives.

   
7.3

GRADE CONTROL

The two ore types (Tension vein and Albion vein) that will be mined may require different grade control measures. Blasthole cuttings will be assayed and ore waste contacts will be identified. However due to the sub-vertical orientation of much if the veining and the nugget effect, it may become evident that vertical blastholes cannot provide the necessary accuracy in ore control.

The Albion vein structure is visually distinguishable and therefore the field geologists may be able to identify the ore/waste contacts for mining.

In the Tension vein zones visual control is likely impractical. Blasthole assays combined with ore/waste definition from the block model may be required to outline the ore zones. This could lessen the mount of selective mining possible and could impact head grades sent to the mill.

   
7.4

RECLAMATION AND CLOSURE

No detailed closure plan has been developed since input from NovaGold management, environmental consultants, and regulator agencies must all be taken into consideration. However a few comments have been provided below.

Organic soils - although the organic soils at site are less than a metre thick, it is assumed that in areas disturbed by construction or mining, these soils would be dozed into piles and stockpiled at site. As development rock dump or tailings areas are finished and re-graded these soils can be spread across the surface or placed as “islands' to allow the reestablishment of native vegetation.


 
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Development rock dump - it is assumed that slopes will be re-graded to 3:1 and may or may not be re-vegetated.

Tailings Facility – at closure, free water will be drained from the tailings area and on-going seepage will be collected and eventually discharged once clarified. Permafrost will likely aggrade into the tailings deposits eventually freezing the remaining pore water in place.

Haul Roads - it is assumed that haul roads will be ripped and loosened to allow the reestablishment of native vegetation.

Infrastructure - it is assumed that all structures, such as shops, offices, explosive magazines, and diesel tanks will be decommissioned and removed from site and taken back to Nome. There may be some salvage value for many of these items but such revenues have not been included in the economic model.

Open Pits - since the open pit will largely be un-backfilled, the pit will likely fill with water over time to about the 65m elevation, which is the lowest point along the south side of the pit. Sections of the pit wall up to 85 metres high will therefore be above water level along the north side. No re-grading of the exposed pit slopes is planned.

Drainage systems - it may be desirable to re-establish pre-mining drainage paths, including ensuring the continual flow of water through the pit. The post-mining drainage plan must be discussed with regulators.


 
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TABLE 7.2
PRODUCTION SCHEDULE SUMMARY (Base Case)
1.78 Mtpy Milling Rate

  1
2006
2
2007
3
2008
4
2009
5
2010
6
2011
Totals


Waste                                             Mt

Tension Ore
       Ore Mined                               Mt
       Grade                                        gpt
       Insitu Gold                               oz
       Recovery                                  %
       Recov Gold                              oz


7.06


1.61
1.900
98,611
97.0%
95,653


8.44


1.64
1.661
87,515
97.0%
84,889

9.20


1.57
1.672
84,353
97.0%
81,822
8.46


1.51
1.947
94,481
97.0%
91,646
6.83


1.47
1.801
84,959
97.0%
82,410
7.49


1.36
2.189
95,786
97.0%
92,913
47.484


9.16
1.853
545,705
97.0%
529,334

Albion Ore
       Ore Mined                               Mt
       Grade                                        gpt
       Insitu Gold                               oz
       Recovery                                  %
       Recov Gold                              oz
0.17
3.521
19,001
94.0%
17,861
0.14
2.388
11,024
94.0%
10,363
0.21
2.706
18,508
94.0%
17,397
0.27
3.509
30,744
94.0%
28,900
0.32
3.112
31,526
94.0%
29,635
0.42
3.152
42,985
94.0%
40,406
1.54
3.114
153,789
94.0%
144,562

Total Ore
       Ore Mined                               Mt
       Grade                                        gpt
       Insitu Gold                               oz
       Recovery                                  %
       Recov Gold                              oz
1.78
2.053
117,612
96.5%
113,514
1.78
1.720
98,539
96.7%
95,252
1.78
1.795
102,861
96.5%
99,220
1.78
2.186
125,225
96.3%
120,546
1.78
2.033
116,485
96.2%
112,045
1.79
2.418
138,771
96.1%
133,319
10.695
2.034
699,494
96.3%
673,895

Total Material

Waste:Ore ratio
8.84

3.96
10.22

4.74
10.98

5.16
10.25

4.75
8.61

3.83
9.28

4.20
58.18

4.44

 
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8
  

WASTE DISPOSAL STRATEGY

The two main waste products to be generated by the Rock Creek mining operation will be the development rock stripped from the mine and tailings generated by the milling process. Each of these products will be disposed of separately.

     
8.1

ORGANIC SOILS

Although the organics soils at site are relatively thin and will be small in volume, they will be salvaged in areas that will be disturbed by construction or mining. Within the plant site, tailings dam footprint, and pit area, the organic soil will be dozed into piles and hauled to local stockpiles for final reclamation purposes. Removal of organic soils may be specified for critical foundation design reasons. The organic soils will not be recovered from beneath the development rock dump areas or within the tailings impoundment area.

     
8.2

DEVELOPMENT ROCK MANAGEMENT

Development rock stripped from the pit will be placed into development rock dumps located adjacent to the pit. If pit sequencing permits, some waste may be backfilled into mined out portions of the pit. Some development rock will also be used periodically for raising the tailings dam.

Development rock will be characterized to confirm the neutralizing potential. In order to prevent the release of seepage and surface runoff from the development rock dumps, they have all been located within the watershed for the Rock Creek Project. Diversion ditches surrounding the site will prevent runoff from undisturbed areas from coming in contact with the development rock.

The mine plans entails the development of three individual development rock dumps, as shown in Figure 8.1. Table 8.1 summarizes the development rock material balance, highlighting the volume of the different development rock dumps. The in-situ density of the rock is 2.6 t/m3 with a loose bulk density of 2.0 t/m3, assuming a net swell factor of 30%. The siting of the various dumps is based on the following factors:

     
  Minimizing truck haul distances outside the pit
  Ensuring a minimum 50m setback from the dump toe to the pit crest
  Minimizing the number of separate dump areas
 
Maintaining development rock placement to within the project watershed already being disturbed by mining activity.

 
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TABLE 8.1
DEVELOPMENT ROCK MATERIAL BALANCE

Location Phase 1
(kt)
Phase 2
(kt)
Phase 3
(kt)
Total
(kt)
Tailings Dam
North Dump (N1)
North Dump (N2)
South Dump (S1)
Pit Backfill
600

600
200

9,400

13,700
6,000

4,000
8,020
3,000

2,000
14,000
8,020
17,300
6,200
2,000
Total 1,370 29,100 17,020 47,490

8.3

TAILINGS MANAGEMENT

A single tailings facility will be used to store the mill tailings, provide seasonal water inventory for processing needs, and act as a storm water runoff buffer.

The tailings will be retained behind a centerline constructed containment dam, which will be constructed using mined development rock and will incorporate a graded filter zone to mitigate piping issues. Specifications for the dykes have not been developed at this time, but Figure 4.4 provides a cross-sectional schematic while Section 4.1.3 describes the anticipated geotechnical design criteria.

As the tailings are discharged into the tailings pond, the larger particle will tend to rapidly segregate, forming a sloping beach. The silt-sized fraction, typically less than 200-mesh (74 µm), will tend to wash out into the center of the pond and gradually settle. Beach dry densities could be in the range of 1.55 t/m3 while the finer interior deposit densities could be in the range of 0.6 t/m3. For sizing the tailings facility, an average density of 1.3 t/m3 has been used.

The total mined ore tonnage is about 10.7 million tonnes, which will generate about 8.2 million m3 of tailings.

Figure 8.2 provides a fill curve for the tailings management facility layout shown in Figure 8.1. The graph provides the cumulative tailings area storage capacity versus tailings elevation on the y-axis. The figure also shows the volume of dam fill required corresponding to different dam elevations. It is assumed that the dam would be built two


 
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metres higher than the tailings elevation to provide a freeboard allowance. In order to store 8.3 million m3 of tailings, the tailings elevation would be about 73 metres (240 ft). The dyke would be built to 75m elevation (246 ft.) and require about 7 million m3 of rock fill or about 14 million tonnes of rock.

To store the first year of tailings (about 1.8 million tonnes or 1.4 Mm3) before raising the dam, a tailings elevation of about 60m will be developed, requiring a dam crest of 62m (203 ft.). This will require about 3 million tonnes of development rock initially.

FIGURE 8.2: TAILINGS STORAGE CURVE

 
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9  MINE EQUIPMENT
     
9.1

MINING EQUIPMENT

The major mining equipment fleet for the Rock Creek Project will be limited in size. The annual quantities of combined ore and development rock to be mined range from 8.6 to 11.0 million tonnes.

     
  9.1.1

Loading Equipment

Due to limited production requirement and the need for operating flexibility, the Rock Creek Project is well suited as a front-end loader operation. A 12 m3 front end loader, equivalent to a CAT 992G loader, will provide sufficient capacity for the ore and development rock loading operations. Table 9.1 summarizes the basis for the loading productivity calculation, which indicates an hourly capacity of 1250 t/h. Loader availability is assumed to be 90%.

Due to the production demands, two loaders will be purchased. The second loader will also be used to feed ore to the mill from the dead storage areas in the stockpile when the mine loader is operating on waste stripping operations.


 
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TABLE 9.1
LOADING EQUIPMENT PRODUCTIVITY

LOADING PRODUCTIVITY - Rock Creek        
      Material Porphry Porphry Porphry  
Manufacturer Cat Cat Cat  
Model 992G 992G 992G  
Truck Size 773D 775E 777D  
MATERIAL              
      Material Weight (Bank) - dry dkg/BCM A density tests 2650 2650 2650  
      Moisture Content % B density tests 2% 2% 2%  
      Wet Density wkg/BCM C A x (1+B) 2703 2703 2703  
      Percent Swell factor % D estimate 30 30 30  
      Swell Ratio % E 1 / (1+C) 77 77 77  
      Material Weight (Loose-wet) wkg/LCM F C x E 2079 2079 2079  
               
LOADING UNIT              
      Rated Bucket Capacity - heaped m3 E Handbook 10.0 12.0 12.0  
            Bucket Fill Factor % F estimate 90 90 90  
      Actual Bucket Capacity lcm G E x F 9.0 10.8 10.8  
      Actual Bucket Capacity BCM   G x E 6.9 8.3 8.3  
      Material Weight (Loose) wmt/LCM H F / 1000 2.08 2.08 2.08  
            Tonnes/ Pass wmt J G x H 18.7 22.5 22.5  
               
TRUCK              
      Rated Truck Size (Heap) - wet wmt K Wajax 54 64 91  
      Rated Truck Size (Heap) - dry dmt   K / (1+B) 53 63 89  
      Rated Truck Capacity (Heap) m3 L Handbook 35.2 41.5 60.1  
            Truck Fill Factor % M estimate 100% 100% 100%  
               
      Loaded Truck Size wmt N K x M 54.0 64.0 91.0  
      Loaded Truck Capacity m3 P L x M 35.2 41.5 60.1  
      Theoretical Passes - Tonnage passes   N / J 2.9 2.9 4.1  
      Theoretical Passes - Volume passes   P / G 3.9 3.8 5.6  
      Actual Passes passes Q Select to not overload 3 3 4  
      Actual Tonnage   R Q x J 56.1 67.4 89.8  
      Actual Volume   S Q x G 27.0 32.4 43.2  
      Truck Capacity Utilized (tonnage)     R / K 104% 105% 99%  
      Truck Capacity Utilized (volume)     S / L 77% 78% 72%  
      Average Load cycle time per pass sec t1 estimate 45 45 45  
      First bucket time sec t2 estimate 10 10 10  
      Truck Spot Time sec t3 estimate 30 30 30  
      Load Time per Truck sec T (Q-1) x t1 + t3 + t2 130 130 175  
      Load Time per Truck min TT T / 60 2.17 2.17 2.92  
               
Loader Productivity (maximum) trucks/hr U 60 / TT 27.7 27.7 20.6  
      Loader production per 60 min hour wt/60min hr LP U x R 1555 1866 1848  
      Utilization 45/60 min X Op Factor 75 75 75  
      Op eff, moves, waiting due to Fleet mismatch % Y estimate 90% 90% 90%  
      Loader production per NOH 45/60 min LP50 LP x X x Y 1049 1259 1247  
      Total Hours per Shift (TH) hrs/shift V Planned 12 12 12  
                     Delay - lunch hrs/shift d1 Planned 0.50 0.50 0.50  
                  Delay - coffee breaks hrs/shift d2 Planned 0.50 0.50 0.50  
Delay - inspection, shift change hrs/shift d3 Planned 0.50 0.50 0.50  
      Total Operating Delays hrs/shift DD d1+d2+d3 1.50 1.50 1.50  
      Working Hours per Shift (NOH) hrs/shift W V - DD 10.50 10.50 10.50  
      Effective Operating Hrs/Shift (EOH) hrs/shift Z W x X x Y 7.1 7.1 7.1  
      Loader production per shift t per shift Zz W x LP50 11,018 13,222 13,096  
      Loader production per day tpd tpd Ee x Zz 22,037 26,444 26,192  
      Actual Trucks Loaded trucks/shift Aa ZZ / R 196.3 196.3 145.8  
      "swings per day" factor       1178 1178 1166  
Planned Loader Production              
      Actual truck load wt / truck R R (wet weight) 56.1 67.4 89.8  
      Actual production per shift wt / shift Bb Aa X R 11,018 13,222 13,096  
      Tonnes per Eff Op Hour (EOH) wt / op hr Cc Bb / Z 1,555 1,866 1,848  
      Tonnes per Working Hour (NOH) wt / stat hr Dd Bb / W 1,049 1,259 1,247  
      Tonnes per Total Hour (TH) wt / shift hr   Bb / V 918 1,102 1,091  
               
Yearly production              
      Scheduled Shifts / day # shift/week Ee Planned 2 2 2  
      Scheduled Operating Days/Year # days/yr Ff Planned 365 365 365  
      Total Shifts / Year # shift/yr Gg Ee x Ff 730 730 730  
      Loader Mechanical Availability % Hh estimate 90% 90% 90%  
      Use of Availability (standby hours) % SB estimate 95% 95% 95%  
      Scheduled Hours per year per year TH Ee x 24 8760 8760 8760  
      Downtime hours per year DT AH x (1-Hh) -876 -876 -876  
      Standby Hours per year SB (TH-DT )x (1-SB) -394 -394 -394  
      Gross Operating Hours (GOH) per year GOH TH-DT-SB 7,490 7,490 7,490  
      Operating Delays per year OD GOH x (DD/V) -936 -936 -936  
      Net Working Hours (NOH) per year NOH GOH - OD 6,554 6,554 6,554  
      Annual Loading Capacity per Loader wt per shovel TPY NOH x Dd 6,877,137 8,252,564 8,173,968  
      Tonnes per Month (average over year) wt per month TPMa TPY / 12 573,095 687,714 681,164  
      Tonnes per Day (average over year) wt per day TPDa TPY / Ff 18,841 22,610 22,394  
 
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  9.1.2

Hauling Equipment

The target for sizing the truck fleet is to operate with a fleet of about 4 trucks. This minimizes the number of operators yet provides flexibility during periods of truck downtime.

The optimum truck size is about 90 to 100 tonnes, equivalent to the CAT 777 truck. Typically three trucks would be operating at all times while the fourth truck would be being serviced.

     
  9.1.3

Drilling and Blasting Equipment

It is assumed that the ore and development rock will utilize the same powder factors and blasting patterns. In actuality the mineralized zone may be more fractured and could require less blasting effort.

The drill and blast operation is based on a powder factor of 0.2 kg/t and a 150mm diameter blasthole while the bench height is 5 metres and blasthole spacing will be about 4 metres. It is assumed that most holes in the permafrost will be dry and that ANFO can be used. If wet holes are encountered then plastic liners are an option or an emulsion blasting agent can be used. A blasting truck and ANFO truck will be purchased.

     
  9.1.4

Mine Support & Miscellaneous Equipment

Other support equipment that will be required at the Rock Creek Project includes the following:


          Dozers, 302 kW 2  
        Graders, 160 kW 1  
        Water truck, 15000L 1  
        Fuel & Lube truck, 2000 L 1  
        Mechanic service vehicles 1  
        Pickup trucks 4  
        Crewcab trucks, 4x4, 1t 2  
        Light Plants 3  
        Float trailer & tractor 1  
        Fork Lift 1  

 
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  9.1.5

AN Storage

Blasting agents will be barged in to Nome during the summer months therefore about 8 months of ammonium nitrate storage is required. A total of about 1300 tonnes of AN will require storage with a peak annual consumption of 2200 tonnes of AN. Several options for AN storage are available.

Option 1: Ammonium nitrate can be stored in one or two-tonne tote bags. The AN storage pad would have a gravel surfacing and a HDPE liner. A small auger could be used to transfer AN into the trucks.

Option 2: A steel frame structure with a concrete floor can be used. The building would a concrete floor and approx. 1-metre high concrete walls. The structural steel may need to be epoxy coated and the building may require a dehumidifier system.

Option 3: AN shipped to site may be stored directly in silos.

AN storage will be in accordance with MSHA regulations. For the purposes of this study, the Option 2 ammonium nitrate storage approach is assumed.

     
  9.1.6

Explosive Magazine

Explosive will be stored in two separate magazines one for detonators and the other for primers and boosters. The magazine area will be bermed off and located away from the rest of the project facilities, as shown in Figure 8.1. Explosives will be stored according to MSHA regulations.


 
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10  MILLING
     
10.1

METALLURGICAL STUDIES

The following is a summary of metallurgical test work that has been undertaken over the last several years on samples from the Rock Creek site. Metallurgical summary/interpretation reports were available for review, but no prime sources were available with the exception of the McClelland Laboratory report from 2000, and some recent scoping centrifugal gravity concentrator test results. Earlier results are summarized in a number of documents but in many cases the source of the samples is unclear. Some of these “abstracted” results are quite interesting but the full significance is uncertain because of the lack of provenance for the data.

In his February 2000 review, Mr. Phil St. George reviewed the McClelland report and commented on earlier test work. Some of the comments summarized below are derived from his report.

     
  10.1.1

Early Studies

Early metallurgical test work by Newmont indicated >80% of gold reports to a gravity concentrate with a 48-mesh grind.

Metallurgical test work by Placer Dome indicated 92% and 93% recovery with cyanidation and floatation respectively. The grind for this test work is not reported in Mr. St. George’s review.

Samples for both of these studies were taken from the sheeted vein area and were surface samples. Norwest did not review test reports for this work. Albion ore type was likely not tested.

A brief anonymous report produced during the acquisition of Alaska Gold by NovaGold as part of the due diligence also reports on metallurgy and geology. In it Placer Dome work carried out in 1989-1990 is reported on. Grinds of 52-60 micron are reported to give recoveries in the low 90%'s by flotation and cyanidation. This is probably the same data as reported by Phil St George. In 1990 pilot plant runs were carried out on feed material assaying 0.35-1.82 g/t (head assay) and 0.35-1.93 g/t (calculated head). There is poor correlation between measured and calculated head, and this is frequently a sign of coarse, gravity recoverable gold being present. In total, pilot plant runs 101-107, treated 93.2 tonnes of 0.84g/t (mean assayed head) or 1.25g/t (mean calculated head) to give


 
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81.8-85% recovery. Additional data from this test work was indications of ore specific gravity (2.65) and Work Index (16.9-18.1 kWh/tonne).

No mention of the Newmont work is made in this summary; however a KD Engineering report is mentioned as part of the Independent Mining Consultants (IMC) 1996 review. It seems this report looked at a gravity/flotation plant on a 100 mesh ground product. The conclusions were that this was uneconomic but no indications of head grade or treatment rate was available, although the mineable resource was quoted as 5.293 million tons at 0.074 oz/ton.

Other anonymous review document repeats the above data in various forms. It also comments on the Placer Dome work, and states that gravity recovery in the pilot plant typically recovered over 50% of the Gold. It would seem that leaching flotation concentrate suggested that preg-robbing carbonaceous species were present and the plant would have to be a CIL plant for leaching flotation concentrate.

     
  10.1.2

1999 Program

In 1999 two samples were made from rotary drill cuttings and tested in McClelland Laboratories in Reno Nevada. Composite 1 was made up of high grade, quartz rich intervals from the Albion shear (AS). Composite 2 was made up of higher-grade intervals from the sheeted veins (SV) to the southeast of the Albion shear (RMR-3 and RMR-5). These samples were subjected to both gravity concentration and whole rock cyanidation testing. Work focussed on Composite 1 to save costs and because it was thought to have more difficult metallurgy due to finer grained gold and sulfosalts within quartz veins.

Having looked at a number of grind sizes it was concluded that 80% passing 65 mesh was optimum for conventional (very low mass pull) gravity recovery. This conclusion was based on hand panning ground samples. Pilot scale (75kg) testing was carried out on 65 mesh products and it was concluded Composite 1 indicated a potential for approximately 37% recovery, and Composite 2 for 86% recovery (concentrate+middlings). This would tend to indicate that gravity alone would not give acceptable recoveries for Albion shear material.

Cyanidation was carried out on both ore and gravity tailings. On ore, recoveries were 89.4% and 94.9% at a 200 mesh grind for Composite 1 and 2 respectively. This compares with 62.7% and 60.7% at 10mesh (as received). When treating gravity tailings directly (65mesh), total recoveries (Gravity + Cyanidation) of


 
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80.5% and 91.8%, whereas after regrinding to 200 mesh (from 65 mesh) combined Gravity + Cyanidation, results gave recoveries of 87.2% and 98.1% for composites 1 & 2 respectively. Clearly improved recoveries on both composites are possible by treating gravity tailings by cyanidation, but this would result in all tailings being contaminated by cyanide, and thus requiring treatment for cyanide removal/destruction. In general required leach times were long (96 hrs) on ore but significantly lower on gravity tailings (48hrs) even on the un-ground material. Cyanide and lime consumption were low to moderate for un-reground material, rising to high for fine reground material.

TABLE 10.1
GRAVITY CONCENTRATION RESULTS

  Conc                  Mids     
Recovery          Recovery    
Total Gravity
Recovery
Grind Size
P80 (micron)
Mass
Pull
Composite 1 (AS)

Composite 1 (AS)

Composite 1 (AS)

Composite 1 (AS)

23.1%                  8.0%     

41.2%                  6.2%     

12.0%                  3.7%     

26.8%                  3.2%     

31.1%

47.4%

15.7%

30.0%

425µ

212µ

150µ

106µ

3.9%

4.1%

1.4%

2.6%
Composite 2 (SV) 76.8%                  9.4%      86.2% 212µ 7.2%

TABLE 10.2
PILOT GRAVITY & CYANIDATION OF MIDDLINGS/TAILS RESULTS

Tests used a 65 mesh
(212 µ) grind
Gravity Conc
Recovery
Mids & Tails Cyn
Recovery
Total Recovery
Composite 1 (AS) 34.3% 52.9% 87.2%
Composite 2 (SV) 76.7% 21.4% 98.1%

 
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TABLE 10.3
DIRECT AGITATED CYANIDATION TESTING

  Gold
Recovery
Grind Size
P80
(micron)
Leach Time
COMPOSITE 1 (AS)

COMPOSITE 1 (AS)
62.7%

89.4%
1700µ

74µ
96 HOURS

96 HOURS
COMPOSITE 2 (SV)

COMPOSITE 2 (SV)
60.7%

94.9%
1700µ

74µ
96 HOURS

96 HOURS

  10.1.3

2001 Program

A series of gravity concentration tests were completed in November 2001 at Knelson's testing center in Langley, BC to determine the Gravity Recoverable Gold (GRG). Two composites were an Albion vein sample (281B) and a tension vein sample (281A). The results of the test work are summarized in Table 10.4.

TABLE 10.4
KNELSON GRAVITY CONCENTRATION RESULTS

  Gravity
Recoverable
Gold (GRG)
Grind Size
P80 (micron)
Mass Pull Passes through
circuit
Tension Vein 88.9% 100µ 3.40% 4 passes
Tension Vein 79.6% 100µ 2.55% 3 passes
Albion Vein 38.8% 115µ 1.89% 3 passes

  10.1.4

2003 Program

Early in 2003 it was realized that the conventional approach to all gravity recovery (low weight recovery to produce a high grade product) would not be satisfactory on Albion shear material as recoveries looked to be below 40%. With the advent of continuous centrifugal concentrators it was decided to investigate the potential of pulling a moderate weight recovery to produce a throw away tailing. A series of samples representing the Albion shear zone was selected and submitted to Falcon Concentrators (at PRA) for small-scale tests to evaluate the potential for gravity recovery.


 
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The test work carried out in a centrifugal batch concentrator treated ground products by multiple pass through the unit. The first concentrate was upgraded by hand panning. Engineers at Falcon, based on their experience with their equipment at full and laboratory scale, believe that the pan concentrate closely indicates the recovery likely achieved by a full scale batch concentrator, and that the four pass recovery indicates the recovery likely to be achieved by a continuous concentrator. The results are summarized in Table 10.5. In May 2003, a series of gravity concentration tests were completed at PAR laboratories using the Falcon concentrator in both batch and continuous modes. The data from this study is included in Appendix E.

The sample tested typifies the Albion zone, as it would be mined rather than a high-graded selection of Albion quartz from within the Albion zone. The Rock Creek deposit is broadly modeled into two zones, Albion and QMS, each of which have different grades and very different recovery characteristics. However, in a real mining situation, the higher grade of the two zones (Albion zone) will actually consist of material from both zones because the higher grade zone contains blocks and stringers of the other zone that are too small to differentiate while mining. Therefore the intent is to test a sample that represents the entire width of the high-grade (Albion) zone, including the included blocks of the low-grade zone.

The selected sample intercept is from DDH RKDC02-116. It is 30 meters long (fifteen 2-meter samples) in a 140‹ Az, –55‹ inclination core hole, through approximately 17 meters true width of the Albion zone. This interval is depicted in Appendix E and core photos of typical portions are included. Table 10.5 summarizes the test results. Figure 10.1 provide a graphical correlation between recovery and grind size (P80).

TABLE 10.5
FALCON GRAVITY CONCENTRATION RESULTS (2003)

  Calc'd Head
Grade (g/t)
Total
Recovery*2
Grind Size
P80
(micron)
Pan
Au*3
Mass
Recovery*1
Albion composite 5.07 g/t 50.0% 223µ 17.4% 27.6%
Albion composite 5.63 g/t 63.4% 184µ 20.5% 25.0%
Albion composite 4.60 g/t 69.4% 97µ 24.2% 17.3%
Albion composite 3.43 g/t 90.6% 53µ 51.2% 26.6%

Notes: *1:     The mass recovery is before cleaning. It is anticipated that the mass recovery could be reduced significantly without significant gold loss, but this is subject to test work.
  *2: The total recovery including “Pan Gold” On cleaning the gold not pulled into the high grade (pan)concentrate would be recovered to a middlings product for leaching or sale.
  *3: The pan concentrate represents gold that is recoverable into Dore on site. Note that the 1999 test work had a significantly higher projected recovery (from a different sample of Albion shear).

 
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Additional work is required to confirm the supposition that the recovery achieved in a 4-pass batch operation can be cleaned to a shippable grade.

Although good results were achieved at the 53 micron grind, grinding this fine is very costly. No recent work has been done on flotation (and we have no details of the work that was done by Placer Dome in 1989) but flotation is a well proven technique in pre-concentrating gold, and probably at a significantly coarser size.

FIGURE 10.1
GRAVITY RECOVERY VERSUS GRIND SIZE

  10.1.5

Discussion of Metallurgy

Historical data on what is probably tension vein material indicates that good recoveries can be achieved by a variety of techniques, including flotation and cyanidation. This early work does not appear to have focused on gravity, although Placer apparently claimed better than 50% recovery was achieved by gravity in their flowsheets.

The tension vein material gave good recoveries at coarse grinds in the 1999-2000 test work, with gravity only recoveries being 86% and potentially another 12% being achieved with cyanidation of gravity tailings. In contrast the Albion shear recoveries were below 40% by gravity, but increased to almost 90% with the addition of cyanidation of gravity tailings.

The avoidance of chemical use will simplify environmental permitting. However the benefits of the increased recoveries achieved by the addition of non-gravity processes needs to be critically evaluated. The treatment of gravity tailings with


 
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cyanide improves recoveries significantly especially on the Albion shear material, this would result in all tailings being contaminated with cyanide, and although cyanide can be destroyed/detoxified by oxidation this is expensive and potentially difficult to permit.

Some form of pre-concentration would enable most tailings to be disposed of without cyanide contamination and without treatment of the full tailings flow for cyanide detoxification. The recent work carried out with Falcon concentrators shows that good recoveries (+90%) can be achieved by using a continuous concentrator, but the quality of concentrate after cleaning was not determined (insufficient sample). It is expected that a concentrate of as little as 1% or 2% weight pull may be achievable after cleaning of the rougher gravity concentrate. If concentrate can be reduced to 1or 2% of feed weight (after removal of the “conventional” gravity gold) the low-grade gravity concentrate will be sufficiently high grade to support shipment of concentrate offsite. The major problem with the Falcon gravity test work was that the good recoveries were achieved after fine grinding: there was a straight-line relationship between recovery and fineness of grind as shown previously in Figure 10.1. Fine grinding is costly, especially with expensive power costs.

It is possible that a gravity concentrate may be made that would support shipment off-site to either a smelter or for processing through a cyanide leach plant on a custom basis. Alternatively the concentrate could be leached onsite and the small volume of “sensitive” tailings would be detoxified and then put into special storage.

An alternative to pulling a bulk gravity concentrate with a continuous Falcon (or similar) concentrator, would be the recovery of a flotation concentrate from the conventional gravity tailings (cyclone overflow). This flowsheet has been widely used historically and has been applied to a number of recent projects (e.g. Julietta (Bema Gold) in the Russian Far East). Placer Dome reported good flotation recoveries from 1990, but this may have been on tension vein material. One benefit of flotation is that it may be possible to achieve good recoveries at a relatively coarse grind than required for gravity alone. Many porphyry copper mines carry out rougher flotation at a grind of 150 microns. In flotation most chemicals used are precipitated on the surface of the concentrated minerals and thus do not report to tailings. Frothers are in the tailings solution, however, these are volatile alcohol-like molecules that research has shown at the concentrations used do not present a pollutant to aquatic habitat.


 
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10.2

FUTURE METALLURGICAL WORK

Future metallurgical work can be broken down into two principal objectives:

     

  
(1)Metallurgical characterization of the Rock Creek resource so that parameters such as hardness and recoveries can be established for the resource. This is what is ideally done on all projects.
     
 
(2) Development work on the Albion shear vein material (and any other difficult material identified in 1 above). This work is required to establish:
     
 
Continuous concentrator recoveries and their relationship with primary grind and is an extension of the recent Falcon test work. Alternative machines should be tested (Inline Pressure Jig, Knelson, etc.) should be tested to determine which machine gives the best performance.
 
Cleaner concentrate recoveries and grades achievable from the continuous concentrator concentrates.
 
Cyanide leaching characteristics of cleaned continuous concentrator concentrates.
 
Flotation recovery and grind relationship to recover gold from the batch concentrator tailings. This includes cleaning rougher concentrates to evaluate the benefit (if any) of utilizing flotation rather than continuous gravity concentrators to increase recovery of gold from the Albion shear.
 
Cyanide leaching flotation concentrates.
 
Testing the behavior of tension vein material on any attractive process identified from the Albion shear test work. This will enable overall metallurgy to be predicted.
     
 
The precise details of the above program need to be determined in conjunction with the mining plan and significance of the various areas of the mine and their contribution to the resource.
     
10.3

PLANT DESIGN CONSIDERATIONS

The very high recoveries achieved by a simple gravity plant at a coarse grind on the tension vein ore indicate that a relatively simple plant may be possible. However this is complicated by the requirements to treat Albion shear ore, which is totally different in metallurgical performance. The ores are not really compatible to co-treatment, but neither campaigning material nor building parallel circuits is an attractive alternative.

 
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The requirements for tension vein is for an all gravity plant to recover approximately 86% gold to a clean concentrate smeltable to Dore after cleaning. The primary grind is coarse at 80% passing 65 mesh. Tailings from this all gravity circuit may be further treated to enhance recoveries. The Albion ore processed through this circuit would have a recovery below 40%, which would not normally be acceptable.

With a coarse grind requirement of 80% passing 65 mesh a single stage “European” style SAG mill (length to diameter ratio of >1:1 probably 1.5:1 would be able to achieve this objective most economically. Alternatively fine crushing and a single stage ball mill could be used. The conventional North American style SAG mill (length to diameter ratio of < 1: 2) is unlikely to achieve the required grind without the inclusion of a secondary ball mill, and would thus be a more expensive circuit.

To enable improved gravity recoveries to be achieved from Albion shear material grinds significantly finer than 65 mesh are required, the best results (+90% recovery) being achieved at a grind of 80% passing 270 mesh (53 microns). On material with a work index of 16.5 this approximately doubles grinding power requirements for an additional 11.3 kWh/tonne which at $0.12/kWh is an additional $1.35/tonne in power and perhaps an additional 0.5kg of balls worth about $0.25/tonne. The improved recovery of 50% on 3g/t Albion shear material is worth over $16/tonne, and if this material represents 20% of mill feed, the additional operating costs of $1.60/tonne on ALL mill feed produce $3.20/tonne additional revenue on average. This ignores the cost of processing the additional rougher concentrate but also ignores any additional revenue that may be achieved from tension veins. The higher percentage of Albion ore in mill feed the greater the benefit of fine grinding to recover a continuous concentrate. If the finer grind is justified on the basis of this improved recovery, then a SAG/Ball grinding circuit would be required at additional costs over the single stage “European style” SAG mill.

Whilst the additional grinding appears to be justified by the additional recoveries achieved on Albion ore, similar improvements in recoveries were achieved by cyanidation at the coarse grind required for tension vein material. Insufficient factual information is available to comment on potential benefits of flotation except to say that recovery of a lower grade concentrate is likely without having to go to the fine grind required for gravity.

   
10.4

BASIS OF CAPITAL AND OPERATING COST ESTIMATES

The capital cost estimates are based on soliciting and receiving quotations for major items of process equipment from reputable suppliers of such equipment. Costs of installing this equipment are based on a recent detailed estimates carried out on a similar sized gold plant in Nunavut, Canada. Costs were modified to take into account specific changes in plant


 
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design as well as revised to reflect different labour, concrete and other unit costs, as well as revised to reflect differences in flowsheets.

It is interesting to note that the alternative flowsheets examined, which included very significantly different grinding requirements did not impact the capital costs very significantly. The corollary of this is that changing plant throughput has a minor impact on capital cost as the main change on plant throughput would be on grinding requirements.

Operating costs were worked up from first principles. The cost of power has a major impact on plant operating costs.


 
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11
  

PROJECT INFRASTRUCTURE AND SERVICES

The plant site layout is shown in Figure 8.1 and located to the west of the open pit and to the north of the tailings facility. The plant area includes space for the primary crusher and crushed ore stockpile, the mill, a maintenance shop, an administration mine dry building and fuel storage. Ammonium nitrate and explosives will be stored remotely away from the main plant area.

No detailed site investigations or geotechnical investigation has been done to confirm the suitability of these sites. In addition condemnation drilling should be done to confirm that no mineable ore zones underlie these areas.

     
11.1

ADMINISTRATION OFFICES

The administration building will consist of single story pre-fabricated construction. Such structure can be transported to the site in containers and assembled on prepared concrete slabs.

The administration building will be about 1000 m2 in size however no detailed layouts have been prepared at this time.

     
11.2

MINE MAINTENANCE SHOP & WAREHOUSES

The mine maintenance shop would consist of a pre-fabricated structure. It would include:

     
  two large mobile equipment repair bays
  a smaller light vehicle repair bay
  an overhead crane system
  a wash bay supplied with a high pressure monitor
  a welding bay a tire bay
  a lubricant distribution system
  offices for maintenance staff
     
 
Two warehouses or storage areas would be required adjacent to the truck shop. One building would provide heated storage while the other would provide cold storage. Sprung structures could be used for the warehouse facilities.

 
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11.3

MINE DRY

A mine dry would be located adjacent to the administration building and central to the site facilities. The dry would serve both the mill and mine operating personnel.

This single store structure would also use prefabricated construction. Where buildings are not directly connected, enclosed arctic corridors will be used for personnel access from one building to the other.

An assay laboratory would be included in the mine dry building. This lab would perform routine analysis of blasthole samples and processing samples and require about 150 m2.

   
11.4

ELECTRIC POWER SUPPLY

The power supply system has been sized to accommodate the process loads from the mill and ancillary services. It is estimated that a total power requirement of 4 to 5 MW will be needed.

There are two potential options for electric power supply for the Rock Creek project.

Option 1: One option is to tie into the local Nome Joint Utility System power grid.

Option 2: The second power supply option is to generate power on site using a dedicated diesel genset.

For the purposes of this study, power supply from the local power utility has been assumed as the base case.

Figure 11.1 is a single line diagram of the current power grid provided by Electric Power Systems Inc (EPS) of Anchorage. The highest voltage delivered along Teller Highway is about 12 kv.

The Rock Creek Project will require about 4-5 MW of electric power, most of which is dictated by the milling operation. It is recommended that the power be delivered to the site via a 25 kv powerline. In order to do this a step-up transformer will be required at the generating station, the power line to Teller Highway must be upgraded from 12 kv to 25 kv capability, and a new power line must be installed along the Glacier Creek By-pass road to the mine site.

Power distribution around the site would probably be done at 6 kv and therefore a step-down transformer would be required at the main substation.

 
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The estimated cost for electric power from the local utility is US$0.12/kwh however further discussion with the power utility is required to confirm the responsibility for installing the step-up transformer and upgrading the power line. The project capital cost estimate includes costs for all of these items. If the utility were to undertake these capital improvements, then likely they may incorporate them into their electric power charges and the project capital could be reduced accordingly.

It is recommended that a trade-off study be undertaken to further evaluate the electric power options. Some concerns raised with on-site power generation are the need to store larger quantities of diesel fuel at site and the amount of emissions generated and subsequent permitting impacts. Given the relatively short project life of six years, an on-site diesel generator may still have significant salvage value at project closure.

From discussions with Wartsila, a 7 MW sized diesel power plant could have an installed cost of $575/kW or about $4 million (continental US cost). Adding a 15% allowance for a Nome installation, the capital cost could be in the range of $4.6 million. Diesel power generation costs will be dependent on fuel cost; the relationship is shown in Table 11.1 The estimated diesel fuel cost delivered to Rock Creek could be about $1.00-$1.10/gallon, resulting in a power cost of $0.06/kwh. Adding in labour, maintenance, and other costs could result in a net cost of about $0.10/kwh. Lube oil consumption of a Wartsila 18W32 unit would not, at full load of 7057 kW, between 25 to 35 gallons per day of lube oil.

Such power plants are typically prefabricated and modularize off site, such that the pipefitting work and detailed installation at site is minimized. Erection and assembly of large pieces is required, using a flat floor, slab on grade type construction. Additional capital costs would be incurred for larger on-site fuel storage and the associated spill containment measures required.


 
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TABLE 11.1
DIESEL GENERATED ELECTRIC POWER COST
(Fuel Only Component)

Fuel Cost Power Cost per
kwh
0.60 $0.0350
$0.70 $0.0409
$0.80 $0.0467
$0.90 $0.0526
$1.00 $0.0584
$1.10 $0.0643
$1.20 $0.0700
$1.30 $0.0759
$1.40 $0.0816

11.5

FUEL STORAGE

All diesel fuel is barged in to Nome during the summer season. Near the port several large diesel tank storage areas are available. The two main ones are:

Bonanza Fuel, Inc; is a 100% wholly owned subsidiary of Nome Native Community Enterprises, Inc. The tank farm is located on the port-pad adjacent to the City of Nome Causeway. The configuration of the tank farm is six 611,000 gallon capacity tanks for fuel oil #1, #2, unleaded gasoline and jet fuel.

Crowley Marine Services, The tank farm is also located on the port-pad adjacent to the City of Nome Causeway. The capacity of the tank farm is about 5.5 million gallons.

Discussions with these suppliers indicate that their current storage capacity is fully utilized. It is estimated that annual diesel fuel requirements for Rock Creek Project will be in the range of 3.8 million litres (1 million US gallons). The Rock Creek capital cost estimates assumes that one of these suppliers would construct additional storage at their cost.

Diesel fuel would be delivered to the Rock Creek site via tanker truck. Diesel storage at site will be minimized to reduce potential environmental liability. Assuming a one-week supply, the site would contain about 73,000 litres (20,000 gallons) storage.

Gasoline consumption at site will be less, mainly for service vehicles. On site storage is assumed to be 6000 litres (1500 gallons).


 
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Both of these fuel storage tanks would be contained within lined and bermed off areas, with 110% spill capacity.

Diesel fuel costs delivered to site can be estimated as shown in Table 11.2, based on data derived from The Nome Joint Utility System.

TABLE 11.2
DIESEL FUEL COST BREAKDOWN



US$ per gallon
Base Price $0.52
transport, insurance $0.19
price adjustment $0.18
wharfage $0.03
Total Cost $0.92
Other costs, insurance, environ, $0.09
tankage cost $0.04
delivery to Rock Creek site $0.03
Total Fuel Cost (US$) $1.08/gallon

$ 0.28/litre
from: Nome Joint Utility System

 

Lubricants will be delivered to site in drums and stored in a secured area. The lubricants will be distributed with hose reels in the truck shop.

   
11.6

SEWAGE COLLECTION AND TREATMENT

There are several approaches to deal with sewage and grey water at site, depending on the number or personal assigned at site.

Option 1: Due to the frozen ground typical of the permafrost environment, a septic tile field is likely not a practical alternative.

Option 2: Another option is to operate with septic holding tanks that can be pumped out periodically with the effluent hauled into Nome.

Option 3: The other option is to utilize a Rotating Biological Contactor (RBC) to process the sewage on an-ongoing basis. The clean effluent from the RBC would be reused in the milling process via the tailings pond. Recycling of this water is done at Red Dog and


 
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Diavik although one must monitor the RBC performance by sampling and analyzing effluent prior to mixing with the tailings stream.

For the purposes of the cost estimate, the use of an RBC has been assumed.

   
11.7

PROCESS WATER SUPPLY

Due to uncertainty in the quality and quantity of local groundwater sources, it is assumed that groundwater pumping wells would be inadequate as a source of water supply. A couple of alternatives are available. It is estimated that normal process water consumption may be in the range of 1200 m3/day or about 420,000 m3 per year.

A large water inventory will be maintained in the tailings pond by collecting site runoff. Since the open pit is likely in permafrost ground, groundwater discharge into the pit may be low. Furthermore during winter there will be little surface runoff to collect. The entire project area is about 2 km x 1.5 km, or 3.0 km, which would generate an annual runoff volume of 1,200,000 m3 based on precipitation of 400mm per year. Even with some evaporative losses, there should be sufficient water using this approach. More detailed water balance modelling is needed at the next stage of design to confirm this.

Fire water will be provided from either runoff collection sumps or from the tailings pond.

   
11.8

POTABLE WATER

Due to uncertainty in the quality and quantity of local groundwater sources, it is assumed that potable water would be supplied via tanker truck and/or bottled water from Nome. Potable water supply to the facilities would be by gravity flow from a lined, above ground portable water storage tank.


 
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12

COST ESTIMATE

The following section describes the estimated projects costs, sub-divided into capital, sustaining capex, and operating costs and owners costs.

     
12.1

CAPITAL COSTS

The capital cost estimate is based on the project scope described previously. The capital cost takes into account the location of the project, the local environment, and the accessibility of the site. Capital costs include all expenditures up to and including commissioning of the project facilities. Ongoing capital expenditures are included in sustaining capital, described in Section 12.2.

     
  12.1.1

Basis of Estimate

The cost estimates have been developed from engineering estimates, data from Norwest's in-house databases, and cost quotations from Nome or North American suppliers. All costs are in 2003 US dollars.

All equipment costs are based on procuring new equipment. There may be opportunity for cost savings by purchasing used equipment.

The capital costs include both direct costs and indirect costs. Direct costs include all labour, materials, and equipment necessary for construction.

Indirect costs:
Are those cost incurred to support the direct construction activity and include overheads, freight, tools, insurance, transportation, supervision, and mob and demob. Indirect costs are calculated as a percentage of the direct costs based on the criteria shown in Table 12.1.


TABLE 12.1
INDIRECT COST BASIS

  % of Direct Cost
Infrastructure 5%
Buildings 5%
Mining & Equipment 5%
Milling 4%
Tailings 5%
Development Rock Dumps 5%

 
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EPCM costs:
Are based on a percentage of the direct costs according to the criteria shown in Table 12-2.

TABLE 12.2
EPCM COST BASIS

  % of Direct Cost
Infrastructure 5%
Buildings 5%
Mining & Equipment 3%
Milling 5%
Tailings 4%
Development Rock Dumps 4%

   

Allowances:
Are based on a percentage of the various costs according to the criteria shown in Table 12.3. The allowance provides for items in the scope of work that have not been defined in detail at this stage of study and to cover cost increases since quotations from suppliers have not been solicited. The expectation is that the allowance is part of the capital cost and therefore will be spent.

No contingencies have been applied in the cost estimate. Contingency is often used to set financing limits to ensure sufficient funds are available to complete the project. For example the owner may want to ensure that he has sufficient capital to fund the project with a 90% probability. The expectations are that contingencies need not necessarily be spent if no unforeseen circumstances arise.

TABLE 12.3
ALLOWANCE BASIS

  % of Cost
Infrastructure 8%
Buildings 8%
Mining & Equipment 5%
Milling 8%
Tailings 8%
Development Rock Dumps 20%
EPCM 10%
Indirects 10%
Owners Costs 5%
 
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  12.1.2

Project Development Costs

Infrastructure
Infrastructure capital is estimated to be $1.7 million. This cost would cover roads, ditching, plant site grading, water and power supply etc. The costs are detailed in Appendix H, Table H1, items 101 to 121.

Buildings and Support
$2.5 million is the estimated cost for constructing the offices, shops, warehouse, dry, and explosives facilities. Of this total the truck shop accounts for $1.1 million. The details are listed as items 201 to 213 in Table H1 of Appendix H.

Major Mining Equipment
This classification covers the drilling, loading and hauling units as well as the major support equipment consisting of dozers, graders and water trucks. The water truck is assumed to be a used unit. All the other equipment are assumed to be new units, delivered and assembled at the mine. Budgetary quotes, including freight to Nome were obtained. Therefore, indirects, EPCM and allowances were not included in the cost estimate. Fleet capital costs were based on the following equipment costs:

TABLE 12.4
MINING EQUIPMENT CAPITAL COST

Equipment Type Unit Price at Mine ($ 000’s)
Front End Loader: Cat 992G 1,470
Haul Truck: Cat 777D 1,065
Blasthole Drill: Drilltech D245KS 646
Dozer: Cat D10R 865
Grader: Cat 16H 590
Water Truck: Used Cat773 200

   

For the loaders, haul trucks and drills, the number of units to be purchased was determined from the calculated Net Operating Hours (NOH) available annually per unit. This was based on the estimated equipment availability, the working hours available and the productivity. The calculation methodology is shown in the PDC, Appendix A. Detailed calculations for all the equipment may be found in Tables 12.11 and 12.12.

The grader requirements were based on a percentage of the haul truck hours. A factor of 15% was estimated. In other words for every hour the truck is working, 0.15 hours of time must be spent grading the haul roads.


 
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Dozers are required to clean up the loading bench and push broken waste and ore to facilitate loading. A factor of 35% of front end loader hours was used to determine the dozing time required to support the loading unit. A dozer will also be required at the development rock dumps and for other miscellaneous duties. Therefore the calculated number of dozers was rounded upwards to account for this.

The following table illustrates the annual equipment requirements. As can be seen, one loader is required in 2006 and two loaders in 2007 and 2008. While only one loader is required for production from 2009 onwards, both loaders should be kept on site. Four haul trucks, one drill, one grader and two dozers will also be required for the project.


 
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The capital cost of the major mining equipment will be $10.4 million. The equipment would be barged in from Seattle. Due to the limited barging season, the equipment may have to be brought in during the summer of 2005.

Ancillary Equipment
Other ancillary equipment such as a fuel truck, pickups, service trucks, cranes, etc to support the mine’s activities are estimated to cost $1.0 million. The equipment list is detailed in Appendix H Table H1, items 308 to 324.

Mill Capital
Mill capital is estimated to be $16.5 million. This covers site preparation, the building, and all crushing, separating, smelting and handling equipment.

Tailings
Tailings facilities are estimated at $0.8 million. This would cover the access road, ditches, ponds and dam construction. The activity costs are detailed in Appendix H Table H1 under items 501 to 507.

Development Rock Dumps
Ditching and access roads are estimated to be $32,000, as itemized in Appendix H Table H1

EPCM
Based on the various factors discussed previously, EPCM is estimated at $1.1 million, itemized in 701 to 706 in Table H1 of Appendix H.

Indirects
Based on the various factors discussed previously, indirect costs are estimated at $1.0 million, itemized in 801 to 806 in Appendix H Table H1.

Owners Cost
This category covers activities such as geotechnical and environmental consultants, metallurgical testing and permanent employees hired before production start-up, etc. The anticipated cost is $2.1 million. The costs are itemized in 901 to 911 in Table H1 of Appendix H.

Allowances
Based on the various factors discussed previously, allowances are estimated at $2.1 million, itemized in 801 to 806 in Appendix H Table H1.


 
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Summary of Initial Capital Costs
The initial capital cost of the Rock Creek Project is estimated at $39.2 million, as summarized in Table 12.6.

TABLE 12.6
SUMMARY OF INITIAL CAPITAL COST

Category Cost ($ millions)
100 Infrastructure 1.7
200 Buildings and Support 2.5
300 Major Mining Equipment 10.4
300 Ancillary Support Equipment 1.0
400 Milling and Processing 16.5
500 Tailings 0.8
600 Development Rock Dumps 0.0
700 EPCM 1.1
800 Indirects 1.0
900 Owners Cost 2.1
990 Allowances 2.1
   Total 39.2

12.2

SUSTAINING CAPITAL COSTS

Since the mine life is 6 years, major mining equipment will not have to be replaced. To cover small equipment needs, an allowance of $100,000 per year has been included in the cash flow analysis. There is also an allowance of $200,000 annually to cover tailings dam construction for supply of screened filter zone material.

   
12.3

OPERATING COSTS

The operating costs are subdivided in Mining, Milling, General & Administration and Reclamation.

Mining operating costs are based on hourly costs for major equipment supplemented with cost factors for mining support equipment. Mill operating costs are based on unit rates for typical mill operations of a similar throughput size. G&A costs are based on staffing manpower costs and factors to account for other costs. Reclamation costs are estimated for annual remediation work and final reclamation at the end of the mine life.


 
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  12.3.1

Manpower

The total project manpower peak is estimated at about 123 personnel, as summarized in Table 12.7. A detailed breakdown of the staff with the associated salary structure is included in Table 12.8. A detailed breakdown of the mine and mill operators with the associated wages is included in Table 12.9 and Table 12.10.

 TABLE 12.7
MANPOWER SUMMARY

Admin Staff 17
Mining - Operating & Maintenance 60-63
Milling - Operating & Maintenance 43
                 Total Personnel 120-123

   

It is assumed that all of these people would be working at the project site however there may be opportunity to station some of them at offices in Nome. Relocation of some positions to Nome would reduce the cost for office space and other services at site.

For completeness, some staff and operating positions are identified but not filled. In these cases a decision has been made based on discussions with NovaGold to omit the position assuming that other personnel would fulfill these roles either on an intermittent, part-time, or overtime callout basis.


 
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TABLE 12.8
STAFFING DETAIL

                   # Salary Burden @
34%
Total Payroll
per person
Admin Staff        
Mine Manager                   1 $125,000 $42,500 $167,500
Mine Accountant/Controller*                    1 $70,000 $23,800 $93,800
Accounting Clerks AP/AR/Payroll*                    1 $35,000 $11,900 $46,900
Purchasing*                   1 $50,000 $17,000 $67,000
Safety Superintendent                   1 $50,000 $17,000 $67,000
First Aid/Security                   4 $40,000 $13,600 $53,600
HR/IR Manager*                   1 $70,000 $23,800 $93,800
                   10      
Technical Service        
Chief Engineer                   1 $80,000 $27,200 $107,200
Environmental Manager (shared)                   1 $90,000 $30,600 $60,300
Surveyors                   2 $50,000 $17,000 $67,000
Technician/Grade Control                   1 $50,000 $17,000 $67,000
Geologist                   1 $60,000 $20,400 $80,400
Environmentalist                   1 $50,000 $17,000 $67,000
                   7      
Total Staff                                                         17      

 
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TABLE 12.9
MINE OPERATIONS DETAIL (YEAR 1*)

                   # Wage Rate
$/hr
Annual Payroll
per person, incl
burden
Mine Operations      
General Foreman 1 $80,000 $107,000
Shift Foreman 4 $60,000 $80,400
                           Truck Drivers                   16 $21.00  
   Shovel/Loader Operators                   4 $21.00  
      Dozers Operators                   8 $18.00  
      Grader Operators                   4 $21.00  
                     Drillers/Helper                   4 $18.00  
                  Blasters/Helper                   2 $19.00  
  43    
       
Mine Maintenance      
Shift Foreman 4 $60,000 $80,400
Maintenance Planner 1 $50,000 $67,000
         Shop Mechanics                   1 $23.00  
            Field Mechanics                   4 $23.00  
               Field Electrician                   3 $23.00  
                        Field Welders                   4 $21.00  
       
  17    
   Total Mine Operations                   60    
*Workforce varies between 60 and 66, and averages 63, during the mine life.

 
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TABLE 12.10 MILL
OPERATION DETAIL

      # Wage Rate
$/hr
Annual Payroll
per person, incl
burden
Mill Operations      
General Foreman 1 $90,000 $120,600
Shift Foreman 4 $70,000 $93,800
Metallurgist 1 $80,000 $107,200
Chief Assayer 1 $53,000 $71,000
Assayer 1 $47,000 $63,000
Total staff
8    
       
       
Crusher Operator 2 $21.00  
Grinding Operators 4 $23.00  
Recovery Operator 4 $23.00  
Flotation Operator 4 $23.00  
Refiner 1 $25.00  
Day Crew 2 $21.00  
Laboratory labourer 4 $22.00  
  21    
       
Mill Maintenance      
Leadhand Millrights 1 $25.00  
Millright 3 $25.00  
Welder 4 $25.00  
Maintenance Labour 4 $21.00  
Leadhand Electricians 1 $25.00  
Electrician 1 $25.00  
  14    
Total Mill Operations 43    

 
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  12.3.2 Mining Operating Cost
       
    Mining costs were determined on an activity basis. These activity cost categories are:
       
    Loading
    Hauling
    Drilling
    Blasting
    Grading
    Dozing
       
   
These activity costs include overhaul and maintenance parts and supplies, fuel and lube, tires and wear parts (as applicable) and equipment and maintenance labour. The costs were based on hourly rates from the estimator’s guide publication, 2003 edition “Mine and Mill Equipment Costs” prepared by Western Mine Engineering Inc.. The hourly rates were factored up to account for the mine being in Alaska.These rates and the calculation methodology are shown in Table 12.11.

 
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In addition to the above activity costs, additional labour costs are included for shop workers, mining staff and indirect costs. Indirect costs are based on 25% of the staff costs. This would cover shop and office supplies, building maintenance and heating etc. These annual activity costs are shown in Table 12.12. The mining cost varies from $0.78 to $0.86/t mined, depending on the strip ratio and the total material being mined. Over the project life, the average is $0.82/t mined.

 
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  12.3.3

Mill Operating Cost

The mill operating cost has been determined based on manpower, grinding media and liners, electric power, miscellaneous costs, and maintenance supplies.

Manpower is based on a mill manpower of 43 persons as shown in Table 12.10.

Grinding media is based on a Rod work index of 14, a ball index of 16, impact work index of 7.5, and an abrasion index of 0.35. Only the ball index is based on any test work. SAG mill liners are replaced every 12 months while ball mill liners are replaced ever 18 months.

Electric power is based on a grinding power of 18.5 kwh per tonne plus an additional allowance of 5.5 kwh/t for other power consumption. Therefore each incremental increase in power cost of $0.01/kwh equates to $0.24 per tonne milled.

The breakdown for 1.78 million tonnes per year capacity is listed in Table 12.13.

TABLE 12.13
MILL OPERATING COST SUMMARY

  US$ per tonne
Manpower 1.62
Grinding media and liners 1.40
Electric Power 2.88
Miscellaneous costs 0.20
Maintenance supplies 0.30
                           Total Cost $6.40

  12.3.4

Reclamation Operating

An allowance of $$0.10/t ore is included for annual reclamation activities such as re-sloping, re-vegetation etc. In the final year, an additional lump sum of $2.0 million is included for demolition of permanent buildings and structures, and excavation and backfilling of any contaminated areas, road, development rock dump, tailings pond, and pit final reclamation.

     
  12.3.5

Administration Operating

Administration costs are based on the staffing level outlined in Table 12.8. The staff labour cost is $1.3 million. To cover indirect costs such as office supplies, building maintenance, heating, communications, insurance, local taxes etc. a factor


 
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    of 25% was applied to the staff labour cost. The total administration operating cost is then $1.7 million per year.
     
  12.3.6

Summary of Operating Costs

As can be seen in Table 12.14, the total operating cost varies between $11.58/t Ore and $12.85/t Ore. Over the project life it averages $12.08/t Ore. Based on the gold recovered, the operating cost varies between $177.13/oz and $225.94/oz and averages $191.70/oz.


 
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TABLE 12.14
OPERATING COST SUMMARY

                   
Year
Project Year
Units 2006
1
2007
2
2008
3
2009
4
2010
5
2011
6
Total Average
                   
                   
Production                  
Ore Mt 1.78 1.78 1.78 1.78 1.78 1.79 10.70 1.78
Waste Mt 7.06 8.44 9.20 8.46 6.83 7.49 47.48 7.91
Total Material Mt 8.84 10.22 10.98 10.25 8.61 9.28 58.18 9.70
                   
Gold Recovered oz 113,514 95,252 99,220 120,546 112,045 133,319 673,895 112,316
                   
Mining k$ 7,442 8,250 8,564 8,257 7,370 7,655 47,538 7,923
Milling k$ 11,405 11,405 11,405 11,405 11,405 11,426 68,450 11,408
Reclamation k$ 178 178 178 178 178 2,179 3,070 512
Admin G&A k$ 1,688 1,688 1,688 1,688 1,688 1,688 10,130 1,688
Total Operating Costs k$ 20,714 21,521 21,836 21,528 20,641 22,948 129,188 21,531
                   
Mining   $0.84 $0.81 $0.78 $0.81 $0.86 $0.83   $0.82
Milling $/t ore $6.40 $6.40 $6.40 $6.40 $6.40 $6.40   $6.40
Reclamation $/t Ore $0.10 $0.10 $0.10 $0.10 $0.10 $1.22   $0.29
Admin G&A $/t Ore $0.95 $0.95 $0.95 $0.95 $0.95 $0.95   $0.95
Total Operating Costs $/t Ore $11.62 $12.08 $12.25 $12.08 $11.58 $12.85   $12.08
                   
Mining $/oz $65.56 $86.61 $86.31 $68.49 $65.78 $57.42   $70.54
Milling $/oz $100.47 $119.73 $114.94 $94.61 $101.79 $85.71   $101.57
Reclamation $/oz $1.57 $1.87 $1.80 $1.48 $1.59 $16.34   $4.55
Admin G&A $/oz $14.87 $17.73 $17.02 $14.01 $15.07 $12.66   $15.03
Total Operating Costs $/oz $182.48 $225.94 $220.07 $178.59 $184.22 $172.13   $191.70
 

 
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13

FINANCIAL ANALYSIS

Project cash flows have been determined at a price of $325/oz, from the production schedule discussed in Section 7 and the capital and operating costs developed in Section 12. The mill will process 1.78 million tonnes/ year providing saleable gold output of about 112,300 oz/ year.

Project economics are determined on both a before and after tax stand alone project basis (possible corporate tax losses and expenses incurred before the production decision have not been considered). Royalties from aboriginal controlled lands and state mining taxes have also been calculated.

     
13.1 FINANCIAL ASSUMPTIONS
     
  13.1.1

Projections

All revenues and costs are in 2003 constant US dollars. Cash flows are discounted to July 1, 2003 at 5%.

     
  13.1.2

Net Smelter Returns

The project will produce both dore bars on site and send concentrate off site for smelting and refining. Dore bars refined on site are assumed to incur a charge of $2.00/oz to cover transportation, insurance, refinery deductions etc. to point of sale.

Concentrate sent off site is assumed to have a concentrate grade of 200 gm/DMT (Dry Metric Tonne) or 6.43 oz/DMT and a moisture content of 8%. Payment is made for 90% of the gold, with a processing charge of $100/DMT. The concentrate is assumed to be shipped out in "supersacs" at $35.00/WMT (Wet Metric Tonne) and $25.00/WMT for freight. The total deductions reduce the net gold price from concentrate to approximately $267/oz or 82% of the full price of $325/oz. About 15% of the total net smelter return (NSR) is from gold in concentrate.

     
  13.1.3

Royalties and State Mining Tax

Production from land controlled by the Bering Straits Regional Native Corporation (BSNC) is subject to an NSR royalty of 2.5% and a 5% Net Profits Interest payment. It is assumed that the net profits definition is the same as for the State of Alaska’s Mining License Tax. The actual production from BSNC lands has not been determined on an annual basis for this evaluation. It is assumed that 44% of

 
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the annual production is from BSNC land. At the Feasibility stage, it may be necessary to flag ore blocks with location to allow better tracking of royalty requirements.

Since the remainder of production is from patented mining claims owned by a NovaGold subsidiary, it is assumed that no production is from state land and so the Alaska Production Royalty on production does not apply.

All mining income in Alaska is subject to the Mining License Tax. This is a graduated tax based on the level of net income, reaching 7% above $100,000. As a new mine, Rock Creek should qualify for the 3.5 year new mine exemption from the Mining License Tax. This has been assumed in the analysis.

     
  13.1.4

Income Taxes

Federal and State income taxes have been calculated on a project basis. Assets continue to be depreciated after the project ends. A tax credit is therefore continued to be calculated after 2011 until project depreciation terminates. NovaGold may want to determine the effect of possible tax losses or incentives that might be applied to Rock Creek on a corporate basis both at the start of and at the end of the project.

Mining companies qualify for a depletion deduction. It is assumed that the percentage method will provide a greater benefit than the cost method for Rock Creek, therefore this method is used. For gold mines the rate is 15%. The actual depletion deduction is limited. It will be the minimum of 15% of gross income before depletion or 50% of taxable income before depletion.

For regular tax calculations, 70% of exploration and development costs are deducted from income tax in the year incurred and the remainder over a 60 month period. For the alternative minimum tax, the total is deducted over 10 years. It is assumed that the items detailed in Table 13.2 as exploration and development are expenses for the purposes of this calculation.

It is assumed that all hard assets, shown as infrastructure, mining equipment and allowances in Table 13.2 will be Asset Class 10.0, which has a 7 year depreciation life. It will qualify for the Modified Accelerated Cost Recovery System (MACRS). For regular tax, it is depreciated on a 200% declining basis. Under the alternative minimum tax, it is at 150%.


 
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Alaska income tax will also be reduced by an exploration incentive tax credit. It is assumed that exploration conducted by the company in 2003 qualifies for this tax credit.
     
  13.1.5

Working Capital

Working capital is included in the cash flow analysis since product and supplies inventories must be built up before revenues are realized. It is assumed that two months of cash costs at the mine will have to be covered at the start of 2006 until receivables are converted to cash. Working capital is maintained through the project life. This amount is recovered in the final year when working capital can be liquidated. It is assumed that 100% of the working capital is recovered.

     
  13.1.6

Salvage Value

The evaluation assumes there is no salvage value for the equipment.

     
13.2

MILLING OPTION COMPARISON

The project looked at a number of milling scenarios. The processing methods included gravity, gravity with flotation, gravity with cyanidation and gravity with flotation and cyanidation. Batch only and batch with continuous processing were also reviewed. In total, eight processing options, each with varying capital and operating costs and recoveries were evaluated. These options and the results are shown in Table 13.1 and Figure 13.1. The results show that Option 5, Gravity with Flotation has the best result. This option was used for all financial evaluations.


 
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13.3

OPERATING COST PROFILE

Figure 13.2 illustrates the trend in operating costs on a total dollar and $/oz basis. Both total spending and the unit cost show a declining trend until 2009, as the strip ratio declines. In 2010 the strip ratio increases to 4.65, bring the total cost up as well.


 
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13.4

CAPITAL COST PROFILE

Table 13.2 illustrates the annual capital requirements to bring the project into production and to continue through the project life. $2.2 million is required in 2004. The majority of the spending will occur in 2005, at $35.5 million ($38.9 million including working capital).


 
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TABLE 13.2
CAPITAL SUMMARY (For tax asset classes and cash flow analysis)

                         
Year Unit   2003 2004 2005 2006 2007 2008 2009 2010 2011 Total
Project Period Cost Units -3 -2 -1 1 2 3 4 5 6  
                         
Exploration   k$                    
901: Resource Definition Owners Cost 0     0                 0
                         
                         
Development                        
100: Infrastructre 1,697   k$   848 848             1,697
500: Tailings 832   k$     832 200 200 200 200 200   1,832
600: Waste Dumps 31   k$     31             31
701: Infrastructure EPCM 85   k$   42 42             85
705: Tailings EPCM 33   k$     33             33
706: Waste Dumps EPCM 1   k$     1             1
801: Infrastructure Indirects 85   k$   42 42             85
805: Tailings Indirects 42   k$     42             42
806: Waste Dumps Indirects 2   k$     2             2
990: Infrastructure Allowances 136   k$   68 68             136
994: Tailings Allowances 67   k$     67             67
995: Waste Dumps Allowances 6   k$     6             6
902: Feasibility Study Owners Cost 0   k$ 0 0               0
903: Permanent Employees Owners Cost 437   k$   219 219             437
904-906: Environmental Owners Cost 1,222   k$   764 458             1,222
907 - 911: Owners Cost 500   k$   250 250             500
                         
Total Development   k$ 0 2,234 2,941 200 200 200 200 200 0 6,174
                         
Infrastructure                        
200: Buildings and Support 2,465   k$     2,465             2,465
400: Milling and Process 16,497   k$     16,497             16,497
702: Buildings and Support EPCM 123   k$     123             123
704: Milling and Process EPCM 825   k$     825             825
802: Buildings and Support Indirects 123   k$     123             123
804: Milling and Process Indirects 660   k$     660             660
991: Buildings and Support Allowances 197   k$     197             197
993: Milling and Process Allowances 1,320   k$     1,320             1,320
                         
Total Infrastructure   k$ 0 0 22,210 0 0 0 0 0 0 22,210
                         
Mining Equipment 100.00   =New or Used Equipment Factor          
301: Front End Loaders 12 m3 1,470   k$ 0 0 1,470 1,470 0 0 0 0 0 2,940
302: Haul Trucks 90 t 1,065   k$ 0 0 4,260 0 0 0 0 0 0 4,260
303: Blasthole Drill 150 mm 646   k$ 0 0 646 0 0 0 0 0 0 646
305: Dozers 302 kW 865   k$ 0 0 1,730 0 0 0 0 0 0 1,730
306: Graders 160 kW 590   k$ 0 0 590 0 0 0 0 0 0 590
307: Water Truck 15,000 L 200   k$     200             200
304, 308-324 Other Support Equipment 1,003   k$     1,003             1,003
703: Mining Equipment EPCM 30   k$     30             30
803: Mining Equipment Indirects 50   k$     50             50
992: Mining Equipment Allowances 50   k$     50             50
Sustaining Capital           100 100 100 100 100   500
                         
Total Mining Equipment     0 0 10,029 1,570 100 100 100 100 0 11,999
                         
Allowances                        
996: EPCM 110         110             110
997: Indirects 96         96             96
998: Owners Cost 108         108             108
                         
Total Allowance     0 0 314 0 0 0 0 0 0 314
                         
Total Mining Equipment & Infrastructure   k$ 0 0 32,553 1,570 100 100 100 100 0 34,523
Contingency 0.00   k$ 0 0 0 0 0 0 0 0 0 0
Total Capital   k$ 0 2,234 35,494 1,770 300 300 300 300 0 40,698

 
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13.5

PROJECT ECONOMICS

Under these assumptions, gold price and costs, project economics are as follows:

TABLE 13.3
PROJECT ECONOMICS (Base Case)

    Before Tax After Tax
NPV@5% $20.1 million $16 million
   IRR 18.8% 16.2%

 
The project has a payback period of 4.0 years. When production commences in 2006, after tax net cash flows vary between $7.4 and $17.7 million annually, as illustrated in Figure 13.3. Most of the cash flow is generated between 2009 and 2011, when grades, hence gold recovered are the highest. Unfortunately, the impact on the economics is not as great had the higher cash flows come in the earlier years. This is also demonstrated in Figure 13.4. The graph shows the cash margin after all costs, taxes and capital is highest in those three years, being $113.94/oz in 2009 up to $132.91/oz in 2011 compared to an average of about $88.00/oz between 2006 and 2008.

 
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13.6

FINANCIAL SENSITIVITIES

Sensitivity analyses were carried out on the gold price, operating and capital costs, and the gold grade. The results on the net present value and internal rate of return are presented in figures 13.5 and 13.6.

The project economics are most sensitive to the gold price and the grade of the ore body. Each 1% change affects base case the NPV by about $1.2 million and the IRR by approximately 0.8%.

The base case NPV will change by $0.7 million and the IRR by 0.5% for each 1% change in operating cost. The capital cost changes affect the NPV by about $0.4 million and the IRR by 0.3% for each 1% increase or decrease in the capital cost.


 
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13.7

RISKS & OPPORTUNITIES

At this screening level of study, there probably is a +/- 30% level of certainty in the costs. As Rock Creek progresses towards a feasibility level study, the cost accuracy will be within a tighter range. Costs that may be lowered with further engineering or with supplier quotes or used equipment include the milling operating and capital costs and mining capital.

The milling cost would be reduced to about $6.00/t milled from $6.40/t milled if the power cost can be reduced to $0.10/kwh from $0.12/kwh. Negotiations with the power provider could achieve this or alternative electricity generation strategies may result in a lower cost.

Capital for the mill and mine development are engineering estimates based on current construction practices and costs. Actual supplier quotes may provide lower estimates. The present mill capital cost is based on typical mill construction and installation techniques done at the mine. There may be opportunities to lower the cost by having the mill engineered and constructed in modules off site. These modules could then be assembled on site quickly and at a lower cost.

The major mining equipment capital cost is based on vendor budgetary quotes for new equipment. These are typically on the high side. Depending on economic conditions at the time, suppliers may offer discounts to be competitive. There may also be very reliable used equipment available, which could be used in the mine. This would have a significant impact on the capital cost.

The average gold price in 2003 has been higher than the $325/oz used in the current evaluation. Many reputable commodity forecasters are using a long term value closer to $350/oz. CIBC World Markets estimates $350/oz in 2003, $375/oz in 2004 and $350/oz long term.

Considering the above opportunities, Rock Creek Project economics can be enhanced. If the milling cost was $6.00/t milled, initial capital was $5 million lower and a gold price of $345/oz over the project life was forecast, the internal rate of return would be 27.2 %. This cumulative impact is shown in Figure 13.7.


 
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13.8

UPSIDE TARGET SCENARIO

NovaGold geological staff indicate that there is a reasonable expectation that the block model resource and block model gold quantity at Rock Creek will increase due to additional resource drilling within the main orebody, drilling along untested extensions of the ore zone, and improved geological modeling and interpretation techniques resulting in improved ore grades. NovaGold estimates that this could result in the addition of 300,000 to 400,000 ounces of block model gold to the pit resource. Norwest has not undertaken any geological modeling to confirm these potential upside estimates since all resource modeling is the direct responsibility of NovaGold.

The incremental capital expenditure for additional ore production from Rock Creek would likely be minimal. The major mining equipment fleet would only be six years old at the end of the current mine plan and would not need replacing for another four to five years.

If one assumes a total block model gold resource of approximately 1.1 million ounces, three more years of mine production at average ore grades and average strip ratios would occur. A comparison between the base case mine plan and the upside target scenario is described below.

TABLE 13.5
UPSIDE TARGET SCENARIO

         Base Case Plan Upside Target
Scenario
Mine Life 6 years 9.5 years
Total Ore 10.7Mt 16.8Mt
Gold, insitu 699,500oz 1,100,200oz
Gold, recovered 673,900oz 1,059,100oz
                          
IRR 16.2% 21.7%
NPV (0%) $30.3 $61.8
NPV (5%) $16.0 $34.5

 
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14
RECOMMENDATIONS FOR FEASIBILITY DESIGN STUDY
     
 
The following is a summary of key recommendations that should be addressed for feasibility design and project permitting.
     
 
Geology, Resource, and Condemnation Drilling
 
Resource Classification - the current resource estimate should be classified into the categories of Measured, Indicted, and Inferred. The basis for the subsequent feasibility economics will be the Measured and Indicated resources however at this time there is no indication of how the resource is distributed amongst these categories.
     
 
Resource Infill Drilling - a certain level of resource drilling is needed to test areas where the ore zones are still open and could impact on siting of infrastructure. As well additional drilling may be needed to improve confidence in the resource estimates suitable for feasibility study purposes.
     
 
Condemnation drilling - a series of test holes should be drilled beneath potential infrastructure sites to confirm that no mineable ore exists in these areas. In addition potential southward extensions to the ore zone should be evaluated.
     
 
 Block Modelling - once the new drilling is complete, the block model must be updated. There are certain opportunities to define ore envelopes using geological controls that may improve the representativeness of the model.
     
 
Geotechnical
 
Development Rock Dumps - shallow test pit should be excavated with a backhoe in the dump foundations to confirm the foundation soils and bedrock depth. Ten to twelve pits, excavated by backhoe with a permafrost bucket if necessary, should be dug to at least four metres below ground level or to competent bedrock is shallower. Pits should be inspected and logged by a geotechnical engineer to characterize the materials and any excess ice conditions. Disturbed samples of various horizons should be collected for classification testing in the laboratory.
     
 
Pit Slopes - At least six core holes should be drilled in the proposed walls around the pit. These may be vertical or inclined into the pit at the planned wall angle to maximize representative data over the ultimate height of the wall. All cores should be oriented, photographed, and logged in detail including RQD and hardness, as well as detailed joint descriptions. Representative samples of both weathered and intact rock

 
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core should be selected for point load and unconfined compression testing as well as grain size, moisture content and Atterberg Limits as appropriate.
     
 
Tailings Area - At least four core holes should be drilled under the proposed tailings dyke alignment to characterize the weathered rock strata and at least the upper 10 metres of intact rock. These holes should also be logged and sampled. Several shallow test pits should be excavated with a backhoe in the tailings dam foundations to confirm the foundation soils and bedrock depth. Ten to twelve pits, excavated by backhoe with a permafrost bucket if necessary, should be dug to at least four metres below ground level. Pits should be inspected and logged by a geotechnical engineer to characterize the materials and any excess ice conditions. Disturbed samples of various horizons should be collected for classification testing in the laboratory.
     
 
Plant Site - At least two core holes should be drilled at the proposed plant site to characterize the weathered rock strata and at least the upper 10 metres of intact rock. These holes should also be logged and sampled.
     
  Permafrost
 
Permafrost Characterization - at least two deep and two shallow thermistors should be installed at the project site. The shallow thermistors can be used to characterize the depth of the active layer. The deep thermistors can be used to measure the depth of the permafrost in the pit area and confirm if groundwater will be a concern during mining. If groundwater is encountered, then at least two piezometers should be installed to determine groundwater pressures. It may be necessary to install electric or pneumatic piezometer if freezing of the standpipe is a concern.
     
 
Tailings Freeze-back Modelling - thermal modelling should be undertaken to assess the long-term freeze back of the tailings deposit. The amount of freezing may impact of the type of closure scenario proposed for the tailings area. At least two piezometers and thermistor strings should be installed in the tailings area.
     
  Environmental & Permitting
 
Groundwater quality - samples should be collected to determine the quality of water that may seep into the pit and will be re-used in the process.
     
  Optimization Studies
 
Project siting evaluations - as part of the engineering and regulatory process, more detailed siting evaluations may be useful to the regulatory agencies and could include

 
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    alternative analysis of site infrastructure locations, utilities supply, and avoidance of surface water impacts.
     
  Mine Planning
 
Pit Optimization - based on the updated block model and the new drilling results, new pit limits may need to be defined.
     
  Tailings Characterization
 
Acid Base Accounting - a series of ABA tests should be undertaken to characterize the acid generating potential of the milled ore. If ARD potential exists then more comprehensive kinetic tests should be undertaken.
     
 
Site Water Balance - further study should be undertaken on the site water balance to determine whether the collection of local runoff and maintaining a large water inventory in the tailings pond will be sufficient for year round mill operation.
     
  Development Rock Characterization
 
Acid Base Accounting - a series of ABA tests should be undertaken to characterize the acid generating potential of the development rock. If ARD potential exists then more comprehensive kinetic tests should be undertaken.
     
  Infrastructure
 
Power generation - a trade-off study should be done to evaluate whether on-site electric power generation is economically preferable to utilizing delivered power from the local power utility. Negotiations should be initiated to discus the anticipated power charges from the local utility given the system upgrades required to deliver power to the Rock Creek site. Confirmation of the supply voltage to Rock Creek is also required.
     
 
Diesel Fuel Supply: discussions should be initiated with local fuel suppliers to determine fuel supply logistics, including current storage capabilities, infrastructure improvements required, and fuel delivery costs.
     
 
Existing Nome Capabilities: complete a comprehensive database of construction equipment, construction and operating services, and accommodation that can be acquired in the town of Nome. This information will be useful in the feasibility design to determine the extent of local support and the need to mobilize from off-site.

 
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  Manpower
 
Staff Salaries: an informal survey of Alaskan mines should be undertaken to better estimate staff salaries that will be required to attract qualified personnel to the project.
     
 
Hourly Wages: further assessment should be done to determine local labour rates and the availability of skilled and unskilled operating labour in Nome.
     
 
Housing: further assessment should be done to determine the availability of permanent and temporary housing in Nome. This may impact on the success of attracting operating personnel and the costs associated with maintaining the construction workforce.

 
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15
  

PROJECT EXECUTION SCHEDULE

The Project Execution Schedule shown in Figure 15-1 describes the work required to undertake continued development, construction and commissioning of the Rock Creek Project. The development activities can be sub-divided into three phases:

     
  Pre-Approval Stage: this work includes the completion of feasibility level engineering and a bankable feasibility study. Other tasks include additional resource drilling programs, metallurgical testing, environmental baseline monitoring, and permitting.
     
  Project Implementation Stage: this work includes activities associated with detailed design and construction geotechnical site investigations, detailed engineering, equipment procurement, project construction and commissioning.
     
 
NovaGold Decision Points: in order to ensure timely advancement of the project certain corporate decisions or commitments must be made. A delay in making some decisions can result in delays in project advancement. It may be possible to mitigate some of these delays by committing funds for equipment procurement subject to cancellation clauses.
     
  Some of the key milestones shown in the development schedule are:

Preliminary Economic Study completed: 3rd quarter 2003
Feasibility contractor selected: 4thquarter 2003
Feasibility study completed: 2nd quarter 2004
NovaGold project decision: 2nd quarter 2004
Permitting, initiate process: 3rd quarter 2004
Permitting, completion: 3rd quarter 2005
Detailed engineering commences: 3rd quarter 2004
Major equipment and components ordered: 4th quarter 2004
Equipment delivery to site: 3rd quarter 2005
Construction start: 3rd quarter 2005
Rock Creek Project commission & production: 1st quarter 2006
   
The required start-up date for the project is January 2006. The key factor impacting on the development schedule are the open water seasons allowing barge access to bring major equipment and components to site. From today to the start of production there are three

 
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barge seasons remaining; summer 2003, 2004, and 2005. Effectively only the 2005 summer provides the most realistic shipping target. In order to freight materials to site in the summer of 2004, design engineering and equipment orders must be placed 3 to 6 months prior to that date, or in the 1st and 2nd quarters of 2004. It is envisioned that the project decision would only be made during that period. Therefore all engineering and procurement activities are geared towards meeting the summer 2005 shipping date.

The fact that all of the project components will be shipped that summer will result in significant amount of construction activity in the 3rd and 4th quarters of 2005. The logistics of maintaining and coordinating a relatively large construction workforce will require specific attention to project management. It will be important to have significant portion of the detailed engineering completed beforehand to minimize re-work or the need to fly in supplies once the open water season ends.

As well construction activities should target to allow short-term contractors to demobilize some of their equipment in the fall of 2005 or stand-by charges may possibly be incurred if demobilization can only be done the following summer.

     
15.1

OUTSIDE AGENCY RESPONSIBILITIES

Outside agencies may be required to undertake construction activities outside the envelope of site activities. Specifically these relate to:

     
 
Glacier Creek Bypass Road: this road would most likely be constructed by the State of Alaska and would be required prior to the 2005 summer shipping season. It would be difficult to transport major components to site using the existing Glacier Creek Road.
 
Powerline Upgrades: Depending on the arrangements made with the Nome power utility, a 25 kv step-up transformer and powerline upgrades would be required and would likely be undertaken by the utility. NovaGold could complete construction of the powerline from Teller Highway to the Rock Creek Project site. For construction it may be possible to maintain 12 kv power to site since the 25 kv power is only required at the start of production.
 
Diesel Storage: Depending on the arrangements made with the local fuel supply depots, some additional tankage may be required. This tankage would need to ready in the summer of 2005 so that it can be filled prior to the end of summer.

 
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15.2

CONSTRUCTION LOGISTICS

Construction of the project facilities will require specific attention to detail due to the seasonal access to Nome.

The town of Nome may have limited short-term accommodation and therefore it may be necessary to develop a temporary construction camp for construction personnel.

The availability of construction equipment in Nome is limited and therefore most of the required equipment will have to be mobilized to Nome. Timely demobilization of unneeded equipment will be important in that standby charges may be incurred if equipment is forced to winter unnecessarily in Nome in order to demobilize the following summer.

Site building will probably be pre-fabricated using Butler type buildings, foldaway structures, or Sprung structures. Lab and Offices can also be Atco type trailers.

It is envisioned that the mill would be stick built although there may some opportunities for some offsite modularization. Possibly the control rooms, MCC's, centrifugal concentrators, tables, etc. can be skid mounted. The larger SAG and Ball mills have to be mounted on concrete foundations and such mills may be a very heavy impractical lift given the crane capabilities at site.

This means that concrete pours occur in the early summer prior to the main mill construction and the fixing of flowsheets very early on to allow foundation design. The risk of a late break-up followed by an early freeze could cause problems. The subsequent winter start-up will be more difficult than a summer start-up due to erratic operation and possible tailings line freeze ups.


 
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Figure 15.1: Project Execution Schedule

 
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Dates Referenced Herein

This ‘40FR12G’ Filing    Date    Other Filings
Filed on:10/29/03None on these Dates
9/22/03
8/13/03
7/1/03
5/22/03
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