Adanac Molybdenum Corporation #200-2055

Transcription

REPORT ON
FEASIBILITY STUDY UPDATE
RUBY CREEK PROJECT,
NORTHERN BRITISH COLUMBIA,
CANADA
Submitted to:
Adanac Molybdenum Corporation
#200-2055-152nd Street
Surrey, BC V4A 4N7
Prepared by:
Rick Alexander, P.Eng
December, 2007
Ruby Creek Feasibility Study Update
December, 2007
-i-
TABLE OF CONTENTS
SECTION
PAGE
TABLE OF CONTENTS .......................................................................................... I
1.0
SUMMARY.................................................................................................. 1
2.0
INTRODUCTION....................................................................................... 11
3.0
RELIANCE ON OTHER EXPERTS .......................................................... 13
4.0
PROPERTY DESCRIPTION AND LOCATION ........................................ 14
5.0
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
AND PHYSIOGRAPHY............................................................................. 18
6.0
HISTORY .................................................................................................. 20
7.0
GEOLOGICAL SETTING.......................................................................... 24
7.1
7.2
8.0
9.0
10.0
DEPOSIT TYPES...................................................................................... 29
MINERALIZATION .................................................................................... 30
EXPLORATION......................................................................................... 32
10.1
10.2
10.3
10.4
11.0
12.0
13.0
ADANAC 2004.......................................................................................32
ADANAC 2005.......................................................................................32
ADANAC 2006.......................................................................................33
2007 Drilling Program............................................................................35
DRILLING.................................................................................................. 36
SAMPLE METHOD AND APPROACH..................................................... 40
SAMPLE PREPARATION, ANALYSES AND SECURITY ....................... 42
13.1
13.2
13.3
13.4
14.0
15.0
Regional Scale ......................................................................................24
Local and Property Scale ......................................................................26
Field Sample Preparation Procedures...................................................42
13.1.1 Adanac 2004, 2005 and 2006....................................................42
13.1.2 Prior to Adanac ..........................................................................43
LABORATORY SAMPLE PREPARATION PROCEDURES .................44
13.2.1 2004, 2005 and 2006 ADANAC.................................................44
13.2.2 1979 — 1980 Placer ..................................................................44
13.2.3 Pre-1979 Kerr Addison ..............................................................44
Analytical Procedures............................................................................45
13.3.1 Adanac 2004, 2005 and 2006....................................................45
13.3.2 1979 — 1980 Placer ..................................................................46
13.3.3 Pre-1979 Kerr Addison ..............................................................46
Quality Assurance and Quality Control..................................................46
13.4.1 Adanac Procedures ...................................................................46
13.4.2 QA/QC Results –2006 Samples ................................................47
DATA VERIFICATION .............................................................................. 49
ADJACENT PROPERTIES....................................................................... 50
December, 2007
16.0
MINERAL PROCESSING AND METALLURGICAL TESTING ................ 51
16.1
16.2
17.0
Mineral Processing ................................................................................51
Metallurgical Testing..............................................................................52
16.2.1 Kerr Addison’s Pilot Plant (Britton 1969 – 1971) .......................53
16.2.2 SGS-MinnovEX Test Work ........................................................53
16.2.3 B.C. Mining Research Limited Test Work ..................................54
16.2.4 G&T Test Work ..........................................................................54
MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE ............. 56
17.1
17.2
17.3
17.4
17.5
17.6
17.7
17.8
17.9
17.10
17.11
17.12
17.13
17.14
18.0
- ii -
The Database ........................................................................................57
The Geological Model............................................................................58
Wireframe Validation .............................................................................61
Data Preparation and Compositing .......................................................62
Declustering...........................................................................................63
Spatial Trend Analysis...........................................................................63
High-Grade Treatment...........................................................................63
Variogram Analysis................................................................................64
17.8.1 Variography Objectives and Approach ......................................64
17.8.2 Summary of Variography Parameters .......................................65
Update Block Model Parameters...........................................................65
Grade Interpolation................................................................................67
17.10.1
Grade Interpolation Methods and Objectives.................67
17.10.2
Ordinary Kriging Plan.....................................................67
Density Assignment...............................................................................68
Mineral Resource Classification ............................................................69
Mineral Resource Summary ..................................................................70
Mineral Reserve Estimate and Mine Design .........................................72
OTHER RELEVANT DATA AND INFORMATION ................................... 82
18.1
18.2
18.3
18.4
Tailings Facilities, Waste Rock Dumps and Site Water Management...82
18.1.1 Tailings Characterization ...........................................................82
18.1.2 Tailings Facility ..........................................................................82
18.1.3 Waste Rock Dumps ...................................................................84
18.1.4 Site Water Management Plan....................................................85
18.1.5 Operating and Monitoring Controls............................................86
Mine Closure Plan .................................................................................87
18.2.1 Tailings Facility ..........................................................................87
18.2.2 Waste Dump/East Ruby ............................................................87
18.2.3 Open Pit.....................................................................................88
18.2.4 Temporary Mine Closure ...........................................................88
Environmental Considerations...............................................................89
Capital Costs .........................................................................................91
December, 2007
18.5
18.6
18.7
19.0
20.0
21.0
22.0
- iii -
Operating Costs.....................................................................................94
Market....................................................................................................96
18.6.1 Supply Fundamentals ................................................................97
18.6.2 Supply Outlook ..........................................................................97
18.6.3 Demand Fundamentals .............................................................99
18.6.4 Molybdenum in Steel .................................................................99
18.6.5 Other Molybdenum Applications..............................................100
18.6.6 Substitutes ...............................................................................102
18.6.7 End-Use Industry Analysis.......................................................102
18.6.8 Price Outlook ...........................................................................103
Economic Model ..................................................................................105
18.7.1 Net Present Value and Internal Rate of Return Summary .......105
18.7.2 Sensitivity Analysis ..................................................................106
18.7.3 Summary of Results ................................................................107
INTERPRETATIONS AND CONCLUSIONS.......................................... 109
RECOMMENDATIONS........................................................................... 109
REFERENCES........................................................................................ 110
DATE AND SIGNATURE PAGE............................................................. 113
LIST OF TABLES
Table 1.1
Table 1.2
Table 1.3
Table 1.4
Table 1.5
Table 1.6
Table 1.7
Table 1.8
Table 4.1
Table 6.1
Table 11.1
Table 11.2
Table 17.1
Table 17.2
Table 17.3
Table 17.4
Table 17.5
Table 17.6
Table 17.7
February 22, 2007 Mineral Resource Estimate
2007 Updated Mineral Reserves
Capital Cost Summary
Operating Cost Summary (First Five Years of Full Production)
Operating Cost Summary (After Five Years of Full Production)
Molybdenum Price Outlook
Production and Operating Costs (First Four Years of Production)
IRR and NPV Summary Results
List of Mineral Claims
Drilling History of the Ruby Creek Deposit (Dec 31, 2006)
Drill Holes in the 2006 Ruby Creek Datamine Database
Drill Hole Data in the July 23, 2007 Mineral Resource Update
2007 Ruby Creek Molybdenum Project 2D and 3D Geometries
Primary and Secondary Lithological Units
High Grade Thresholds for % Mo by Zone
Variography Parameters for Zones 6, 60 and 7
Block Model Dimensions for Ruby Creek Resource Model
Kriging Plan Parameters
December, 2007
Table 17.8
Table 17.9
Table 17.10
Table 17.11
Table 17.12
Table 18.1
Table 18.2
Table 18.3
Table 18.4
Table 18.5
- iv -
Rock Type Bulk Density Assigned to the Block Model
November 22, 2007 Mineral Reserve Estimate
Production Schedule
Mine Department Personnel
Operating Cost Summary (First Five Years of Full Production)
Operating Cost Summary, (After Five Years of Full Production )
LIST OF FIGURES
Figure 4-1
Figure 4-2
Figure 7-1
Figure 10-1
Figure 11-1
Figure 13-1
Figure 13-2
Figure 16-1
Figure 17-1
Figure 17-2
Figure 17-3
Figure 17-4
Figure 17-5
Figure 17-6
Figure 17-7
Figure 18-1
Location Map
Mineral Claims Map1
Regional Geology Map of the Atlin Area
Plan View of 2006 Drill Hole Collars
Adanac Field Sample Preparation Procedures
Kerr Addison’s Laboratory Sample Preparation Procedure
Simplified Flowsheet
Section View of 3D Mineralization Zones and Rock Types
2006 Global Distribution of Raw Sample Lengths (Palmer, 2006)
Plan View of Block Model Geometry
Summary Schedule
Ultimate Pit Design
Phase Pushback Sequence in Plan
Phase Pushback Sequence in Section
Real Molybdenum Prices and World Supply and Demand Balance
LIST OF APPENDICES
Appendix A
Appendix B
Economic Model
Certificate of Qualification
December, 2007
1.0
-1-
SUMMARY
The Ruby Creek Molybdenum Project (“the Project”) is a large porphyry molybdenite deposit
located approximately 24 km northeast of Atlin, British Columbia, which potentially would
include an open pit mine and ore processing facility.
This report was prepared by Rick Alexander, P.Eng (Alexander) at the request of Adanac
Molybdenum Corporation (“Adanac”), in order to prepare a feasibility study to incorporate all
new information on the project including current resource and reserve estimates and to outline the
standards of disclosure required for mineral projects under National Instrument (“NI”) 43-101.
The work has included a resource estimate, mine design and the development of a metallurgical
process. This NI 43-101 compliant Technical Report has been based on work by Adanac’s in
house professional staff, G&T, SGS MinnovEX, Wardrop, Golder, Klohn Crippen, and CPM
Group.
Discovered in 1905, it has been explored on several occasions (between 1968 to 1981) but failed
to advance into production on account of molybdenum prices holding in the US$2.00 - US$4.00
per pound molybdenum range during that period. The present owners, Adanac, acquired a 100%
interest in the Project, (no royalties) in 2002 through staking 189 units, which covers the upper
southwest part of Ruby Creek Valley and much of the adjacent Boulder Creek Valley. Six
mining claims were converted into a Mining Lease for development on March 27, 2007.
In April 2005, Adanac commissioned a team of engineering consultants to complete the
component studies of a Preliminary Feasibility Study (as defined by NI 43-101) on the Project.
Based on the results of that study, it was recommended that the project proceed to the Feasibility
Study level. As such, in September 2005, the following engineering consulting companies were
commissioned to complete the component studies for a Feasibility Study (as defined by NI 43101):
•
Wardrop Engineering Inc. (Wardrop):
-
•
Feasibility Study Report, 2006 (“the 2006 Feasibility Study”)
Golder Associates Ltd. (Golder):
-
Ruby Creek Molybdenum Project - Mining Feasibility Study, 2006
-
Technical Report - Mineral Resource Estimate Ruby Creek Molybdenum Project,
February 8, 2006
-
Pit Slope Stability Considerations for the Ruby Creek Project, 2006
December, 2007
•
Wardrop Engineering Inc.:
-
•
-2-
Feasibility, Process and Infrastructure Design and Cost Estimate, Ruby Creek, 2006
Klohn Crippen Berger Consultants Ltd. (Klohn Crippen):
-
Feasibility Design of Tailings Facility, Waste Dumps and Site Water Management,
Ruby Creek Project, 2006
Alexander has revised the original 2006 Feasibility Study Report to incorporate new and updated
information provided by the authors of the component studies, which include:
•
Adanac:
-
•
•
Klohn Crippen:
-
Site Water Management Design Report, February 2007;
-
Tailing Facility Detail Design Report, April 2007
G&T Metallurgical Services Ltd (G&T):
-
•
An Assessment of Metallurgical Response, Ruby Creek Project, December 2006
CPM Group (CPM Group):
-
•
Internal reports prepared by professional engineers employed by Adanac and
reviewed by independent qualified persons (QP) to confirm the work was performed
in accordance with good and accepted engineering practice.
Sustainability of Recent Molybdenum Prices, a Molybdenum Industry Analysis,
October 2007 (“the October 2007 CPM Marketing Report”)
Golder:
-
Mineral Resource Estimate Update, Ruby Creek Molybdenum Project, July 2007
This revision to the 2006 Feasibility Study is entitled “Feasibility Study Update, Ruby Creek
Project, Northern British Columbia, Canada” (“the Feasibility Study Update”).
The Project is planned as an open pit mining operation with on-site ore beneficiation. The mineral
resource is a porphyry molybdenum deposit in multiple phases of felsic intrusions. Based on the
ore characteristics and reserve magnitude, the mill has been designed to operate at an average rate
of approximately 23,000 metric tonnes per day for a life of 21 years. The site has no developed
infrastructure with the exception of a single lane access road.
December, 2007
-3-
Mining and processing are scheduled on a 24 hour per day, 7 days per week schedule. The
Project will be operated as a fly in/fly out camp accommodation program, with crews working on
a two week turn-around basis. The operating management and labour positions are expected to
peak at 250 employees. Whereas it is located at an elevation of some 1,400 metres above sea
level, the Project is also easily accessible by road from Atlin.
Golder completed an NI 43-101 compliant Technical Report entitled “Technical Report - Mineral
Resource Estimate, Ruby Creek Project, British Columbia” and dated February 8, 2006 (“the
February 8, 2006 Golder Technical Report”), which describes the geology and this was the
second time that Adanac reported resources to the public. Golder then completed an update to the
February 8, 2006 Golder Technical Report estimate to incorporate the results of the 2006 drilling
program. An updated Mineral Resource Estimate for the Ruby Creek Property was provided to
Adanac on February 22, 2007 as a letter report (“the February 22 Mineral Resource Estimate”),
which was reproduced in Golder’s report entitled “Mineral Resource Estimate Update, Ruby
Creek Molybdenum Project” and dated July 23, 2007 (“the July 23, 2007 Mineral Resource
Update”). The July 23, 2007 Mineral Resource Update does not include any drilling information
collected during the 2007 drilling exploration program for the Ruby Creek Project. The July 23,
2007 Mineral Resource Update was used in the 2007 Feasibility Study Update .
The February 22, 2007 Mineral Resource Estimate, as reported in the July 23, 2007 Mineral
Resource Update, is tabulated by cut-off grades from 0.02 to 0.1 % molybdenum (%Mo) for
measured, indicated and inferred mineral resource categories and is summarized in Table 1.1.
December, 2007
-4-
Table 1.1
Resource Category
Measured
Indicated
Measured
+
Indicated
Inferred
Cut-off (%Mo)
Tonnage
%Mo
Mo lb
0.020
0.030
0.040
0.050
0.060
0.070
0.080
0.090
0.100
0.020
0.030
0.040
0.050
0.060
0.070
0.080
0.090
0.100
0.020
0.030
0.040
0.050
0.060
0.070
0.080
0.090
0.100
0.020
0.030
0.040
0.050
0.060
0.070
0.080
0.090
0.100
55,831,000
54,300,000
49,106,000
41,389,000
30,151,000
21,909,000
14,556,000
10,411,000
6,612,500
387,278,000
238,954,000
163,801,000
109,444,000
61,471,000
37,664,000
18,813,000
9,848,100
4,286,900
443,108,000
293,254,000
212,907,000
150,834,000
91,621,000
59,573,000
33,369,000
20,259,000
10,899,000
135,737,000
48,456,000
24,973,000
11,631,000
5,194,300
2,626,200
1,239,700
821,310
493,790
0.068
0.069
0.073
0.078
0.086
0.094
0.104
0.112
0.122
0.042
0.052
0.060
0.067
0.077
0.084
0.094
0.102
0.113
0.045
0.055
0.063
0.070
0.080
0.088
0.098
0.107
0.118
0.032
0.045
0.054
0.064
0.077
0.089
0.103
0.113
0.125
83,698,000
82,600,000
79,029,000
71,172,000
57,165,000
45,403,000
33,374,000
25,706,000
17,785,000
358,593,000
273,935,000
216,669,000
161,658,000
104,350,000
69,749,000
38,987,000
22,145,000
10,680,000
442,290,000
356,535,000
295,699,000
232,831,000
161,513,000
115,151,000
72,360,000
47,851,000
28,464,000
95,759,000
48,072,000
29,730,000
16,411,000
8,817,500
5,152,900
2,815,000
2,046,000
1,360,800
Cut-off %Mo grades were classified as greater than or equal to and range from 0.01 to 0.10 in increments of 0.01. The
mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources
and Mineral Reserves.
December, 2007
-5-
The combined Measured and Indicated Mineral Resource estimate using a 0.04 %Mo cut-off is
212,907,000 metric tonnes (T) with a grade of 0.063 %Mo and 295,699,000 pounds of
molybdenum. Comparing the February 22, 2007 Mineral Resource Estimate to the previous
estimate in the February 8, 2006 Golder Technical Report shows an increase of some 6,532,000
tonnes of measured and indicated resources and an increase of 10,095,000 pounds of contained
molybdenum, with no significant change in grade. The changes from the February 8, 2006
Mineral Resource Estimate is attributed to the 2006 inclined drilling program (with respect to
grade and tonnage) and re-interpretation of the 2006 mineralized geometries (affecting tonnage
only).
An updated Mineral Reserve Estimate was developed based on the February 22, 2007 Mineral
Resource Estimate and on a new and updated mine design, using a 0.04 %Mo mining grade cutoff and a 0.03 %Mo milling grade cut-off. This is provided in Table 1.2.
Table 1.2
2007 Updated Mineral Reserve
Phase 1
Phase 2
Phase 3
Phase 4
Total
Ore to Mill
Proven
Probable
Tonnes
%Mo
Tonnes
%Mo
19,455,000 0.089 3,065,000 0.082
3,903,000 0.070 6,819,000 0.075
20,250,000 0.056 48,463,000 0.050
271,000
0.049 29,166,000 0.056
43,879,000 0.072 87,513,000 0.055
Stockpile Ore
Proven
Probable
%Mo
Tonnes
%Mo
Tonnes
6,081,000
0.049
2,608,000 0.042
4,996,000
0.049
7,519,000 0.046
185,000
0.027
2,893,000 0.026
12,000
0.027
1,999,000 0.026
11,274,000 0.049 15,019,000 0.039
Total
Tonnes
31,209,000
23,237,000
71,791,000
31,448,000
157,685,000
Upon completion of open pit mining of Phases 1 to 4 an estimate of 157,564,000 tonnes with a
grade of 0.058 %Mo is extracted based on the updated mine design.
In their 2006 work, Wardrop made use of detailed pilot plant information developed by two of the
previous operators of the property (Kerr Addison Mines Limited and Placer Development
Limited), and additional documentation provided by Adanac and its consultants. The information
available from the earlier work (1969 to 1981) was updated and supported by supplementary data
developed in years 2004 and 2005. SGS-MinnovEX Technologies Inc. (MinnovEX) was
contracted by Adanac to develop comminution and flotation studies to be used as the basis for
updated mill process design. These were appended to the 2006 Feasibility Study.
Klohn Crippen has completed the detailed design of a compacted cyclone sand tailings dam for
the Project, which is expected to be sufficient to support the planned operations over the duration
of mine life. It will be sited downstream of the proposed mill site location. Waste dumps have
been sited so as to be compatible with plans for surface water management, including a seepage
%Mo
0.077
0.059
0.051
0.054
0.058
December, 2007
-6-
recovery dam and pond that is located downstream of the main dam structure. Testing of tailings
supernatant water shows that most elements meet the British Columbia Water Quality Guidelines
(BCWQG) for the protection of freshwater aquatic life.
The 2006 Feasibility Study was based on a 20,000 tonnes per day conventional mill process, fed
with ore from an open pit mine. While this overall design concept has not changed in the 2007
Feasibility Study Update, the on-going detailed engineering is considering an average milling rate
of 1100 tonnes per hour, as opposed to the previous rate of 906 tonnes per hour.
To take advantage of the expected higher metal prices in the early years of the Project, an initial
phase of the mining operation has been included in the mining schedule which will maximize
molybdenum production during the first four years of operation. During this phase of the mining,
the mine cut-off grade was raised to 0.06 %Mo. However, material grading between 0.04 and
0.06 %Mo would be stockpiled for processing at a later date. By processing a higher grade feed to
the mill during this phase, net revenues are maximized allowing a faster pay-back of the initial
capital investment.
The total initial capital cost for the development of the Project is estimated to be CDN$640
million. The cost estimate has been carried out to an accuracy of +15%. The estimate was
updated from the 2006 Feasibility Study.
A summary of the major capital costs is shown in Table 1.3.
Table 1.3
Description
Project Development and Infrastructure
Estimated Cost (CDN$)
39,700,000
Facilities Construction and Commissioning
341,300,000
Materials and Equipment
134,500,000
Engineering and Project Management
Subtotal
48,800,000
564,300,000
Contingency
55,600,000
Construction Risk
20,100,000
Total
640,000,000
Mining costs were developed by Golder and Adanac from first principles and with various
equipment suppliers to determine the most appropriate operating costs for the mining operations.
Estimated hourly equipment operating costs developed were compared to actual costs at similar
December, 2007
-7-
mines. Labour costs were estimated from information collected from other similar mining
operations.
The mine has been designed and costed as an owner-operated mine. The average unit mining
cost was determined to be CDN$1.39 per tonne mined or CDN$3.94 per tonne of ore milled, for
the first five years of full production and CDN$2.47 for the remaining years of operation.
Process operating supply costs are based on budgetary prices from vendors of the consumables
and reagents.
The costs for General and Administration (G&A) includes mine management, transport,
insurance, warehouse and security personnel and general management.
Tailings dam maintenance costs have been estimated from published costs from other mines
using similar construction techniques.
The total operating cost for mining is estimated to be CDN$13.08 per tonne of ore milled for the
first five years of full production and CDN$8.11 per tonne of ore milled for the remaining years
of operation. Commissioning and pre-stripping activities completed in Year 1 have been
accounted for in the capital cost estimates. The operating costs for mining, processing, power,
tailing dam operation, general administration and Adanac’s related cost were prepared by in
house professional engineers and reviewed by independent qualified persons to confirm that the
work conforms to good engineering practice. Tables 1.4 and 1.5 present the operating cost
summary for first four years and after five years of full production to end of life (year 21).
Table 1.4
Operating Cost Summary
(First five years of full production)
Description
Operating Cost
(CDN$/tonne of ore milled)
Mining (average)
Processing
Power
Tailings Dam Operation (average)
G&A
Owner Cost
3.94
2.40
5.57
0.26
0.10
0.81
Total Average Operating Cost:
13.08
December, 2007
-8-
Table 1.5
(After Five Years of Full Production)
Description
Operating Cost
(CDN$/ tonne of ore milled)
Mining (average)
Processing
Power
G&A
Owner Cost
2.47
2.40
2.25
0.26
0.10
0.63
8.11
CPM Group prepared a detailed market study and price forecast for molybdenum in US$/lb in
their October 2007 CPM Marketing Report. This is presented in Table 1.6.
Table 1.6
Year
Price per pound (US$)
2008
34.00
2009
32.25
2010
28.00
2011
23.00
2012
21.75
2013
19.50
2014
16.00
2015
15.00
2016 onwards
14.75
Estimates of production, operating costs, projected metal prices and possible gross margins for
the first four years of production are summarized in Table 1.7.
December, 2007
-9-
Table 1.7
Production and Operating Costs (First Four Years of Production)
Year
lbs Mo x 106
Total
Cost/Tonne
(CDN$)
Total
Cost/lb Mo
(US$)
Projected
Price
(US$/lb)
Projected Gross
Margins (US$/lb)
2009
8.143
13.48
7.42
32.25
24.83
2010
13.816
12.25
6.29
28.00
21.71
2011
12.463
12.38
6.87
23.00
16.13
2012
12.147
13.68
7.78
21.75
13.97
After the initial four years of production, the mine cut-off grade is predicted to reduce to
0.04 %Mo and the operating plan will focus on maximizing overall project cashflow.
Adanac's goal is to become a producer as soon as possible as the timing of the project is critical to
fully take advantage of the current high molybdenum prices. The decision has been made to
initially use diesel-electric power generation in order to expedite the proposed development
schedule. This has added significantly to both capital and operating costs but enables the project
to start-up at least two years earlier than otherwise could be achieved. Adanac believes that
connection to grid electric power from Yukon will occur by the end of 2013.
Salient points of the overall study are:
• Mine Life:
21 years
• Milling Rate:
23,000 tonnes per day
• Strip Ratio:
1.11 (tonne waste)/1.0 (tonne ore)
• Tonnage Milled:
157.6 million tonnes, average grade 0.058%Mo
• Molybdenum in concentrate:
81.7 million kilograms
• Pre-production Capital:
CDN$640.0 million
• Average Operating Cost:
US$7.60/lb Mo for first five years of full
production and. US$7.99/lb Mo afterwards
A pre-tax economic model has been developed from the estimated costs and the open pit
production schedule. The Base Case has an Internal Rate of Return (IRR) of 18.9% and a Net
Present value (NPV) of CDN$295.0 million, at an 8% discount rate, with a 21-year mine life.
The payback of the initial capital is 3.2 years.
December, 2007
- 10 -
The timing of the Project is critical if Adanac is to take advantage of the higher molybdenum
prices early in the scheduled mine life. Additional economic sensitivities were run for variations
in price, capital cost and operating cost, and the comparative indicators are summarized in
Table 1.8.
Table 1.8
Case Description
IRR
NPV @ 8%
(CDN$
millions)
Payback Period
(Years)
18.9%
295.0
3.2
Historical Average Mo Price (Last 3 years’)
30.3%
1,014.7
2.9
Low Case Mo Price Scenario
12.3%
120.7
5.9
High Case Mo Price Scenario
24.6%
444.1
2.6
Capital Cost +15%
14.8%
213.0
3.8
Capital Cost -15%
24.8%
377.2
2.6
Operating Cost +15%
15.2%
185.6
3.5
Operating Cost -15%
22.4%
404.5
2.9
Specific Economic Sensitivities
Mine operation with hydroelectric power starting
earlier in year 3 (Base Case is year 6)
19.5%
325.0
2.6
Increase in in-situ grade by 15%
27.6%
533.4
2.4
Base Case
Sensitivities
Detailed engineering and procurement have been underway since November 2006. Adanac
announced a positive production decision subsequent to the receipt of the Environmental
Assessment Certificate on September 10, 2007 from the Province of British Columbia. Adanac is
in the process of arranging equity and debt financing to build the mine.
Based on the results of this analysis, the Project contains a valuable molybdenum-bearing mineral
resource that can be economically extracted using proven mining methods and processing
technologies, at current labour, equipment and material costs and also based on the projected
prices of Molybdenum in the future.
It is recommended that Adanac continue to develop the Project through detailed engineering and
construction.
December, 2007
2.0
- 11 -
INTRODUCTION
Adanac Molybdenum Corporation (“Adanac”) is currently evaluating the development of the
Ruby Creek Molybdenum Project (“the Project”). The Project is planned as an open pit mining
operation with on-site ore beneficiation. The mineral resource is a porphyry molybdenum deposit
in multiple phases of felsic intrusions. Based on the ore characteristics and reserve magnitude,
the mill has been designed to operate at an average rate of approximately 23,000 metric tonnes
per day for an life of at least 21 years. The site has no developed infrastructure with the
exception of a single lane access road.
Mining and processing are scheduled on a 24 hour per day, 7 days per week schedule. The
project will be operated as a fly in/fly out camp accommodation program, with crews working on
a two week turn-around basis. The operating management and labour positions are expected to
peak at 250 employees.
This report was prepared by Rick Alexander (“Alexander”) at the request of Adanac, in order to
prepare a feasibility study to incorporate all new information on the project including current
resource and reserve estimates and to outline the standards of disclosure required for mineral
projects under National Instrument (“NI”) 43-101. The work has included a Mineral Resource
estimate, mine design and the development of a metallurgical process. This NI 43-101 compliant
Technical Report has been based on work by Adanac’s in-house professional staff, G&T, SGS
MinnovEX, Wardrop, Golder, Klohn Crippen, and CPM Group.
The report updates and summarizes the technical content from several previous reports, as
follows:
•
Golder Associates Ltd., Ruby Creek Molybdenum Project Mining Feasibility Study,
March, 2006 (Qualified Person: Kirk Rodgers, P.Eng.).
•
Golder Associates Ltd., 2007 Mineral Resource Estimate Update Ruby Creek
Molybdenum Project in British Columbia, Canada, July 23, 2007 (Qualified Person Paul Palmer, P.Eng., P.Geo.).
•
Golder Associates Ltd., Pit Slope Stability Considerations for the Ruby Creek Project,
Adanac Moly Corp, Atlin, BC, February 08, 2007 (Qualified Person - Al Chance,
P.Eng.).
•
Wardrop Engineering Inc., Ruby Creek Feasibility, Process and Infrastructure Design and
Cost Estimate, March 2006 (Qualified Person - Rick Alexander, P.Eng.).
•
Klohn Crippen Consultants Ltd., Ruby Creek Project — Feasibility Design of Tailings
Facility, Waste Dumps and Site Water Management, February 8, 2006 (Qualified
Person-Howard D. Plewes, P.Eng.).
December, 2007
- 12 -
•
Klohn Crippen Consultants Ltd., Ruby Creek Project — Site Water Management Design
Report, February19, 2007 (Qualified Person-Howard D. Plewes, P.Eng.).
•
Tailings Facility Detailed Design Report, Kohn Crippen Berger Ltd., April 5, 2007
(Qualified Person-Howard D. Plewes).
Golder has also performed additional technical work in the area of mineral resource estimation.
Alexander the Qualified Person within the Meaning of NI 43-101 has prepared, or supervised the
technical matters covered by this report.
A site visit inspection was completed by Alexander on June 21, 2005 for one day and on several
occasions between February and December 2007, site visits by Golder qualified person’s have
been completed in 2005, 2006 and 2007 to review the data collection procedures and a review the
results from the sampling quality assurance and quality control (QA/QC) program for
molybdenum and trace element assaying that is completed by Adanac.
Currencies are expressed in Canadian Dollars unless identified otherwise. (United States Dollars:
US$.
December, 2007
3.0
- 13 -
RELIANCE ON OTHER EXPERTS
I have relied upon, and believe that I have a reasonable basis to rely upon, the marketing
information contained in the molybdenum market study and economic model prepared for
Adanac by CPM Group.
With respect to the ownership of the mineral and placer tenures described in Section 4.0 –
Property Description and Location, I have relied upon, and believes that I have a reasonable basis
to rely upon, the title opinion of Fraser and Company LLP, (November 30, 2007).
December, 2007
4.0
- 14 -
PROPERTY DESCRIPTION AND LOCATION
The Ruby Creek property is in the Atlin Mining Division and originally consisted of a single,
irregularly shaped block of 20 claims, comprising 189 units, covering the upper southwest part of
the Ruby Creek valley and much of the adjacent Boulder Creek valley. The claims (numbered
510307, 510308, 510309, 510310, 510315 and 530317) were converted into a Mining Lease on
March 27, 2007 to facilitate project development.
Subsequently, Adanac has acquired selected mineral and placer tenures adjacent to the Mining
Lease or in it’s vicinity, to compliment its original holdings. All mineral tenures are 100% owned
by Adanac and are not subject to any royalties or carried interests.
The property is located 24 km northeast of Atlin, British Columbia. Atlin is 175 km southeast of
Whitehorse, Yukon Territory. The approximate geographic centre of the original claims on the
National Topographic 1:50,000 Map Sheet 104N/11 is universal transverse mercator (“UTM”)
Zone 8, 590,000 m east and 6,620,000 m north. Figure 4-1 shows the location map of the project.
Six claims were converted into a Mining Lease on March 27, 2007.
Figure 4-1
Location Map
December, 2007
- 15 -
Fraser and Company LLP (“Fraser”) is of the opinion that the recorded owner and expiry
periods for the mineral and placer claims are as listed in Table 4.1.
Table 4.1
List of Mineral Claims
December, 2007
- 16 -
* 140154 is the owner number of the Corporation, who is the recorded owner of these claims.
Fraser’s search does not indicate any recorded liens, encumbrances or agreements against the
above claims.
The records of the MTO (which is the source of this information) are subject to the provisions of
the Mineral Tenure Act (British Columbia) and the regulations thereunder.
Fraser confirms that a mining lease over the following six mineral tenures has been issued as of
March 27, 2007 and granted for an initial term of 30 years. The mining lease number is 555153:
- 510307
- 510308
- 510309
- 510310
- 510315
- 510317
The title can be affected by a number of other factors, which the Mineral Titles Office would
have no record of and there is no statutory requirement to record with the Mineral Titles Office
bills of sale, agreements or other documents that may affect title to the mineral claims.
The original mineral claim package is shown in Figure 4-2. The main mineralization is located
on the Mining Lease in the northeast area of these claims. This information was provided by
Adanac.
December, 2007
- 17 -
Figure 4-2
Mineral Claims Map1
N
Approximate Ruby Creek
Property Boundary
1
Provided by Adanac
December, 2007
5.0
- 18 -
ACCESSIBILITY, CLIMATE, LOCAL RESOURCES,
INFRASTRUCTURE AND PHYSIOGRAPHY
The resource is a molybdenum porphyry deposit located at the headwaters of Ruby Creek, which
flows south into Surprise Lake. The property lies at an elevation of about 1,400 metres, and the
deposit underlies a relatively unvegetated alpine cirque, near the head of the valley. The walls are
moderately steep but the floor is glacially scoured and flat. The deposit is easily accessed by
40 km of road from Atlin. The first 19 km to the bridge at Surprise Lake is fully maintained; the
remaining 21 km is a single lane exploration road that is not maintained.
Other than road access, there is no infrastructure on the property other than a core logging
building and core storage racks. The town of Atlin, located 24 km southwest of the property, is a
source for fuel, groceries, accommodation, charter aircraft services, etc. Atlin is accessible by a
two-hour drive on an all-weather road south from Whitehorse, Yukon, the territorial capital and
major supply centre for the region. Whitehorse is serviced by daily commercial flights to
Vancouver and other cities.
The Ruby Creek area lies east of the Coast Range Mountains and within a zone generally
described as having an interior type of climate. In general, the winters are severe and the summer
months are cool. Summer is enhanced by long hours of daylight; during June and July, daylight
lasts up to 20 hours.
Temperature was measured concurrently at Atlin and Ruby Creek in 1979 and 1980. A weather
station was installed on the property (June 2004) to provide meteorological information needed
for environmental baseline studies.
Atlin long-term temperature records indicate:
•
Mean annual temperature is about 1°C;
•
Mean daily temperatures are above freezing from April to October;
•
Temperatures range from -50°C to +31°C; and
•
From the limited comparative weather information (Atlin and Ruby Creek) the site has:
-
Mean annual temperature that is approximately -2°C, which is 3°C lower than Atlin;
Mean summer temperatures that are about 5°C lower than Atlin;
Mean daily temperatures (June to September) are usually above freezing; and
Temperature extremes that range from -53°C to +28°C.
Freezing temperatures can be encountered at any time of the year in Atlin and Ruby Creek.
December, 2007
- 19 -
Precipitation data at the Project site is available for the summer months of 1979 and 1980, and
from the weather station installed in June 2004. Comparison of the summer rainfall at Atlin and
Ruby Creek during 1979 and 1980 indicated the Ruby Creek had 1.6 to 2.2 times the rainfall
measured at Atlin. The estimated mean annual precipitation at Ruby Creek is 715 mm.
The surface rights cover an approximate area of 24.65 square kilometres and therefore is of
sufficient size to allow mining operations to be carried out.
December, 2007
6.0
- 20 -
HISTORY
The history of the Ruby Creek deposit has been provided by Pinsent (2005) and previous
Technical Reports (Blower, 2005; and Palmer, 2006 and 2007). Excerpts from those reports are
included below.
The Ruby Creek deposit has been known since 1905, but underwent only limited exploration until
1966, when it was staked by Adanac Mining and Exploration Limited (no relation to Adanac) and
Canadian Johns Manville Limited. Adanac Mining and Exploration Limited acquired the
controlling interest the following year and cored an aggregate length of 12,775 m in 80 drill
holes. It optioned the property to Kerr Addison Mines Limited (Kerr Addison), in 1970.
Kerr Addison completed a further 47 diamond drill holes for a total length of 5,626 m and drove
589 m of drift, 246 m of cross-cut and 281 m of raise in the high-grade core of the deposit. It
extracted 9,545 tonnes of mineralization, primarily from six raises and processed it on site to
assess the significance of a well-defined nugget effect caused by coarse-grained molybdenite
mineralization and confirm a viable metallurgical process. Chapman, Wood and Griswold
Limited completed a feasibility in 1971 and deemed the deposit un-economic on account of
prevailing low prices for molybdenum (US$1.90/lb). As a result, Kerr Addison dropped its
option.
The following year, 1973, Climax Molybdenum Corporation of British Columbia Limited
(Climax) drilled and/or deepened a further 9 diamond drill holes for an aggregate length of
2,672 m and developed the first comprehensive geological model for the deposit (White et al.,
1976). Climax dropped its option and thereafter, the property remained dormant until metal
prices improved in the late 1970s.
In 1978, Placer Development Limited (“Placer”) optioned the property and initiated a full-scale
technical and socioeconomic evaluation. In 1979, they completed a further 6,028 m of diamond
drilling in 49 holes in-and-around Kerr Addison's proposed initial pit area, and the following year
drilled a further 27 diamond drill holes with a total length of 4,858 m, in the ultimate pit area.
Although Placer carried out nearly all of the work required for a feasibility study, the price of
molybdenum dropped sharply in 1983 and the project was shelved. The company held on to the
option for a few years but eventually returned it to Adanac Mining and Exploration Limited.
The claims lapsed in the late 1990s and Andris Kikauka staked the deposit for Adanac Gold Corp,
now Adanac Molydenum Corp. (“Adanac”) in 2002. The following year, Adanac started to
compile the considerable amount of historic data that documented the previous work on the
property. Following a positive scoping study early in 2004, Adanac did further drilling on the
property. The objectives were to fill gaps in existing data distribution, to get new assay
December, 2007
- 21 -
information in order to assess the quality of the old assay data and to improve the understanding
of the deposit. The programme was designed with input from AMEC Americas Ltd. (“AMEC”)
and, later in the same year, Adanac completed 38 diamond drill holes totalling 9,087 meters.
In April 2005, the mineral resources for the Ruby Creek deposit were outlined in the Technical
Report – Mineral Resource Estimate Ruby Creek Project (Blower, 2005) by AMEC.
During 2005, Adanac completed a 19 diamond drill hole program, totalling 4,984 m, based on
recommendations in the Technical Report – Mineral Resource Estimate Ruby Creek Project
(Blower, 2005) and with additional input from Golder. Of the 19 diamond drill holes, 7 were
drilled within the proposed pit area to infill past campaigns, obtain metallurgical samples and 12
[including five dual purpose holes (exploration / geotechnical)] were drilled on the extremities of
the property to delineate the fringes of the deposit and to provide geotechnical data for the
proposed pit slope design. The geotechnical holes included two holes drilled in the south wall,
one hole in the west wall, one hole in the northwest wall, and one hole in the north to intersect the
Adera Fault area
During 2006, Adanac completed a 16 diamond drill hole program totalling 3,921 m in and around
the previously established deposit (central deposit area) as well as the south and southwest edges
of the deposit. Three of the holes were located on the fringes of the deposit to improve Adanac’s
understanding of the overall shape, depth and extremities of the deposit. The remaining 13 holes
were located to infill past campaigns and drilled at approximately -50° dip and 270° Azimuth. In
addition, during the 2006 drilling program, 8 drill holes (AD-357, AD 359, AD-361, AD-363,
AD-364, AD-366 to AD-368) from the inclined drilling program were surveyed with an optical
televiewer camera to determine dominant dip and dip directions of the mineralized veins.
The 2006 database for the Ruby Creek deposit is composed of 266 drill holes with a total of
46,912 m of drilling information. Between the years of 2004 and 2006, Adanac has drilled a total
of 73 of the 266 drill holes.
In 2006 a NI 43-101 Technical Report was completed based on the 2005 data and earlier and was
entitled “Mineral Resource Estimate Ruby Creek Molybdenum Project” and dated February 8,
2006 (“February 8, 2006 Mineral Resource Estimate”).
In late 2006 and early 2007 the 2006 Datamine Database was used in developing the updated
Mineral Resource Estimate for the property and is entitled “2007 Mineral Resource Estimate
Update Ruby Creek Molybdenum Project in British Columbia, Canada”, dated July 23, 2007
(July 23, 2007 Mineral Resource Update”). The July 23, 2007 Mineral Resource Update was
used in the development of this Feasibility Report which is based on 266 drill holes (71 from
Adanac) drilled between 1966 and 2006.
December, 2007
- 22 -
Outlined in Table 6.1 is the previous NI-43-101 2006 Mineral Resource Estimate for Ruby Creek
based on the 2006 Technical Report (Palmer, 2006).
December, 2007
- 23 -
Table 6.1
January 2006 Mineral Resource Estimate
Ruby Creek Molybdenum Project
Resource Category
Measured
Indicated
Measured
+
Indicated
Inferred
Cut-off (%Mo)
Tonnage
%Mo
Mo lb
0.020
40,636,000
0.077
68,982,000
0.030
40,386,000
0.077
68,557,000
0.040
38,942,000
0.079
67,822,000
0.050
35,834,000
0.081
63,990,000
0.060
28,836,000
0.088
55,942,000
0.070
22,593,000
0.094
46,820,000
0.080
15,225,000
0.104
34,907,000
0.090
10,601,000
0.112
26,175,000
0.100
6,908,000
0.121
18,427,000
0.020
432,936,000
0.041
391,325,000
0.030
261,618,000
0.051
294,149,000
0.040
167,433,000
0.059
217,782,000
0.050
113,435,000
0.066
165,052,000
0.060
63,101,000
0.076
105,725,000
0.070
37,572,000
0.083
68,751,000
0.080
17,543,000
0.094
36,355,000
0.090
8,808,000
0.103
20,001,000
0.100
4,151,000
0.113
10,340,000
0.020
473,572,000
0.044
460,307,000
0.030
302,004,000
0.054
362,706,000
0.040
206,375,000
0.063
285,604,000
0.050
149,269,000
0.070
229,042,000
0.060
91,937,000
0.080
161,667,000
0.070
60,165,000
0.087
115,571,000
0.080
32,768,000
0.099
71,262,000
0.090
19,409,000
0.108
46,176,000
0.100
11,059,000
0.118
28,767,000
0.020
151,326,000
0.034
113,429,000
0.030
61,837,000
0.048
65,437,000
0.040
33,067,000
0.060
43,740,000
0.050
23,225,000
0.067
34,305,000
0.060
13,375,000
0.076
22,409,000
0.070
6,490,000
0.088
12,591,000
0.080
3,166,000
0.102
7,120,000
0.090
1,915,000
0.113
4,771,00
0.100
1,143,000
0.124
3,124,000
mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources and
Mineral Reserves.
December, 2007
7.0
- 24 -
GEOLOGICAL SETTING
The regional and local geological setting descriptions of the Ruby Creek deposit are provided in
reports by Pinsent (2005), Blower (2005) and Golder’s Mineral Resource Estimate Reports, (2006
and 2007). Excerpts from those reports are provided below.
In Golder’s opinion, the geology of the Ruby Creek deposit and the mineralization controls are
sufficiently well understood and sufficiently reliable to be used in the development of the
resource estimation.
Additional analysis has been completed during the 2006 drilling program in order to better
understand the sub-vertical and sub-horizontal vein orientations and molybdenum mineralization
using in-hole optical televiewer cameras. The 2006 drilling results indicated that high grade
mineralization did occur along both sub-vertical and sub-horizontal veins. These vein orientation
results were also applied to the mineral resource estimate process. However, more inclined
drilling data is required to better understand the relationship between sub-vertical and subhorizontal vein orientations and higher grade molybdenum mineralization.
7.1
Regional Scale
The Ruby Creek deposit is a disrupted, dome-shaped occurrence formed late in the development
of a localized plutonic complex. It is associated with granitic to quartz monzonitic rocks of the
Surprise Lake Batholith, east of Atlin. Regional geology is shown in Figure 7-1.
December, 2007
- 25 -
Figure 7-1
Regional Geology Map of the Atlin Area
The geology of the Atlin area was mapped by Aitken (1959) and the regional setting of the
deposit is discussed by Christopher and Pinsent (1982). Described simply, the Atlin area is
underlain by deformed and weakly metamorphosed ophiolitic rocks of the Pennsylvanian and/or
Permian-aged Cache Creek Group (Monger, 1975). These rocks, which include serpentinites and
basalts as well as limestones, cherts and shales, are thought to be the source of much of the placer
gold found in the Atlin area. The stratigraphic rocks are cut by two younger batholiths. North of
Pine Creek, the stratigraphy is cut by a Jurassic-age granodiorite to diorite intrusion known as the
Fourth of July Batholith and, north and south of Surprise Lake, it is cut by a Cretaceous-age
granitic to quartz monzonite intrusion known as the Surprise Lake Batholith.
The rocks are locally strongly faulted and the Ruby Creek deposit is located near the intersection
of two major, pre- to post-mineral fault systems. The deposit location is partially offset by the
Adera fault system which trends from southwest to northeast down Ruby Creek and defines much
of the southern boundary of the fourth of July Batholith. The deposit is also controlled by the
Boulder Creek fault system. This runs due north up Boulder Creek and cuts across the head of
the Ruby Creek drainage. The Boulder Creek fault appears to have helped localize emplacement
of the deposit, which is intimately associated with late-stage porphyritic and aplitic plutonic rocks
intruded into a marginal phase of the Surprise Lake Batholith.
December, 2007
- 26 -
Ruby Mountain, immediately to the south of the deposit, is underlain by Late Tertiary to
Quaternary flows from a volcano that erupted and filled the lower part of the Ruby Creek
drainage with columnar basalt and volcanoclastic debris. The volcanic rocks unconformably
overlie placer gold-bearing gravels. The origin of the gold is uncertain; however, most of it
probably comes from quartz-carbonate veins hosted by shears that cut Cache Creek Group strata.
7.2
Local and Property Scale
The Ruby Creek deposit underlies the valley floor near the head of Ruby Creek. It is largely
buried and has very little surface expression. There is little outcrop in the lower part of the valley
and molybdenite is only rarely found in float and/or in veins outcropping in the bed of the creek.
The geology underlying the valley floor is largely derived from drill data.
The Ruby Creek area is underlain by two separate pulses of plutonic rock. The first pulse, which
includes the contact phase between the two batholiths, consists of a highly variably textured unit
that grades from Coarse-grained Quartz Monzonite (CGQM) south of the Adera fault through a
number of texturally transitional phases including Transitional and/or Hybrid Coarse-grained
Quartz Monzonite (CGQM-T; CGQM-H) and Crowded Quartz Feldspar Porphyry (CQFP) to
Sparse Quartz Feldspar Porphyry (SQFP) upward and outward from the deposit. The latter is well
exposed north of the Adera fault, near the diorite contact.
The CGQM is weakly to moderately deformed pink to grey equigranular, coarse grained (0.5 to
3.0 cm) quartz monzonite consisting of equal amounts of orthoclase, plagioclase and grey quartz
(Christopher and Pinsent, 1982). The feldspar is commonly seriate and, locally, includes a small
amount of fine-grained (2 to 4 mm) matrix. CGQM grades to SQFP with increase in matrix
content, and increased isolation of constituent phenocrystic crystals, particularly orthoclase and
quartz.
The first phase also includes a distinctive Mafic Quartz Monzonite Porphyry (MQMP) unit that is
present east of the deposit. This distinctive grey rock type has a seriate (1 to 4 mm locally)
porphyritic texture. It is composed largely of chalky white plagioclase, disseminated biotite and
phenocrysts of ragged plagioclase and lesser quartz. These rocks were fractured and deformed
prior to emplacement of the second pulse of magma.
There are three main mappable phases to the second pulse. They include Crowded Quartz
Monzonite Porphyry (CQMP), Sparse Quartz Monzonite Porphyry (SQMP) and Fine-grained
Quartz Monzonite (FGQM).
The CQMP has an average of 50% (2 to 6 mm) subhedral to euhedral plagioclase, orthoclase,
quartz and biotite phenocrysts in an aphanitic matrix. The SQMP variety is similar, but has fewer
December, 2007
- 27 -
(10% to 30%) phenocrysts. The SQMP is fresher and generally less deformed than the
surrounding rocks and has a much finer, more chilled matrix than the sparse quartz feldspar
porphyry described above. The second phase porphyries cut out the older rock units and are
exposed locally in the floor of the valley. They are also found in the subsurface, under the valley
floor, upstream where the CGQM and its variants are intruded by a buried cupola of SQMP. Its
shape has strongly influenced the locus of mineralization, as shown by Placer's 0.06% Mo and
0.1% Mo assay contours at 1448 m elevation. Mineralization surrounds the buried cupola and, to
a lesser extent, covers it.
The FGQM is a variably textured aplite that intrudes the CGQM (and also its variants) and the
MQMP, above and around the sparse and crowded porphyry intrusions. This rock type is not
exposed on the surface, but it is well documented as forming a series of 0.05 to 10 m thick,
approximately flat lying, structurally-controlled sills in the higher-grade (north-eastern) portion of
the deposit. FGQM dykes are found elsewhere around the buried sparse-porphyry cupola;
however, they are generally less frequent and smaller, and occur as narrow dykelets.
In addition to these rock types, recent drilling at the southwest end of the deposit has located a
Megacrystic Porphyry (MFP) unit in the subsurface. This is not well constrained; however, it
appears to be a relatively young phase of the quartz monzonite intrusion. It consists of rare to
abundant large (> 10 mm) euhedral orthoclase phenocrysts in a chilled matrix. Another notable
feature throughout the deposit is the presence of coarse-grained quartz-feldspar pegmatite. This
is not abundant, but covers a wide area as small dykes and structurally controlled sills.
Large-scale fault structures and their splays have provided conduits for mineralizing fluids and
have localized mineralization. The deposit is situated at the intersection of the Adera and
Boundary Creek faults. The Adera fault is particularly important because it offsets the northern
portion of the deposit. It is a composite structure that dips steeply to the northwest, is normal in
character and appears to have down-dropped (to the north) the north-western part of what was
originally a dome, or ring-shaped deposit formed above and around a sparse-porphyry intrusion.
Mineralization has been found in the coarse-grained and related rocks northwest of the
southernmost strand of the Adera fault. However, it has not been found in (probably similarly
aged) sparse quartz feldspar porphyry and related rocks further north. These rocks, which are well
exposed in the creek canyon below Molly Lake, contain abundant barren quartz veins and
disseminated pyrite. They are gossanous, but barren.
In addition, during the 2006 drilling program, 8 drill holes (AD-357, AD 359, AD-361, AD-363,
AD-364, AD-366 to AD-368) from the inclined drilling program were surveyed with an optical
televiewer camera to determine dominant dip and dip directions of the mineralized veins. This
work was completed by the Golder Burnaby office as a separate project from the mineral resource
estimate. At the time of the July 23, 2007 Mineral Resource Update, a review of this data
December, 2007
- 28 -
collected by Golder was reviewed by the qualified person and three dominant vein orientation
sets were identified and are as follows:
•
Set 1: 50-80° dip/170-195° dip direction;
•
Set 2: 50-85° dip/300-310° dip direction; and
•
Set 3: 5-15° dip/350-010° dip direction.
The structural data collected from the inclined holes were based on veins sizes ranging from
<5 mm, 5-10 mm and >10 mm. Set 3, near horizontal veins, had the largest number of veins
identified in the analysis, which is consistent with the current understanding of mineralization.
Sub-vertical mineralization veins (Sets 1 and 2) were identified during the structural analysis, but
previously of unknown orientation.
It is likely that Set 2 could be a subset of the Adera Fault since it is estimated as dipping steeply
to the northwest.
The rocks at Ruby Creek are, for the most part, fresh and much of the alteration that is observed
is post-mineralization, associated with fluids that circulated during post-mineral faulting.
However, there is a small amount of primary alteration. It occurs as sill-like zones of intense
silicification intermixed with bodies of aplite in the higher-grade, north-eastern part of the deposit
and as intermittent feldspar envelopes and flooding around individual mineralized quartz veins
throughout the deposit. In one locality, the silicification can be shown to pre-date both intrusion
of aplite and emplacement of mineralized quartz veins.
Fractured rocks near post-mineral faults, such as the Adera, have commonly undergone late
hydrothermal alteration. They are either weakly or strongly altered to a mixture of sericite,
carbonate, clay and chlorite (without addition of secondary quartz). The altered rocks become soft
and friable and early (1969/72) core recoveries were lower in these areas. Major faults commonly
include breccias cemented by grey gouge of similar composition, with or without smeared
molybdenite. Some of the altered rocks contain fluorite veins. Work by Placer in 1980 shows that
most of the light-coloured clay is predominantly montmorillonite; however, the grey clay in the
main Adera fault zone consists largely of kaolinite.
December, 2007
8.0
- 29 -
DEPOSIT TYPES
Details of the deposit type of the Ruby Creek property are contained in Golder’s Mineral
Resource Estimate Reports, (2006 and 2007) which are based on the reports by Sinclair (1995)
and Blower (2005). Excerpts from those reports are provided below.
Ruby Creek can be classified as a low fluorine porphyry molybdenum deposit (Sinclair, 1995).
These deposits are characterized by stockworks of molybdenitebearing quartz veinlets and
fractures in intermediate to felsic intrusive rocks and associated country rocks. They are typically
low-grade but large and amenable to bulk mining methods.
Porphyry molybdenum deposits vary in shape from an inverted cup, to roughly cylindrical, to
highly irregular. They are typically hundreds of metres across and range from tens to hundreds of
metres in vertical extent. Mineralization is predominantly structurally controlled, consisting
mainly of stockworks or crosscutting fractures and quartz veinlets, with veins, vein sets and
breccias. Molybdenite is the principal ore mineral; chalcopyrite, scheelite, and galena may be
present but are generally subordinate (Sinclair, 1995).
These deposits are thought to originate from large volumes of magmatic, highly saline aqueous
fluids under pressure. Multiple stages of brecciation related to explosive fluid pressure release
from the upper parts of small intrusions result in deposition of ore and gangue minerals in
crosscutting fractures, veinlets and breccias in the outer carapace of the intrusions and in
associated country rocks. Incursion of meteoric water during waning stages of the
magmatic-hydrothermal system may result in late alteration of the host rocks, but does not play a
significant role in the ore-forming process (Sinclair, 1995).
December, 2007
9.0
- 30 -
MINERALIZATION
A detailed description of the mineralization is contained in Golder’s Mineral Resource Estimate
Reports (2006 and 2007), which is a summary based on reports by Pinsent (2005) and Blower
(2005). Excerpts from those reports are provided below.
The Ruby Creek deposit consists of a stockwork of veins of molybdenite and quartz molybdenite
found in all the principal rock types. However, it is best developed in the early stage plutonic
rocks (mafic quartz monzonite porphyry and coarse grained quartz monzonite and its variants)
that overlie and surround the buried sparse quartz monzonite porphyry stock under the Ruby
Creek valley. The veins are most commonly without other metallic phases, although pyrite is
found locally and chalcopyrite has been observed. The veins locally contain traces of scheelite,
orthoclase, fluorite, biotite, sericite, and carbonate.
Mineralization post-dates emplacement of fine grained quartz monzonite in the higher-grade
zone, located on the northeast side of the deposit. In this area, there appears to be a crude positive
correlation between the presence of dykes and sills and the amount of mineralization observed.
However, the same relationship does not hold on top of the cupola or in the south-western part of
the deposit. The deposit consists of a mineralized blanket that covers the sparse quartz monzonite
porphyry stock and dips off in all directions.
The mineralization commonly consists of sulphide veins as coatings on quartz free fractures, and
as coarse and fine rosettes and blebs in both smoky and lesser clear quartz. It also occurs as
streaks and smears in deformed rock and may, locally, be enriched in fault zones. In the higher
grade zone, explored by Kerr Addison, much of the mineralization is in horizontal to subhorizontal veins and fractures from 1 mm to 5 mm wide that are interspersed with veins that are
considerably wider, up to 20 mm wide. The narrow quartz veins are oriented at a high angle to the
(vertical) core axis. Both vein sets are mineralized and blebs of molybdenite commonly occur at
the intersection of cross cutting veinlets.
The near horizontal vein set is locally extremely well mineralized. Veins exposed through
underground development in the 1970's show that coarse rosettes of molybdenite up to 30 mm in
diameter formed in the plane of the vein, and that the spacing between the rosettes is variable,
causing a pronounced nugget effect in drilling.
The crowded and sparse porphyries underlying the higher grade zone are cut by narrow (1 mm to
3 mm) mineralized quartz veins and fractures that also occur at both high and low angles to the
(vertical) core axis. These veins and fractures commonly contain fine grained to powdery
molybdenite. There are fewer high grade rosettes formed at depth.
December, 2007
- 31 -
As outlined previously, optical televiewer surveys were completed in eight inclined boreholes
form the 2006 drilling program in order to determine the dominant orientation of veins in the
central portion of the deposit. The following three dominant sets have been identified from the
survey:
•
Set 1: 50-80° dip/170-195° dip direction;
•
Set 2: 50-85° dip/300-310° dip direction; and
•
Set 3: 5-15° dip/350-010° dip direction.
The dominant vein set identified from the survey is the horizontal set followed by two
sub-vertical sets, Set 1 dipping south and Set 2 dipping northwest. It was identified by the
Adanac geology staff that zones that were dominated by both horizontal and vertical veins sets in
the 2006 drilling program showed an increased molybdenite mineralization. Therefore, the
drilling of inclined boreholes in areas where both horizontal and vertical veining occurs increases
the chance of sampling these zones. Based on the orientation of sub-vertical sets any future
planned inclined boreholes should be drilled perpendicular to dip direction of these sets (azimuth
directions north and southeast).
December, 2007
10.0
- 32 -
EXPLORATION
Much of the past efforts (post 1966) at Ruby Creek have focused on the central portion of the
deposit. The Placer and Kerr Addison exploration campaigns both focused on the central portion
of the deposit to depths approximately 200 m below ground level. Mineralization occurs below
this depth, but is not necessarily amenable to open pit mining. Descriptions of the past
exploration programs prior to Adanac exploration programs has been described in Section 6
(History). The following sections focus on the exploration programs completed by Adanac from
2004 to 2006
10.1
ADANAC 2004
Adanac conducted a major exploration drilling campaign in 2004. The company drilled 38 holes
having an aggregate length of 9,087 m in and around a previously established deposit and
submitted 2,830 samples for molybdenum assaying. Additionally, 256 samples were collected for
specific gravity testing by ALS Chemex Laboratories (ALS Chemex) in Vancouver, BC.
Ten holes were twinned in 2004 (five twinned with Kerr Addison and five with Placer). These
holes were designed to confirm drill holes from earlier campaigns to validate original drilling
results. The remaining holes were located to improve Adanac’s understanding of the overall
shape of the deposit. The results were reviewed, validated and Mineral Resources in compliance
with NI 43-101 standards were first reported in AMEC's Technical Report (Blower, 2005).
10.2
ADANAC 2005
Adanac expanded on their 2004 campaign during 2005. Adanac drilled 19 holes having an
aggregate length of 4,984 m in and around the previously established deposit as well as on the
deposit fringes. Five of the holes were geotechnical holes designed to generate information
necessary for pit slope stability assessment. The remaining twelve holes were located to infill past
campaigns on the fringes to improve Adanac’s understanding of the overall shape, depth, and
extremities of the deposit.
Drill core sampling for the 2005 drilling program was based on the 2004 QA/QC program
developed by Adanac. A total of 1,559 drill core samples were submitted to ACME Analytical
Laboratories (ACME) in Vancouver for analysis.
The seven holes internal to the established deposit, AD-337 to AD-343, were submitted for
molybdenum analysis. The three short distal exploration holes, AD-344 to AD-346 were
submitted for molybdenum as a trace element. The remaining drill holes (including geotechnical
holes), AD-347 to AD-355 were also analysed for molybdenum as a trace element. Additionally,
December, 2007
- 33 -
60 samples from seven drill holes (AD-337 to AD-343) were submitted for assay checks to
ALS Chemex as part of the quality control program.
A total of 615 samples were also collected for specific gravity testing and were submitted to
ALS Chemex. These samples included 332 from 19 drill holes in the 2005 program and 283 from
30 holes in the 2004 drilling program. The results from the specific gravity testing showed similar
results to the testing results in the Technical Report – Mineral Resource Estimate Ruby Creek
Project (Blower, 2005).
All 2005 drill hole collars were surveyed at the end of the drilling program by Underhill
Geomatics Ltd and were provided by Adanac using the NAD 27 UTM co-ordinate system which
was also used in the 2004 drilling program. The NAD 27 UTM same co-ordinate system was
also used in the historical drilling programs and was used in the 2004 and 2005 drilling programs
for consistency. All 5 geotechnical drill holes were inclined dipping and were surveyed by
Golder using an optical televiewer system.
The 2005 drilling program, sampling program and surveying was reviewed and validated and, in
Golder’s opinion, was sufficient to include in the 2006 Mineral Resource Estimate. Golder also
visit the project site during the 2005 drilling program.
10.3
ADANAC 2006
Adanac continued a drill program during 2006. Adanac drilled 16 holes for an aggregate depth of
3,921 m in and around the previously established deposit (central deposit area) as well as the
south and southwest edges of the deposit as illustrated on Figure 10-1. Three of the holes were
located on the fringes to improve Adanac’s understanding of the overall shape, depth and
extremities of the deposit. The remaining 13 holes were located to infill past campaigns and
drilled at approximately -50° dip and 270° Azimuth.
The location of the 2006 drill holes is illustrated on Figure 10-1.
December, 2007
- 34 -
Figure 10-1
December, 2007
- 35 -
Drill core sampling for the 2006 drilling program was based on the 2004 and 2005 QA/QC
program developed by Adanac. A total of 1,238 drill core samples were submitted to ACME in
Vancouver, BC for analysis, including 295 samples from drill holes AD-356 to AD-368, which
were submitted for molybdenum oxide analysis. For quality control purposes, samples from drill
holes AD-369 to AD-371 were also analyzed for molybdenum as a trace element and analysed for
40 other elements. Additionally, 186 samples from 13 drill holes (AD-356 to AD-368) were
submitted to G&T Metallurgical Services for metallurgical testing. A total of 176 samples were
submitted to ALS Chemex for specific gravity testing from the 2006 drilling program. One
specific gravity sample was selected approximately every 50 linear ft.
All 2006 drill hole collars were surveyed at the end of the drilling program by Underhill
Geomatics Ltd. and provided by Adanac using the NAD 83 UTM co-ordinate system as well as in
the NAD 27 UTM co-ordinate system. The NAD 27 UTM co-ordinate system was also used in
the historical drilling programs and in the 2004 and 2005 drilling programs for consistency. The
survey consulting company also provided 10 historical drill hole locations re-surveyed in the
NAD 83 UTM co-ordinate system.
A decision by Adanac to use the NAD 83 UTM co-ordinate system for future mine site
construction resulted in the conversion of the entire database to the NAD 83 UTM co-ordinate
system. This conversion was based on a combination of drill holes surveyed in NAD 83 UTM
and translating the remaining drill hole locations from NAD 27 UTM to NAD 83 UTM by adding
174N (Y), and subtracting 104E (X).
The 13 infill drill holes were inclined and 8 of these were surveyed by Golder using an optical
televiewer camera system. Where possible, the drill holes’ steel collar casings were left behind in
the drill holes.
Excel spreadsheets of the geological and assaying data for the 2006 drilling program were
reviewed by Golder and included in the Ruby Creek Datamine Database. The QA/QC program
was reviewed during the site visit by the qualified person and was consistent with the previous
site visit and, in Golder’s opinion, was sufficient to include in the July 23, 2007 Mineral Resource
Update.
10.4
2007 Drilling Program
At the time of this report the 2007 drilling program was nearing completion. A site visit was
completed by a Golder qualified person on September 25, 2007. At the time of the site visit, 5
drill holes had been completed out of a planned total of 14 drill holes (6,839 m). The details
(location of drill holes, assay grades, etc.) of this 14 drill holes have not been included in this
report nor have any results from this drilling been included in the July 23, 2007 Mineral Resource
Update.
December, 2007
11.0
- 36 -
DRILLING
Summarized in Table 11.1 is the history of the drilling programs from 1966 to 2006. The holes
were typically diamond drill holes except for a small number of early rotary drill holes. Not all of
these drill holes are contained in the Ruby Creek Datamine Database used for the July 23, 2007
Mineral Resource Update.
Table 11.1
Drilling History Of The Ruby Creek Deposit (Dec 31, 2006)
Years
Drill
holes
(m)
Adanac Mining and Exploration, &
John’s Manville
1966 to 1970
80
12,775
Kerr Addison Mines
1970 to 1972
47
5,626
Climax Molybdenum
1973
9
2,672
Placer Development
1979 to 1980
76
10,886
Adanac Gold
(Adanac Moly Corp.)
2004
381
9,0871
Adanac Moly Corp.
2005
192
4,982
Adanac Moly Corp.
2006
16
3,921
-
285
49,950
Company
Total
Notes:
1
2
Includes 2 re-drills of holes
Includes 5 geotechnical holes
The majority of the drilling completed to date on the Ruby Creek deposit is vertical dipping and
the main mineralization to date has been identified as along sub-horizontal veins except in the
central area of the deposit which is a mixture of sub-vertical and sub-horizontal veins. Therefore,
in general the mineralized drill intercepts are representative of the true thickness with some
exceptions in the areas where mineralization is both sub-vertical and sub-horizontal vein hosted.
The exception to this is the inclined drill holes completed in 2005 (geotechnical holes), the
inclined vertical drill holes in 2006 to assist in defining vein orientation and the pseudo-drill holes
(horizontal) representing the underground sampling that was completed by Kerr Addison.
The drill holes to date have been spaced predominantly along an exploration grid with East-West
and North-South grid lines equal to 064° and 154° azimuths using the UTM NAD 27 co-ordinate
systems. The section spacing between the grid lines is 100 ft with East 1 representing 100 ft east
of baseline West 00. The spacing of drill holes varies, but is typically spaced 300-400 ft
(approximately 90-120 m) apart between Sections East 18 to West 28. A higher grade
mineralization zone has been identified around the underground development area, central
deposit, which has been drilled with a density of approximately 100 ft (approximately 30 m)
spacing between Sections East 6 to West 10.
December, 2007
- 37 -
The historical drilling data was originally transferred from paper logs to electronic spreadsheets
by Adanac and then entered into a database that was used for the April 2005 Mineral Resource
Estimate (Blower, 2005). This database was provided to Golder as ASCII files and was
incorporated in the 2006 Ruby Creek Datamine Database. Drill hole information provided by
Adanac from the 2004 to 2006 drilling programs, as Microsoft Excel spreadsheets, were
incorporated in the Ruby Creek Datamine Database.
The 2006 drilling program added an additional 16 drill holes to this Ruby Creek Deposit as
illustrated on Figure 11-1. Figure 11-1 is a plan view of the deposit showing the location of the
266 drill holes from the 2006 Ruby Creek Datamine Database and includes the historical
exploration grid and the UTM NAD 83 Grid. In 2006 all drill hole information was converted
from the UTM NAD 27 co-ordinate system to UTM NAD 83.
The 2006 Ruby Creek Datamine Database naming convention for drill holes is based on the
various drilling campaigns and included four main groups. The original drill hole names have
been re-labelled originally during the 2005 Mineral Resource Estimate (Blower, 2005) with the
same naming convention continued in the 2006 Ruby Creek Datamine Database. All drill holes
with a prefix of KA for Kerr-Addison, CM for Climax Moly, PD for Placer and AD for Adanac
(2004, 2005 and 2006 drilling). Table 11.2 summarizes the drill holes compiled in the 2006 Ruby
Creek Datamine Database used in the July 23, 2007 Mineral Resource Update.
Table 11.2
Drill Holes in the 2006 Ruby Creek Datamine Database
Campaign
Years
Drill Holes
(m)
1966 – 1972
105
16,897
1973
7
1,148
1979 – 1980
66
9,975
Adanac Gold Corporation
(Adanac Moly Corp.)
2004
36
8,984
Adanac Moly Corporation
2005
19
4,982
2006
16
3,921
Kerr Addison Mine (underground
sampling as pseudo (includes raises)
drill holes
1972
17
1,005
266
46,912
Kerr Addison Mines (including
Adanac Mining and Exploration, &
John’s Manville)
Climax Molybdenum
Placer Development Limited
Total
December, 2007
- 38 -
The database contains 17 pseudo-drill holes (includes raises) created from the underground
sampling program, by Kerr Addison, for a total of 1,005 m. The underground sampling program
consisted of representative samples collected from drift rounds in the adits excavated by Kerr
Addison. Access to the underground workings is currently not possible. Therefore, a total of 266
drill holes are currently contained within the 2006 Ruby Creek Datamine Database that has been
used in the July 23, 2007 Mineral Resource Update.
Site visits by Golder qualified person’s have been completed in 2005, 2006 and 2007, during the
active exploration field season at the Ruby Creek Project site. The same drill core logging
procedures have been observed during each field season.
All core logging and sampling by the Adanac geologists is first entered on paper logs and later
entered electronically onto computers for permanent storage as Excel spreadsheets. After the
core was sampled from the core trays, any remaining core (typically half) was stored on woodenrebar rack structures in their original open core trays. The core rack structures are stored outside,
but are protected under wooden roofs. Each core rack structure is labelled with the drill hole
name. There is no core stored on site or available prior to the 2004 drilling program. In my
opinion, the drilling practice, logging, handling and storage of core employed at the Ruby Creek
deposit is standard in the industry and suitable for Mineral Resource Estimation.
Ruby Creek Feasibility Study
December, 2007
- 39 -
Figure 11-1
December, 2007
12.0
- 40 -
SAMPLE METHOD AND APPROACH
In the drill programs completed prior to 2004, the standard practice was to crush all core, saving
only a small, lithologically representative sample from each 10 ft interval. This approach was
taken in order to minimize handling, reduce molybdenite loss through splitting or sawing, and
increase the volume of material sampled (hence improving sampling statistics).
No assay samples or representative drill core samples are available prior to the 2004 drilling
programs completed by Adanac. All samples that were collected from the 2004 to the present
from drilling programs by Adanac have been sawed or split in half and a portion retained to
establish an inventory of archived core samples. A few holes were sampled entirely in order to
compare the data to historical assay results. These archived split samples are stored on the
permanent core racks on the property site.
During the core logging process, samples were selected for assaying. All drill holes were
sampled beginning from the bedrock interface. No overburden samples were collected for
assaying in the 2004, 2005 and 2006 drill programs. Sample selection was a combination of
lithology type and length. The typical sample length was 3.05 m (10 ft) which was the same
length as the drill core run and similar to historical sample lengths previously collected. The
maximum (4.8 m) and minimum (1.2 m) samples lengths collected in 2004, 2005 and 2006
programs were typically the first or last samples collected in each drill hole.
In practice, the drill core was processed in three ways. The competent sections of the drill holes
were either split using a classic hand-cranked core splitter, or sawed using a fast and efficient
(Almonte) core saw. Core intervals that were structurally weak and/or too poorly consolidated to
split were totally crushed and then passed through a riffle splitter three times to homogenize the
sample before being split into to halves. One half was then treated in the same way as the other
half-core crushed samples obtained by cutting and/or sawing, and the other was double-bagged
and stored as a primary crushed reject.
As part of Adanac’s QA/QC program approximately 5% of the material left in the boxes (split
samples) are analyzed as duplicate samples. In addition to the core duplicate samples, several
sets of representative samples (0.5 m to 0.1 m long) were collected from the archived split
samples for specific gravity, acid generating potential determination and metallurgical testing. A
total of approximately 1,000 samples have been collected from the 2004, 2005 and 2006 drilling
programs for specific gravity testing (approximately 50 ft spacing). Therefore, the residual core
left on site is incomplete. Most of the specific gravity and representative samples taken from the
2006 drill core are in the company office in White Rock, British Columbia.
December, 2007
- 41 -
During the 2004 program, three drill holes were totally crushed to reproduce the sample handling
processes that were taken in pre-2004 drilling programs. These holes (AD-303, AD-307, and
AD-308) were collared adjacent to pre-existing holes drilled by Kerr Addison (KA-60-1 17) and
Placer (PD-221, PD-227).
Although the process of sawing or splitting the core can cause some loss of molybdenite, it is
worth noting that the drill hole recoveries returned in the 2004 through 2006 programs were often
higher (>95%) when compared to the losses through down-hole erosion that were experienced by
either Kerr Addison or Placer.
Golder reported that, the sampling procedures used in the 2006 drilling program are consistent
with the 2004 and 2005 practices developed and are consistent with industry practices and were
representative of the drill hole data collected.
December, 2007
- 42 -
13.0
SAMPLE PREPARATION, ANALYSES AND SECURITY
13.1
Field Sample Preparation Procedures
13.1.1 Adanac 2004, 2005 and 2006
As discussed in the previous section, all samples selected for assaying were crushed on site to less
than approximately 10 mm (3/8") using a Nelson Machinery Atlas core crusher that was cleaned
with compressed air before the next sample was crushed. Samples that were crushed were either
sawed half samples or whole samples (if unable to be sawed). The crushed samples were then
weighed (typically 8.0 kg to 10.0 kg for a sawed sample) and subjected to a systematic splitting
process using an industry standard riffle splitter. The original sample (typically 8 kg) was split
into two 4 kg samples and they were, in turn, split into four 2 kg samples. Two of these (one
from each of the 4 kg splits) were assigned to a reject bag and the remaining two were split to
produce four 1 kg samples. Two of these (again, one from each of the original 4 kg splits) were
then mixed to form the main assay sample. Where appropriate, the remaining two 1 kg samples
were also mixed to make a primary crush duplicate.
The same splitting process was used for samples that were unable to be sawed, except a larger
sample was retained since these samples were not stored in the core racks but with the primary
crush duplicates. Those samples (main and duplicate) selected for analysis were then weighed,
assigned their assay tags and sealed using a single-use cinch-tie. The samples were shipped to
ACME in rice-sacks. Golder observed the 2005 and 2006 field sample preparation procedures
and found them to be consistent with industry standards.
Illustrated on Figure 13-1 is a flow diagram of the field sample preparation procedures.
December, 2007
- 43 -
Figure 13-1
Adanac Field Sample Preparation Procedures
Half Core
8.0 kg
4.0 kg
2.0 kg
1.0
4.0 kg
2.0 kg
2.0 kg
1.0
Main Sample,
for Acme Labs.
2.0 kg
1.0
2.0 kg
2.0 kg
Duplicate Sample,
if requested.
1.0
Reject
6.0 kg
for
or 4.0 kg storage
*From Pinsent (2005)
13.1.2 Prior to Adanac
The sampling procedures used prior to Adanac’s 2004 to 2006 drilling programs are found in the
Technical Report - Mineral Resource Estimate Ruby Creek Project (Blower, 2005).
•
Prior to 1970, samples were prepared for assay in various laboratories. There was no
preparation in the field.
•
Pre-1979: Kerr Addison (1970), after constructing a bucking room and assay lab on site,
completed both the field and laboratory sample preparation at the property.
•
1979 - 1980 Placer: Placer's field sample preparation procedures are virtually identical to
Adanac’s 2004 protocols, except that Placer employed a second stage of crushing with a
gyratory crusher to further reduce the 1/4" jaw crusher output to a -8 mesh sample
weighing 1.75 lbs.
December, 2007
13.2
- 44 -
LABORATORY SAMPLE PREPARATION PROCEDURES
13.2.1 2004, 2005 and 2006 ADANAC
Samples from Adanac’s drilling programs from 2004 onwards were submitted to ACME
Analytical Laboratories for molybdenum and trace element analysis. Check sampling and specific
gravity estimates were submitted to ALS Chemex.
Samples collected during the drilling campaigns were prepared for molybdenum and trace
element assaying. They were crushed to 70% passing 10 mesh and splits (250 g) were then
pulverized to 95% passing -150 mesh.
Methods used for sample preparation prior to the 2004 and 2005 drilling program are provided in
the Technical Report - Mineral Resource Estimate Ruby Creek Project (Blower, 2005).
13.2.2 1979 — 1980 Placer
Placer's 800 gram crushed samples were shipped to their laboratory, where the entire sample was
pulverized, but the targeted pulp specifications are not known.
13.2.3 Pre-1979 Kerr Addison
Kerr Addison's laboratory sample preparation procedures in 1970 are shown in Figure 13-1. The
sample preparation procedures prior to 1970 were not available.
December, 2007
- 45 -
Figure 13-2
Kerr Addison's Laboratory Sample Preparation Procedure
13.3
Analytical Procedures
13.3.1 Adanac 2004, 2005 and 2006
The molybdenum assaying method that was applied to all the 2004 to 2006 samples (including
blanks, duplicates and standard samples) by ACME was the multi-element method. This method
takes a 1.0 gram split pulverized sample first digested by aqua regia and then analyzes the
resultant solution for molybdenum (Mo) using Inductively Coupled Plasma Emission Mass
Spectrometry (ICP-MS).
Trace element assaying of the 2005 and 2006 samples by ACME was prepared the same as above
and then a 0.25 gram sample was heated in HNO3-HCLO4-HF to fuming and taken to dryness.
The residue sample was dissolved in HCL and the resultant solution was then analyzed for
41 elements (parts per million) using ICP-MS.
Note that the values returned are the total molybdenum content of the rock, as they combine the
sulphide with any oxide molybdenum that may be present. However, in practice, previous work
showed that there is little or no molybdenum present in oxide form. In Golder’s opinion, the
December, 2007
- 46 -
ICP method is the best practice for measuring the molybdenum content of molybdenite
mineralization in rock.
13.3.2 1979 — 1980 Placer
Placer completed all of their analytical work at their in-house laboratory in Vancouver. The
samples were analysed by Atomic Absorption Spectrometry (AAS) and were reported as % Mo in
MoS2 (the sulphide portion of the total Mo content after removal of the oxide portion)
(Christopher and Pinsent, 1982).
13.3.3 Pre-1979 Kerr Addison
Several laboratories were used for the pre-1979 drilling (Dagbert and David, 1975). Details on
the 1970 analytical techniques are not available. Assaying in 1970 was completed by Kerr
Addison on site in their own laboratory using a colorimetric - spectrophotometric procedure for
most samples, and a gravimetric method for higher grade or mill product determinations
(Chapman, Wood & Griswold, 1971). The on-site lab results were apparently checked with
duplicate assays at Loring Laboratories.
No samples were available prior to the 2004 drilling program; therefore, no check assays have
been completed on sample information prior to 2004. Adanac drilled five twin holes in 2004 to
test the previous assaying results. The results from the twin hole drilling program are discussed
in Golder’s 2006 Mineral Resource Estimate Report.
13.4
Quality Assurance and Quality Control
13.4.1 Adanac Procedures
Adanac has employed a comprehensive program of QA/QC consisting of inserted blanks (5%),
standards (5-10%) and duplicate samples (5%) from the 2004 to the 2006 program and has
employed the same program for the 2007 program. The QA/QC program employed follows the
same QA/QC program as the one developed in 2004 which is described in the Mineral Resource
Estimate Reports by Golder (2006 and 2007) and AMEC (Blower, 2005). There is no
information available on QA/QC procedures or results from the drilling completed prior to 2004.
Samples submitted to ACME included split drill core samples, blanks (two types) and standards
(two types). Golder has reviewed the assaying results from the duplicates, blanks and standards
that were provided by Adanac in 2005 and 2006. A description of 2006 procedures for the
blanks, standards and duplicates is as follows and is provided in the July 23, 2007 Mineral
Resource Update Report Update. Information regarding a review of the 2005 QA/QC program is
December, 2007
- 47 -
provided in the 2006 Technical Report – February 8, 2006 Mineral Resource Estimate Ruby
Creek Molybdenum Project
2006 Blanks
Blank samples were inserted in the sample stream to monitor contamination. Adanac used two
types of blanks: (1) bags of pre-crushed, commercially purchased, poultry grit quartzite (Blank),
and (2) locally derived volcanic scoria from Ruby Mountain, west of Ruby Creek (Blank-S). The
poultry grit blanks did not go through the on-site crusher and were inserted as a standard.
However, those designated as Blank-S, composed of locally derived volcanic scoria from Ruby
Mountain, did go through the crusher and they provide a check on sample to sample
contamination during the crushing process. Approximately 5% of the samples submitted for a
drill hole included blank samples with no molybdenum mineralization.
2006 Standard Reference Material
During the 2006 drilling program, two commercially purchased molybdenum standards were
included as standard reference materials during the submission of drill core samples to ACME for
both standard ICP-MS molybdenum assaying and trace element assaying. These standards
included WCM Cu 111 and WCM Cu 132. The standard assaying results for WCM Cu 111 is
0.83% Cu, 0.117% Mo and 105 g/t Ag. The standard assaying results for WCM 132 is
0.17% Cu, 0.045% Mo and 27 g/t Ag and 0.17 g/t Au.
2006 Reject and Pulp Duplicates
During analysis, ACME also inserted reject duplicates from a second 250 g split from the original
sample, and pulp duplicates from a second split from the original pulp, as requested by Adanac.
13.4.2 QA/QC Results –2006 Samples
The analytical methods employed on the 2006 samples from the Ruby Creek deposit by ACME
were the same as those used during the 2004 and 2005 sampling programs. The ICP-MS
analytical method used on the 2006 samples considered the best practice for measuring the
molybdenum content in the drill core samples.
The assay results from the blanks, standards and duplicates from the 2006 QA/QC program were
reviewed by Golder with the following results:
•
18 Blank samples were reviewed and had an assay range between 0.0005 and
0.004% Mo. Only one sample had an assay of 0.004% Mo with the remaining at or below
0.001% Mo.
December, 2007
- 48 -
•
27 Blank-S samples were reviewed and had an assay range between 0 (below detection)
and 0.006% Mo.
•
39 WCM Cu 111 standard samples were reviewed and had a range of 0.103% Mo and
0.123% Mo with a mean and standard deviation of 0.111% Mo and 0.0049% Mo,
respectively. Only 1 of the 39 samples exceeded 2 standard deviations.
•
39 WCM Cu 132 standard samples were reviewed and had a range of 0.0398 and
0.048% Mo with a mean and standard deviation of 0.0432% Mo and 0.002% Mo,
respectively. Only 1 of the 39 samples exceeded 2 standard deviations.
•
172 duplicate samples were reviewed using Q-Q plots of assay 1 versus assay 2. In
general, these plots showed reasonable agreement with some outliers occurring at all
grades (both low and high).
Therefore, in general, there appears to be no contamination in the samples crushed on site since
the Blank-S samples had a maximum assay of 0.006% Mo. Also, the standards and second
blanks submitted show no obvious contamination and reasonable accuracy and precision with
only 2 of the 78 standards reviewed exceeding 2 standard deviations above their recorded assay
grade. The duplicate samples reviewed also showed reasonable agreement, which indicates
reasonable precision and no contamination between samples. The review of the 2006 sampling
program indicates that the samples were sufficiently accurate, free from contamination, precise,
under control and were included in the 2006 Ruby Creek Datamine Database for the calculation
of the July 23, 2007 Mineral Resource Update.
December, 2007
14.0
- 49 -
DATA VERIFICATION
The data verification checks that were completed on the drill hole data prior to them being used in
the July 23, 2007 Mineral Resource Update is provided as follows.
1. A check of the drill hole data against original spreadsheet records in the database.
2. A review of the 2006 blanks, duplicate (Q-Q plots) and standards.
3. Site visit completed on August 22, 2006 to review core logging and sampling procedures.
No independent samples were collected during the site visit, but visible molybdenum
mineralization was observed and independent samples were collected in 2004 (Blower,
2005) and 2005 (Palmer, 2006).
Verification checks were completed on the 2005 and earlier data and are provided in the
Technical Reports by Palmer (2006) and Blower (2005). The data verification checks completed
on the 2006 data is discussed above with the exception of the drill hole co-ordinate translation
from UTM NAD-27 to UTM NAD 83 and as described in the following sections.
Approximately 5% of the 2006 drill hole Excel spreadsheets were visually reviewed against the
Datamine Database. The drill hole samples in the database prior to 2004 were provided as ASCII
files. These ASCII files were provided from AMEC based on electronic spreadsheets provided
by Adanac. These drill hole samples were added to the Ruby Creek Datamine Database created
in 2006 and were visually checked against the 3D geological and mineralization models created.
No significant discrepancies were encountered during the check. The collar locations for drilling
data from 2005 and earlier have undergone a conversion and translation from UTM NAD-27 to
UTM NAD 83. Some drill hole samples (pre-2004 historical data) did not have information
pertaining to percent recovery and main lithology identification, but were still included in the
Mineral Resource Estimate since assay data was available. All drill holes with missing
information (i.e. no Mo assay values for overburden samples) were flagged with a negative value
(typically -2) and were not included in the mineral estimate.
A review of the 2006 sampling data including blanks, duplicates and standards was completed in
previous sections.
Previous data verification checks completed in 2004 and 2005 and on earlier data are provided in
the 2005 (AMEC) and 2006 and 2007 (Golder) Mineral Resource Estimation Reports.
Golder concludes that the assay and survey data used in the July 23, 2007 Mineral Resource
Update were sufficiently free of error to be adequately used for resource estimation of the Ruby
Creek Molybdenum deposit.
December, 2007
15.0
- 50 -
ADJACENT PROPERTIES
Adjacent properties are not relevant to the review of the Ruby Creek property.
December, 2007
16.0
- 51 -
MINERAL PROCESSING AND METALLURGICAL TESTING
A detailed description of the mineral processing and metallurgical testing is contained in the
Process and Infrastructure Design and Cost Estimate report by Wardrop in 2006. A simplified
flowsheet is shown in Figure 16-1.
Figure 16-1
Simplified Flowsheet
The process design is based on an average of 23,000 tpd mill operation.
throughput would be 7.6 million tonnes per year.
16.1
Average annual
Mineral Processing
The milling process is based on test work done by SGS-MinnovEX and adjusted by test work by
G&T to account for a coarser primary grind. Three stages of crushing, ball milling, froth
flotation, dewatering and drying are used to produce and package a high grade molybdenite
concentrate (MoS2). Run.of.mine (ROM) ore is dumped into the gyratory crusher and crushed to
December, 2007
- 52 -
approximately -200mm. From there it is conveyed to the 130,000 tonne capacity coarse ore stock
pile; of which approximately 40,000 tonnes is live capacity. From there it is reclaimed and
conveyed to screens where the - 50 mm fraction – the finished product- goes to the HPGR units.
Screen oversize is crushed in one of two cone crushers with cone crushed product recycled to the
screens. (Secondary crushing is a closed circuit operation). The -50 mm product is fed to the high
pressure grinding rolls in open circuit. A small amount of “edge recycle” of HPGR product is
allowed for, but otherwise all product is fed directly into the ball mill. Ball mill product is
classified through cyclones. The finished product has a P80 = 275 microns. Oversize material is
recycled to the ball mill. The flotation comprises rougher/scavenger tank cells where rougher
concentrate (2%) is reground in a tower mill to a much finer state (P80 = 20 microns). This
product is cleaned in a series of stages; trace impurities are reduced to a minimum and the final
molybdenum concentrate is a high grade product.
Ruby Creek mineralization contains minimal trace impurities which are pyrite, chalcopyrite,
sphalerite, galena and bismuthinite. The extensive mineralogy work shows almost all are
liberated in the grinding and regrinding stages and can be mostly removed in the cleaning stages
with the uses of depressant D910 (sodium thiophosphate) in reasonable amounts. Other reagents
for flotation are diesel (kerosene) and pine oil which are used at natural pH (7.8 – 8.2) in the
milled pulp.
Final concentrates are thickened and settled with aid of flocculant and dewatered in a pressure
belt press. Filter cake is dried in a rotary hearth drier, cooled in a storage bin and afterwards
packaged in tote bags (~ 2 tonne capacity), weighed and moved to a storage area.
16.2
Metallurgical Testing
Kerr Addison contracted Britton Research to conduct bench scale and pilot plant test work
(August 1969 to January 1971). The pilot plant work was done on site through a 100 ton/day
plant which was fed from almost 10,000 tons of ore produced from extensive underground
development which included muck from raises on specific drill holes. Almost 7000 tons were
processed as several discrete lots from these raises and some lateral development. Results were
consistent and well documented, and a basic flowsheet was established.
SGS-MinnovEX used their proprietary technologies Comminution Economic Evaluation Tool
(CEET) and Flotation Economic Evaluation Tool (FLEET) to design the comminution and
flotation process for Ruby Creek. Initially, 30 core samples were collected from the 2004 drilling
and sent to SGS-MinnovEX in May of 2005. For the Feasibility Study, an additional 70 samples
from the 2005 drilling were sent to SGS-MinnovEX. Adanac’s geologist, Dr. Robert Pinsent,
ensured that these samples were widely spaced across the orebody and the lithology types and
that they would represent fully the variability occurring within the deposit.
December, 2007
- 53 -
The test work was completed and the reports were issued to Adanac and Wardrop in February
2006.
In December 2005, at Adanac’s request, B.C. Mining Research Limited completed tests on
alternative technologies for the regrinding options in the flotation cleaning stages and settling
tests on molybdenite concentrate, rougher tailings, and reground concentrate.
In December 2006, at Adanac’s request, G&T completed pilot scale and locked cycle flotation
studies and modal assessments of the cleaner circuit simulations.
Further details of both programs are provided in the following sections.
16.2.1 Kerr Addison’s Pilot Plant (Britton 1969 – 1971)
•
6251 tons of underground “ore” (average assay 0.12% Mo) were milled in a pilot plant.
Range of feed grade was 0.056% Mo – 0.151% Mo, and sulphide molybdenum
recoveries were in the range 94.6% - 97.1%.
•
Good grade concentrates were produced (average was >57% Mo) and lead was the only
impurity which exceeded the acceptable limit (0.06% or slightly higher)
16.2.2 SGS-MinnovEX Test Work
•
Grinding tests showed the Ruby Creek samples to be uniform with only small variations
in hardness. The ore is described as medium to soft with average work index of
12.5 kWh/tonne.
•
Flotation tests confirmed expectations of being able to produce a final concentrate grade
of 54% Mo grade and overall recovery of 89% - 90% from a mill feed of 0.084% Mo at a
flotation feed size of P80 = 210 microns. The 0.084% Mo mill feed grade is in line with
grade expectations for the first five years of operations.
•
The flotation tests done using small samples often did not allow five stages of cleaning to
be done because of insufficient production of concentrate as the cleaning stages
advanced. SGS-MinnovEX simulated the flotation data by means of its proprietary
FLEET algorithm which indicated that the flotation performance on an average mill feed
(0.084% Mo) would yield a 54% Mo grade at a 90.1% recovery based on primary grind
size of P80 = 180 microns. These data are conservative when compared to results from
the Kerr Addison pilot plant work (1969 – 71).
•
Mineralogical work done on SGS-MinnovEX’s behalf was superficial and brief and
threw no light on Adanac’s concerns for the seemingly lower overall molybdenum
recovery into a final concentrate.
December, 2007
- 54 -
•
SGS-MinnovEX recommended a flash flotation stage in the grinding circuit to remove
liberated MoS2 as soon as possible but were not able to simulate the effect this would
likely have on overall molybdenum recovery.
•
Very little gangue floats by attachment to the molybdenite into the rougher concentrate.
However, SGS-MinnovEX concluded satisfactory liberation still requires two stages of
regrinding, each to achieve P80’s of about 40 and 70 microns within staged cleaning.
(Regrind sizes are similar because cleaning removes the gangue leaving molybdenite
concentrated into the coarse fraction.
16.2.3 B.C. Mining Research Limited Test Work
Fine grinding tests were conducted at the University of British Columbia to assess alternative
technologies for regrinding the rougher flotation concentrate using either a Netzch (ISA) mill or a
Stirred Media Detritor (SMD). The testing shows:
•
Either mill, ISA or SMD, produced similar results. A comparison of different grinding
media indicated that the 2 mm ceramic media resulted in lowest power requirements.
•
Using 2 mm ceramic media, the power input required to grind from a P80 of 200 microns
to a P80 of 40 microns is about 30 kWh/t.
•
Molybdenite concentrate required high-energy (about 200 kWh/t) to grind from a P80 of
31 microns to a P80 of below 20 microns using either mill.
BC Mining Research noted stirred mills are usually operated in open circuit and they produce a
product with a narrow particle size distribution. At a P80 of 40 microns, the size distribution
would be similar to that produced with a ball mill in closed circuit with a classifying cyclone, but
at finer grind sizes stirred mills may produce a narrower size distribution which benefits flotation
and dewatering. The high stress intensities in stirred mills effect particle breakage.
16.2.4 G&T Test Work
A programme of tests to confirm expectations from SGS-MinnovEX’s work backed by detailed
mineralogy (including modal analyses) was done in the latter part of 2006. The work carried out
at G & T Metallurgical Services revealed that acceptable Mo concentrate grades and recoveries
are achievable at a coarser primary grind sizings of about 275 µm K80. This was in contrast to
simulated results from the SGS-MinnovEx work carried out on smaller samples. The projected
optimal primary grind sizing from the SGS-MinnovEX work suggested a finer optimal primary
grind sizing. The mineralogical data proved to be the key. Modal analyses presented details
confirming the degrees of mineral liberation and how these affected flotation performance. In all
cases (SGS-MinnovEX tests and G & T’s tests) every effort has been made to ensure full
representation of material being tested.
December, 2007
- 55 -
•
G & T confirmed the need for only one stage of regrinding (to a P80 + 20 microns)
•
High molybdenite recovery into rougher concentrates attained with low mass pull (<3%).
•
Only three to five stages of cleaning required to produce high grade and high recovery of
molybdenite into final concentrate.
•
Only lead noted to be an impurity which may have to be leached to achieve acceptable
limit.
The comprehensive testing done during 2005 – 2006 produced results which corroborate Kerr
Addison’s pilot plant results (1969 – 1971). High recovery of molybdenite into a high grade
concentrate containing minimal impurities should be expected from a full scale production
operation. Overall molybdenum recoveries in excess of 90% achieved in test work have not been
considered in the project economic modelling.
December, 2007
17.0
- 56 -
MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE
The July 23, 2007 Mineral Resource Update for the Ruby Creek Molybdenum Project was
calculated under the direction of Paul Palmer, P.Geo., P.Eng., of Golder Associates Ltd. as the
Qualified Person. Richard Gaze and Dr. Sia Khosrowshahi of Golder Associates Pty. assisted
with the geostatistical analysis and Mineral Resource Estimate under the direction of Paul Palmer.
This is the second time Paul Palmer has worked for and reported a mineral resource estimate for
the Project.
The July 23, 2007 Mineral Resource Update was based on a 3D geological and block model
constructed with commercial mine planning software (Datamine and Vulcan) based on the 2006
Datamine Database which includes all data collected from 2006 and earlier. No 2007 drilling
information is included in the July 23, 2007 Mineral Resource Update.
The Project limits were 588396 to 590896 East, 6619174 to 6621674 North and 700 m to 1900 m
elevation (NAD-83 UTM co-ordinates). The previous mineral resource estimates used drill hole
collars in the NAD-27 UTM co-ordinate system. The July 23, 2007 Mineral Resource Update
used drill holes collars in the NAD-83 UTM co-ordinate system. This initially required
transferring all the drill hole data (i.e. collar information) and all the previously created 3D
geological and mineralization geometries, and generating a new block model based on the NAD83 UTM co-ordinate system. The block model used in the July 23, 2007 Mineral Resource
Update was not rotated, measured 20 m east by 20 m north by 12 m (RL-elevation high) and subcelled to a minimum of 5 m by 5m by 3 m.
The Ordinary Kriging (OK) interpolation method was used for resource estimation of %Mo using
variogram parameters defined from the geostatistical analysis that was completed for the 2006
Mineral Resource Estimate. In addition, new variogram parameters were identified from the
structural data that was collected from the 2006 inclined drilling program and optical televiewer
analysis outlined in previous sections. Based on this analysis, the new variogram search
parameters were changed to a vertical direction search as opposed to a horizontal direction
search. The new variogram search parameters were only applied to the central deposit area,
which is where the highest drill density and underground data is located. The delineation was
accomplished by creating a new 3D mineralized domain (Zone 60) in the central pit area such that
the vertical variogram search parameters were applied to Zone 60 and horizontal search
parameters were applied to the mineralized domains outside of Zone 60 (partial Zone 6 and all of
Zone 7).
In addition, the 2006 drilling data (16 bore holes) were applied to the current 3D mineralization
geometries and modified to reflect the new data including slight interpretation modifications in
other area. Based on the updated 3D mineralization geometries, two mineral resource estimates
December, 2007
- 57 -
were completed. The first resource estimate (Model 1) is based on using the geostatistical
analysis defined from the 2005 and 2006 Mineral Resource Estimates, which used a variography
orientation that was horizontal. The second resource estimate (Model 2) is based on using both
vertical variography orientation for Zone 60 and horizontal variography orientation for Zones 6
and 7, outside of Zone 60.
Mineral resource estimates for measured, indicated and inferred resources were calculated for
each model using cut-off grades between 0.02 and 0.10 %Mo. The comparison of the tonnage
and grade estimates from each model indicated that the combined measured and indicated
tonnage for Model 1 was slightly higher (approximately 0.5% higher) when compared to Model 2
and the grades were the same. Therefore, Model 1 was considered still the most appropriate
method in estimating the mineral resource for the Project and was used in the July 23, 2007
Mineral Resource Update. The Model 2 mineral resource estimate supports the current estimates
and should be reviewed again if additional inclined drilling data is collected.
17.1
The Database
The database that was used in the July 23, 2007 Mineral Resource Update was constructed
initially in Datamine and was completed on drill hole information with the exception of historical
sample data from underground development. The majority of the data (approximately 60%) is
historical drill hole information (Kerr Addison, Climax and Placer) that has now been updated
with drilling programs by Adanac in 2004, 2005 and 2006 and includes a total of 71 drill holes.
The Datamine Database used in the mineral resource estimate is composed of 266 drill holes and
15,328 samples and is summarized in Table 17.1.
Table 17.1
Drill Hole Data in the July 23, 2007 Mineral Resource Update
Years
Drill Hole
Prefix
No. of Drill
Holes
No. of
Samples
Percentage of
Database (%)
1966 – 1972
KA
105+17
pseudo -drill
holes
6,073
39.5
1973
CL (100 Series
7
365
2.4
1979 – 1980
PD (200 Series)
66
3,145
20.5
Adanac Gold Corporation
(Adanac Moly Corp.)
2004
AD (300 Series)
36
2,886
18.8
2005
AD (300 Series)
19
1,621
10.6
2006
AD (300 Series)
16
1,238
8.1
266
15,328
100.0
Campaign
Kerr Addison (Adanac
Mining and Exploration,
John’s Manville and Kerr
Addison)
Climax Moly
Placer Development
Total
December, 2007
- 58 -
The database is comprised of vertical, inclined drill holes and underground sampling data
(pseudo-horizontal drill holes). Each sample in the database has the following attribute
information: hole ID, from and to sample distance, total sample length, %Mo, lithological
identification (primary and secondary), sample number, percentage recovery and RQD. Samples
in the overburden were generally not populated with %Mo, recovery and RQD and some earlier
samples did not collect RQD information. Sample lengths in the database were on average 10 ft
long (3.05 m). Any data not available was either left as blank in the original database or later
flagged with a negative code (-9 or -99).
Historical drill hole data in the Datamine Database have been outlined in previous sections of this
report, with additional details provided in the Technical reports by Blower (2005) and Palmer
(2006).
17.2
The Geological Model
The geological model that was generated for the Ruby Creek deposit comprised 3D wireframe
geometries of geological interpretations and a 3D block model created in Datamine and Vulcan
software. Six 3D geometries were created: three represented the main lithological units (primary)
and three represented the main molybdenum mineralization. One additional mineralization
geometry (Zone 60) was included in the July 23, 2007 Mineral Resource Update, which was
located in the central pit area. Two surfaces (DTM) were also used to represent the
overburden/bedrock interface and the topographic surface. The 3D block model was composed
of a 20 m (x-easting) by 20 m (y-northing) by 12 m (z-elevation) parent blocks that were further
subdivided into a minimum of 5 m (x) by 5 m (y) by 3 m (z) sub-celled blocks.
The 3D geometries used in the July 23, 2007 Mineral Resource Update were updated from the
January 2006 3D geometries based on the new information from the 2006 drilling program
(16 bore holes) and slight interpretation modifications in other areas.
The 2007 3D geometries were used for coding both drill hole data and the block model that
defined the spatial zones for the estimation of grades and bulk density assignment in the 2007
resource model (Model 1). The naming convention used for the 2007 2D and 3D geometries is
illustrated on Figure 17-1 (Local Grid 00) and summarized in Table 17.2.
December, 2007
- 59 -
Figure 17-1
Section View of 3D Mineralization Zones and Rock Types for Ruby Creek Deposit
December, 2007
- 60 -
Table 17.2
2007 Ruby Creek Molybdenum Project 2D and 3D Geometries
Field
Value
2D and 3D
Wireframe Geometry
Description
Rzone
1
ob_nov7.00t
Overburden Surface
Rzone
2
RZN2.00t
Rock Type 2
Rzone
3
RZN3.00t
Rock Type 3
Rzone
4
RZN4.00t
Rock Type 4
Zone
6
MIN_ZONE.00t
Mineralization Zone 6
Zone
7
zone7_nov7.00t
Zone
60
Rzone
19
Zone
-99
topo_clipSept27.00t
Topographic Surface
Background Mineralization
Rzone 1 represents the geometry between the bedrock and the overburden. Rzones 2, 3 and
4 represent a simplified 3D geometry interpretation of the rock types in the Ruby Creek deposit.
The lithological units that were classified inside Rzones 2, 3 and 4 are summarized in Table 17.3.
The rock types in Rzones 2, 3 and 4 are a simplification of the various lithological units in the
Ruby Creek deposit which, for resource modelling purposes, were acceptable to combine in order
to define densities and tonnages in the block model. Rzones 2 and 3 rock types are
geographically located south of the Adera Fault and Rzone 4 rock types are located
geographically north of the Adera Fault.
Table 17.3
Primary and Secondary Lithological Units
Text Code
Rock Type Code
(RZone)
Description
CGQM
2,4
Coarse-Grained Quartz Monzonite
CGQM-T
2,4
CGQM – Transition Variety
CQFP
2,4
Crowded Quartz Feldspar Porphyry
SQFP
2,4
Sparse Quartz Feldspar Porphyry
CGQM-H
2,4
CGQM – hybrid
MQMP
2,4
Mafic Quartz Monzonite Porphyry
SQMP
3
Sparse Quartz Monzonite Porphyry
CQMP
3
Crowded Quartz Monzonite Porphyry
FGQM
2,4
Fine-Grained Quartz Monzonite
MFP
2,4
Mafic Feldspar Porphyry
BSLT
2,3
Basalt
December, 2007
- 61 -
Zones 6, 7 and 60 represent the 3D geometry interpretations of the main molybdenum
mineralization in the Ruby Creek deposit. The criteria used to classify drill hole samples inside
the mineralization zones are summarized as follows:
•
%Mo greater than or equal to 0.04%;
•
primary or secondary rock identification as either FGQM or aplite dykes; or
•
presence of silicification alteration.
The 3D poly lines corresponding to mineralization zones were modified based on the 2006
drilling and only impacted Zones 6 and 60. Considerations were also made for continuity
between sections especially where data was not available. The 3D poly lines were then
constructed into 3D geometries and validated in Datamine and Vulcan software. Representative
sections are illustrated in Appendix A.
Zones 6 and 60 represent the main mineralization located south of the Adera Fault. Zone 60
(inside of Zone 6) also represents the main mineralization in the central area of the deposit in
which there is a greater density of sub-horizontal and sub-vertical veining and potentially a
breccia/feeder to the deposit. The breccia/feeder zone is located toward the centre of the deposit
and was reported to intersect the underground development sampling completed by Kerr
Addison. The smaller Zone 7 mineralization geometry is dominated by sub-horizontal veining
and molybdenum mineralization located north of the Adera Fault.
Rzone 19 represented the 2D topographic surface over the Ruby Creek deposit. All
Mo mineralization outside of Zones 6, 60 and 7 were coded as Zone -99 and represented the low
grade (%Mo less than 0.04%) background mineralization. Some mineralization pockets with
%Mo greater than 0.04 did occur outside of Zones 6, 60 and 7, but were not included inside these
geometries because they were localized high grade pods with no sample data to confirm
continuity from section to section.
17.3
Wireframe Validation
Prior to block model generation and populating of the block model, the Datamine 3D geometries
were validated in the Vulcan Software using standard wireframe validation check routines and by
slicing sections through the individual wireframes for comparisons against the drill hole database.
The process verified the following:
•
consistency, by testing triangle edges;
•
correct solid triangulation closure;
•
self-intersection, by testing for self-crossing triangles; and
December, 2007
•
17.4
- 62 -
correct spatial location of individual points within the wireframe.
Data Preparation and Compositing
Drill hole data for the Ruby Creek deposit was predominantly sampled on 10 ft lengths (3.05 m).
The global distribution of raw sample lengths is shown as a cumulative log-probability plot,
which is illustrated on Figure 17-2 from Palmer (2006). This plot illustrates that 95% of the data
is represented by the 3.05 m length.
Figure 17-2
2006 Global Distribution of Raw Sample Lengths (Palmer, 2006)
Considering the nominal most frequent raw sample interval (approximately 3 m) and the likely
vertical mining selectivity for the project, a 6 m down hole compositing interval was selected.
This length was a multiple of the nominal approximate 3 m interval and a divisor of the
anticipated mining bench height of 12 m, and the block model vertical dimension of 12 m.
Prior to compositing, the raw sample intervals were generated from the Datamine database and
flagged to the 3D geometries listed in Table 17.3 and assigned codes for mineralization zone and
rock type. The flagged intervals were then uploaded to a new flagged database.
December, 2007
- 63 -
From this flagged database, a 6 m composite file was generated for statistical and geostatistical
analysis using run-length down hole compositing. All composites were broken at the 3D
mineralized zone contacts (Zones 6, 60 and 7) resulting in some composite lengths of less than
6 m.
Any gaps in the down hole sequence representing unsampled intervals were excluded from the
compositing process, with no default values assigned to represent the missing sample intervals.
Due to the use of a length-weighting approach in the interpolation method used for resource
estimation, residual composites arising from the breaking of 6 m compositing at geological
contacts were not removed from the composite dataset.
No filtering of raw data on the basis of the sample recovery was carried out, mainly due to large
amounts of missing values in the recovery data. No relationship between sample recovery and
%Mo was evident from a statistical analysis.
17.5
Declustering
The exploration drill holes in the Ruby Creek deposit were on a nominal pattern of 90 m by 90 m
(approximately 300 ft by 300 ft) spaced drill holes, with in-fill drilling in the central region of the
resource down to 30 m by 30 m (approximately 90 ft by 90 ft) spacing. A location plan of the
drilling is shown on Figure 11-1.
Due to the clustered nature of the drill holes, spatial declustering was carried out using a 90 m by
90 m by 12 m moving window in order to achieve more representative global statistics. A
standard cell declustering algorithm was applied to determine the declustering weights.
17.6
Spatial Trend Analysis
Spatial trend analysis was carried out independent of domain contacts in 2006 to assess grade
trends in %Mo content across the resource and is provided in Palmer (2006). This was done
using fixed block averages, where 6 m composite data was averaged into 90 m by 90 m by 12 m
blocks across the resource, and the mean block grades displayed spatially. This provides an
indication of whether there are particular grade orientations or zonation of elevated grade that
need to be accounted for in the resource estimation. The spatial trend analysis showed some
indication of a South-Easterly grade trend.
17.7
High-Grade Treatment
High-grade treatment for estimation was applied using a spatial restraining method to avoid overestimating the grade of %Mo in the resource domains. In this method, samples above a
December, 2007
- 64 -
designated threshold are flagged to individual blocks in the model and their spatial influence
restricted to a single block. Therefore, instead of capping high grade values, they were restricted
spatially.
Preliminary high-grade thresholds were selected from an examination of cumulative
log-probability plots for each domain, and refined through estimation validations of the
reproduction of the estimated mean domain grade in the Kriging. After the preliminary estimate
and model validations were run, high-grade thresholds were adjusted where required, and the
model re-estimated.
A summary of the high grade thresholds implemented in the grade estimation is provided in
Table 17.4.
Table 17.4
High-Grade Thresholds for %Mo by Zone
17.8
Mineralized Zone
%Mo High-Grade
Threshold
Approx. Distribution
Percentile
6 and 60
0.3
99.5
7
0.2
97.5
-99
0.15
99.7
Variogram Analysis
17.8.1 Variography Objectives and Approach
The objectives of the variographic analysis were the following:
•
•
to establish the directions of major grade continuity for each element in the domains; and
to provide variogram model parameters for use in geostatistical grade interpolation.
The variographic analysis was competed using in-house Golder software and was based on the
6 m composited %Mo data for the combined Zones 6, 60 and 7. Conventional 3D directional
variography was used for spatial continuity analysis using 3D conical search methodology.
Spherical scheme models were used for modelling of the major, semi-major and minor orthogonal
directions of continuity. These models were generated during the 2006 Mineral Resource
Estimate and are provided in Palmer (2006). The variogram parameters used in the February 22,
July 23, 2007 Mineral Resource Update were taken directly from the 2006 analysis in Palmer
(2006). In addition, a new variogram analysis was used, based on the structural data from the
2006 inclined drilling program. The new variography analysis was applied to the mineral Zone
60 to determine if using the new variogram parameters better defined the grade of the mineral
December, 2007
- 65 -
resource estimate in this area. The results from the new variogram analysis were not significantly
different than the 2005 variogram analysis. The grades from Zone 60 were not significantly
different; therefore, the 2005 variogram analysis (horizontal variogram) from 2006 was used in
the February 22, 2007 Mineral Resource Estimate.
17.8.2 Summary of Variography Parameters
Table 17-5 provides a summary of variography model parameters and orientations modelled for
the combined mineralized Zones 6, 7 and 60, determined in the 2006 and 2007 Mineral Resource
Estimates.
Linear orientations are defined in terms of plunge and plunge direction (i.e. plungeÆplunge
direction). Resultant planes are defined in terms of dip and dip direction (i.e. dipÆdip direction).
The ‘separation’ is the angle between the major and semi-major vectors. The ‘rotation’ applies to
the semi-major axis; it is the angle required to rotate the search ellipsoid around the major axis
vector to align the ellipsoid with the interpreted resultant plane for grade interpolation.
Table 17.5
Variography parameters for Zones 6, 60 and 7
Axis Direction
Major Axis
Semi-Major Axis
Minor Axis
Mineralization
Zone
6, 60 and 7
Correlogram Parameters for %Mo Grades
Structure 1
Structure 2
Nugget Diff. Sill Range Diff. Sill Range
0.300
0.340
42.50
0.320
125.00
0.300
0.200
39.00
0.580
137.00
0.300
0.510
15.50
0.170
98.50
Major Axis
Semi-Major
Axis
Plunge/Direction
Plunge/Direction
0 Æ 140
0 Æ 050
Resultant Plane
Dip/Dip
Direction
0 Æ 050
KEY:
R1/R2/R3 Search orientations (conventional left hand rule)
R1 = azimuth rotation clockwise from north
R2 = plunge along R1 direction (+ve = up, -ve = down)
R3 = dip rotation around R1-R2 axis (+ve = anticlockwise, -ve = clockwise}
17.9
Update Block Model Parameters
Nugget/Sill
31
28
31
Major & SemiMajor
Rotation
Angles
Separation Angle
R1/R2/R3
90
140/0/0
A block model was constructed to cover the mineralization in the Ruby Creek deposit and
sufficient surrounding waste for inclusion into open pit designs. The dimensions of the model are
December, 2007
- 66 -
provided in Table 17.6 and illustrated including mineralization Zone 6 (green) and 7 (yellow) on
Figure 17-3.
Table 17.6
Block Model Dimensions for Ruby Creek Resource Model
Minimum
Maximum
Block size
(m)
No. of
blocks
Sub-block size
(m)
Easting (X)
588 396
590 896
20
125
5.0
Northing (Y)
6 619 174
6 621674
20
125
5.0
RL (Z)
700
1900
12
100
3.0
Figure 17-3
Plan View of Block Model Geometry
Blocks were filled inside each of the wireframed geological interpretations in a priority sequence
to create the geological zones.
Two separate variables were established for mineralized zone (Zones 6, 60, 7 and -99) and rock
type (Rzones 1, 2, 3, 4 and 19). Grade interpolation was constrained to the mineralized zone
interpretations. Rock type was used for applying the density and corresponding tonnage.
December, 2007
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Sub-blocks were used in each model to define the geological zone boundaries and the
topography. Sub-blocks were estimated to the parent cell to help achieve acceptable local
estimation quality.
17.10 Grade Interpolation
17.10.1
Grade Interpolation Methods and Objectives
The OK interpolation method was used for the resource estimation of %Mo for the Ruby Creek
Molybdenum Project using variography parameters defined from the geostatistical analysis.
The resource estimation reflects an in situ resource estimate based on a nominal block selectivity
of 20 m by 20 m by 12 m. No allowance has been incorporated for higher SMU selectivity or ore
loss and dilution assumptions.
17.10.2
Ordinary Kriging Plan
The Kriging plan parameters used for grade interpolation %Mo are summarised in Table 17.7
with details provided as follows:
•
All mineralized zones were estimated individually with their own data (hard boundary
conditions). Waste zones were estimated with parameters from the mineralized zone.
•
Sample weights were determined by the variogram model parameters (kriging approach).
•
Block discretisation was set to 5 (x) by 5 (y) by 3 (z) to estimate block grades of 20 m by
20 m by 12 m parent blocks. Estimation of sub-cells in the model was performed to the
parent cell size, so sub-cells received the parent cell estimate of 20 m by 20 m by 12 m.
•
A maximum of four samples per discretised block was used, which equates to a
maximum of 32 samples per estimate.
•
OK was applied in three passes. Pass 1 used a search of 70 m by 70 m by 12 m. Pass 2
used a search of 140 m by 140 m by 24 m to estimate blocks not estimated in Pass 1, or
blocks estimated in Pass 1 with less than two drill holes. Pass 3 used a 280 m by 280 m
by 24 m to estimate blocks not estimated in the first two passes, but using a minimum of
one sample. Pass 3 was designed to fill any remaining gaps in the model. Pass 3 in the
waste zone (zone=-99) was set to the same dimension as the second pass, but with a
minimum of two samples to limit grade extrapolation.
•
Estimation was weighted by the sample length to account for variations in sample length
due to the compositing process.
December, 2007
•
- 68 -
High grades above the thresholds were spatially restrained to a single block for grade
estimation to control the influence of the high grades. Therefore, high-grade values were
restricted to the 20 m by 20 m by 12 m parent block.
Table 17.7
Ruby Creek Deposit Kriging Plan Parameters
Estimation Method
Ordinary Kriging
Search radius (pass 1/2/3)
X
Y
Z
70/70/12
140/140/24
280/280/24
Anisotropy (OK)
X
Y
Z
Discretisation
X
Y
Z
Search type
Minimum No. samples
Maximum No. samples per octant
Defined by Variogram
5
5
2
Octant
4/4/1
4
17.11 Density Assignment
Adanac supplied Golder with the results of specific gravity samples that were analyzed by ALS
Chemex. In all, there is approximately 1,000 samples collected from the 2004 to 2006 drilling
programs. A statistical review was completed on the specific gravity samples (Palmer, 2006)
based on lithological unit and rock type zone (Rzones 2, 3 or 4). The minimum, maximum and
mean specific gravity estimates are summarized in Table 17.8. The mean bulk densities were
assigned to all blocks in the model based on rock type using the values in Table 17.8.
December, 2007
- 69 -
Table 17.8
Rock Type Bulk Density Assigned to the Block Model
Rock type
Min Bulk
Density
Max Bulk
Density
Mean Bulk
Density
Default rock type
2.57
Rzone 2
2.19
2.88
2.57
Rzone 3
2.22
2.63
2.57
Rzone 4
2.37
2.71
2.55
All the specific gravity samples for the current Datamine Database (approximately 22% of
database) were based completely on the 2004 - 2006 drilling samples. No specific gravity
information was available from the historical data. Therefore, a representative amount of specific
gravity samples have been collected from the 2004 - 2006 drilling programs. Additional samples
should be collected from any future drilling programs to continue to expand the dataset,
specifically in areas where no specific gravity sample data has been collected.
17.12 Mineral Resource Classification
Classification of the resource estimate was based principally on sample data density and
geological confidence criteria. Consideration was also given to the classification based on the
two previous mineral resource estimates (Blower, 2005 and Palmer, 2006) in order to achieve
some continuity in terms of changes in proportions of resource categories with the additional
drilling.
The previous mineral resource estimates used a confidence interval scheme for classification as
follows (Blower, 2005):
•
Measured Resource: estimates within +/- 15% at 90% confidence for quarterly
production
•
Indicated Resource: estimates within +/- 15% at 90% confidence for annual production
The application of this method indicated that the following drill densities were applicable with
respect to resource categories:
•
Measured Resource: drill hole spacing of 30 by 30 m or less
•
Indicated Resource: drill hole spacing of >30 by 30 <90 by 90 m
•
Inferred Resource: drill hole spacing of greater than 90 by 90 m.
December, 2007
- 70 -
Based on the mineralization style at Ruby Creek and the continuity identified from the
variography studies, the above drill hole spacings were considered an acceptable method for
resource classification of the Ruby Creek deposit and were used in the July 23, 2007 Mineral
Resource Update.
The July 23, 2007 Mineral Resource Update included resource classification by digitising
polygons on cross-section for both Indicated Resources (drilling <= 90 by 90 m) and Measured
Resources (drilling <= 30 by 30 m) to better define the initial classification boundaries.
The following conditions were also applied to the mineral resource estimate (using Vulcan
scripts):
•
Initially, all blocks were set to an Inferred Resource;
•
Blocks inside the indicated 3D geometry were set to Indicated Resource;
•
Blocks inside the measured 3D geometry were set to Measured Resource;
•
Measured blocks were downgraded to an Indicated Resource if the average weighted
interpolation distance was greater than or equal to 50 m or the number of drill holes used
to estimate the block was less than four;
•
Measured blocks initially defined by the measured 3D geometry but located outside the
main mineralization zones (Zones 6, 60 and 7) were downgraded to an Indicated
Resource; and
•
All blocks estimated with less than two drill holes per block were downgraded to Inferred
Resource.
The following resource classification codes were used in the block model:
•
Measured Mineral Resource (class = 1);
•
Indicated Mineral Resource (class = 2); and
•
Inferred Mineral Resource (class = 3).
17.13 Mineral Resource Summary
The Ruby Creek Mineral Resource Estimate is summarised in Table 17.9 reported at %Mo cutoffs from 0.02 to 0.10% (resource tabulation convention is greater than or equal to the cut-off).
December, 2007
- 71 -
Table 17.9
July 23, 2007 Mineral Resource Update
Resource Category
Measured
Indicated
Measured
+
Indicated
Inferred
Cut-off (%Mo)
Tonnage
%Mo
Mo lb
0.020
55,831,000
0.068
83,698,000
0.030
54,300,000
0.069
82,600,000
0.040
49,106,000
0.073
79,029,000
0.050
41,389,000
0.078
71,172,000
0.060
30,151,000
0.086
57,165,000
0.070
21,909,000
0.094
45,403,000
0.080
14,556,000
0.104
33,374,000
0.090
10,411,000
0.112
25,706,000
0.100
6,612,500
0.122
17,785,000
0.020
387,278,000
0.042
358,593,000
0.030
238,954,000
0.052
273,935,000
0.040
163,801,000
0.060
216,669,000
0.050
109,444,000
0.067
161,658,000
0.060
61,471,000
0.077
104,350,000
0.070
37,664,000
0.084
69,749,000
0.080
18,813,000
0.094
38,987,000
0.090
9,848,100
0.102
22,145,000
0.100
4,286,900
0.113
10,680,000
0.020
443,108,000
0.045
442,290,000
0.030
293,254,000
0.055
356,535,000
0.040
212,907,000
0.063
295,699,000
0.050
150,834,000
0.070
232,831,000
0.060
91,621,000
0.080
161,513,000
0.070
59,573,000
0.088
115,151,000
0.080
33,369,000
0.098
72,360,000
0.090
20,259,000
0.107
47,851,000
0.100
10,899,000
0.118
28,464,000
0.020
135,737,000
0.032
95,759,000
0.030
48,456,000
0.045
48,072,000
0.040
24,973,000
0.054
29,730,000
0.050
11,631,000
0.064
16,411,000
0.060
5,194,300
0.077
8,817,500
0.070
2,626,200
0.089
5,152,900
0.080
1,239,700
0.103
2,815,000
0.090
821,310
0.113
2,046,000
0.100
493,790
0.125
1,360,800
mineral resource estimate has been completed in accordance with CIM standards of Estimation of Mineral Resources
and Mineral Reserves.
December, 2007
- 72 -
17.14 Mineral Reserve Estimate and Mine Design
The final depth of the proposed open pit is planned to be at an elevation of 1,228 m. The North
West rim of the pit will reach Molly Lake (El. 1,600 m). Molly Lake will be drained before the
open pit mining is begun.
Open pit mining will initially be performed using a diesel-powered rotary drill and a hydraulic
shovel with a 22.0 m3 as the primary loading tool with two 12.0 m3 front end loaders (FELs) as
secondary loading tools. A fleet of six 150 ton capacity end-dump haul trucks will be used for
ore transport. Road, dump and pit bench maintenance will be carried out with a fleet of
compatible support equipment. After grid electric power becomes available, a second rotary drill,
(electric) will be added to the fleet, as well as 22.0 m3 electric front shovel and seven additional
150 ton capacity end-dump haul trucks. The rock will be drilled, blasted, loaded and hauled to
either the primary crusher, in the case of ore or designated locations for stockpile ore and waste
rock.
The comminution circuit includes a single gyratory crusher followed by two secondary cone
crushers in closed circuit with a screening plant. Cone crusher discharge feeds two high pressure
grinding rolls (HPGRs) and a single ball mill designed to produce material for flotation. The
flotation circuit includes rougher/scavenger and cleaner stages. To increase the degree of
liberation and improve overall molybdenum recovery, a single regrinding stage is included within
the cleaning section. The final concentrate is thickened, filtered, dried, packaged and shipped
offsite for roasting. Tailings will be impounded and water recycled into the process.
A detailed plan for the property is currently being developed. This plan includes a construction
camp, fresh water supply and sewage collections/treatment, power generation and distribution,
fuel storage, communications, tailings impoundment, waste dump sites and accommodation for
operating personnel, has been developed as an integral part of project development.
Construction is currently underway and the planned start up is scheduled in the year 2009. A
construction work force is estimated to peak at 700 people. A summary schedule is provided in
Figure 17-4.
December, 2007
- 73 -
Figure 17-4
Summary Schedule
The Golder Mine Feasibility Report completed in March 2006 served the basis of parameters for
the detailed mine design work described in this report. For example, the geotechnical work and
analysis required for the open pit design were undertaken by Golder in the Summer of 2005 and
results are contained in their feasibility report.
Open pit optimizations were completed using a range of molybdenum prices from US$5.00 per
pound to US$10.00 per pound. Previous work used molybdenum prices up to US$13.00 per
pound to guide the location of mine facilities and related surface infrastructure. The open pit
optimizations were performed with overall angles of 45 - 48.5° based on the geotechnical
evaluation.
A molybdenum base price of US$10.00 per pound was used for all tonnage calculations and cutoff grade determinations. At this price, the calculated mining cut-off grade was 0.040% Mo and
the processing cut-off grade (internal cut-off) was 0.030% Mo. The mining cut-off was set at
0.040% Mo and low grade material grading between 0.030% and 0.040% stockpiled for
processing later.
December, 2007
- 74 -
A series of four phases or push-backs have been designed for the ultimate pit . The revised open
pit design extends south southwest to incorporate reserves of the phase 5 push back identified in
the Golder March 2006 Feasibility Report.
Phase 1 focused on the near surface higher-grade portion to maximize grade and subsequent early
cash flows, thereby minimizing the payback period, increasing internal rate of return reducing
other financial risks.
All pit phase designs were designed to accommodate a truck size of 200 tonnes capacity (which
allows for a possible upgrade from the currently contemplated fleet of 150 tonne trucks, later in
the mine life), utilizing 30 metre ramp width. The mining schedule uses a cut-off grade of 0.060%
Mo for Phase 1 and 2 to generate a payback pit and speed repayment of mine capital. The
remaining phase 3 and 4 were completed with a mining cut-off of 0.040% Mo and an internal cutoff grade of 0.030% Mo for the low-grade stockpile throughout. The ultimate pit design can be
seen in Figure 17-5.
December, 2007
- 75 -
Figure 17-5
Ultimate Pit Design
The production schedule provides for 20 years of mining and 21 years of processing at a rate of
7.6 million tonnes per year. Based on the mining cut-off, production plan and schedule a Mineral
Reserve was determined. This mineral reserve is presented in Table 17-10, and is estimated at
157.6 million tonnes grading 0.058% Mo.
December, 2007
- 76 -
Mining dilution for the Ruby Creek deposit was assessed at 5.1% along with an accompanying
ore loss of 3% for a net tonnage gain of 1.9% above the undiluted tonnage. Undiluted grades were
subsequently reduced with a 0.020% Mo diluting grade for the 5.1% dilution tonnage.
Table 17.10 illustrates the reserve distribution by phase and classification.
Table 17.10
November 22, 2007 Mineral Reserve Estimate
Phase 1
Phase 2
Phase 3
Phase 4
Total
Ore to Mill
Proven
Probable
Tonnes
%Mo
Tonnes
%Mo
19,455,000 0.089 3,065,000 0.082
3,903,000 0.070 6,819,000 0.075
20,250,000 0.056 48,463,000 0.050
271,000
0.049 29,166,000 0.056
43,879,000 0.072 87,513,000 0.055
Stockpile Ore
Proven
Probable
%Mo
Tonnes
%Mo
Tonnes
6,081,000
0.049
2,608,000 0.042
4,996,000
0.049
7,519,000 0.046
185,000
0.027
2,893,000 0.026
12,000
0.027
1,999,000 0.026
11,274,000 0.049 15,019,000 0.039
Total
Tonnes
31,209,000
23,237,000
71,791,000
31,448,000
157,685,000
A production schedule was developed, that ensures mill feed continuity during and between
mining phases. Table 17.11 shows the mining and processing schedule by year.
%Mo
0.077
0.059
0.051
0.054
0.058
December, 2007
- 77 -
Table 17.11
Ruby Creek Production Schedule
December, 2007
- 78 -
Table 17.11
Ruby Creek Production Schedule (Con’t)
December, 2007
- 79 -
Figure 17-6 and 17-7 present the actual phase (pushback) development sequence in plan and
section, respectively. In general terms, mining proceeds from the highest value ore to the lowest
value ore. The phase 1 pit is located in the northeast part of the deposit. Phase 2 ,3 and 4 expand
the open pit to the south and southwest and deepen the northeast lobe to its ultimate depth.
Specific protocols have been included to account for the identification, handling and disposal of
potentially acid generating waste.
Figure 17-6
December, 2007
- 80 -
Figure 17-7
The expected mine manpower has been detailed in Table 1712. Four departments fall under the
Mine Operations area, as follows:
•
Maintenance;
•
Operations; and
•
Engineering/Geology.
The total number of positions for the initial five years is estimated to be 115 people. This allows
for a more flexible rotation schedule while maintaining operations continuity.
December, 2007
- 81 -
Table 17.12
Mine Department Personnel
Description
Number of Persons
Mine Operations
Mine Superintendent
1
Mine Shift Foreman
4
Drill/Blast Engineer
1
Training Foreman
2
Shovel/FEL Operators
5
Truck Drivers
29
Rotary Drill Operators
5
Support Equipment
16
Blasters & Helpers
3
Trainees/Labourers
2
Subtotal:
68
Mine Maintenance
Maintenance Superintendent
1
Maintenance Planner
1
Shop Foreman
4
Maintenance Clerk
1
Mechanics
15
Welders
4
Apprentices
4
Light Duty Mechanics
2
Tire Boy
2
Lube/Service & Fuel Truck
4
Labourers
2
Tool Crib/Warehouse
0
Subtotal:
40
Engineering/Geology
Chief Engineer
1
Chief Geologist
1
Grade Control Geologist
1
Mine Planner
1
Technician/Surveyor
3
Clerk
0
Subtotal:
7
Total:
115
December, 2007
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18.0
OTHER RELEVANT DATA AND INFORMATION
18.1
Tailings Facilities, Waste Rock Dumps and Site Water Management
Detail design of the tailings facilities and site water management are presented in Klohn
Crippen's reports “Site Water Management Design Report” dated February 19, 2007 and
“Tailings Facility Detail Design Report” dated April 5, 2007. Feasibility level geotechnical
design of the waste rock dumps is presented in Klohn Crippen's report “Feasibility Design of
Tailings Facility, Waste Dumps and Site Water Management” dated February 8, 2006.
The site investigations carried out in 1979/1980 were extensive and provided most of the required
information for design purposes. Additional site investigation work was carried out in 2005 and
2006 to augment this information.
The plant, tailings facility and waste rock dumps will be located near the open pit in the upper
basin of Ruby Creek. The main features of the feasibility design are provided in the following
sections.
18.1.1 Tailings Characterization
The tailings produced by the mill will be relatively coarse with nominally 26% to 33% fines
content. The Neutralization Potential Ratio (NPR) among all the ore types tested ranged from 7 to
42, with a median of 23. The tailings are, therefore, concluded to have a very low potential for
producing acidic drainage that would liberate metals. This is consistent with the low sulphide
sulphur content of the tailings (less than 0.03%). Additionally, the existing pilot plant tailings
facility has remained in a neutral state for 35 years.
Testing of tailings supernatant water show that most elements meet the BC Water Quality
Guidelines (BCWQG) for the protection of freshwater aquatic life. Aluminium and iron are the
only elements that have median values exceeding BCWQG.
The tailings facility plan allows pond water to be discharged to the environment only during peak
spring runoff. With a 20 times dilution rate, the median and maximum concentrations of all
elements fall below the BCWQG. Nevertheless, the tailing facility is suitably sized to store all the
tailings water during the life of the mine with no release to the environment, if required.
18.1.2 Tailings Facility
A compacted cyclone sand tailings dam will be constructed across Ruby Creek to form the
tailings storage facility downstream of the plant. The tailings will be transported by slurry
December, 2007
- 83 -
pipeline at a rate of 21,000 to 23,000 tonnes of solids per day. Free water will be reclaimed from
the pond by a floating pump-barge and returned to the plant. Water reclaim will be maximized to
reduce the volume of stored water.
A seepage recovery dam will be constructed downstream of the tailings dam. It will also serve as
a settling pond for any sediment transported by runoff and dam construction, and provide a
monitoring station for water quality sampling.
At the end of operations, 94 Mm3 of tailings (excluding cyclone sand used for dam construction)
and 30 millions tonnes PAG waste deposited in the facility will reach an elevation of 1394 m. The
total volume of free water will be 7 Mm3, giving a final average tailings/water at an elevation of
1397 m.
The required dam crest elevation for the base case water balance is 1401.1 m. However, for
facility layouts, an ultimate dam crest at 1418.2 m has been selected to accommodate a
conservative water balance with incomplete diversion of the drainage catchments. The ultimate
dam could also provide for 4 more years of mining reserves.
The tailings dam will be built from compacted cyclone sand dam constructed by the centreline
method with a crest width of 30 m, and downstream slope of 3H:1V. The maximum dam height is
140 m and the final crest length is 1500 m. A system of finger drains will be installed at the base
of the cyclone sand to keep the water level in the dam depressed.
A starter dam of compacted till will be constructed at an elevation of 1340 m to store 1.5 Mm3 of
water for mill start up and to store the first year of tailings production. Subsequent dam raises will
be constructed using cycloned tailings compacted in cells. The technique of "cell construction" is
well established and has been used at Highland Valley Copper, Kennecott Utah Copper, Southern
Peru Copper and in the oil sand industry.
All vegetation and soil beneath the dams and tailings basin will be stripped for site reclamation at
closure. The exposed surface extending 500 m upstream of the tailings dam will be inspected for
highly pervious zones (talus or colluvial materials) that could cause increased seepage or piping
of tailings into the foundation. Any highly pervious zones will be excavated and replaced, or
suitably covered by lower permeability compacted till soils.
Several localized pockets of loose soils were encountered at shallow depths of less than 5 m.
Prior to dam construction, additional drilling investigations will be carried out to assess the aerial
extent of these loose zones. If required, the loose zones can be easily excavated. The final ground
surface beneath the dam will be compacted.
December, 2007
- 84 -
Tailings deposited from the crest of the dam will maintain a separation of 400 m to 1000 m
between the dam and the water pond. This will limit seepage and keep water levels in the
pervious cyclone sand dam very low. Seepage into the pervious basalt in the base of the valley
will be restricted by the overlying cover of overburden soils, augmented over time by the 75 m
depth of tailings. Analyses indicate 10 L/s to 33 L/s of seepage through the dam and foundation
will occur under normal operating conditions. This is consistent with experience at the Gibraltar
and Highland Valley Copper tailings facilities.
The static and dynamic stability of the tailings dam exceeds Factors of Safety recommended by
the Canadian Dam Association.
18.1.3 Waste Rock Dumps
Geo-chemical testing indicates that most of the waste rock will not be acid-generating, with a
median NPR in excess of 3.7. This conclusion is corroborated by the neutral paste pH of the rock
samples (typical paste pH 6.8 to 9.1) and the low sulphide sulphur contents (median = 0.04%) of
the waste rock.
Detailed evaluation of the deposit geology and NPR results suggests that select zones of the pit
may product up to 30 Mt of potentially acid-generating (PAG) waste rock. PAG will be placed in
the tailings impoundment so that it is permanently submerged within 1 year of placement.
Non acid-generating (NAG) waste rock and overburden will be deposited in small dumps located
in valleys above the open pit and in a large main dump to the east of the open pit at the upper end
of the tailings facility. The low-grade ore will be stockpiled and will be integrated with the larger
waste dump which is nearest the plant.
Pro-active measures will be implemented during dump operations to minimize the mobilization of
metals under neutral leaching conditions. Upslope diversion ditches will be constructed to convey
runoff around the dumps and completed dumps will be immediately capped with overburden soils
to minimize infiltration.
Final reclamation of the dumps will occur during the milling of low-grade material in the last four
years of mill operation. The waste dumps will be re-sloped and terraced to safe long-term angles
while meeting closure land use objectives. Dump surfaces will be capped with an overburden soil
cover to shed water to a perimeter ditch that will route the water away from the dumps.
The foundations at the dumpsites are composed of competent glacial tills, coarse colluvium or
bedrock. All sites are rated as having a low failure hazard according to the British Columbia
Dump Stability Classification System. Calculated static factors of safety exceed 1.3 for operating
December, 2007
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conditions and 1.5 for closure conditions. Dump displacements for a 475-year return period
earthquake will be less than 1 m.
18.1.4 Site Water Management Plan
Two major diversions will route natural runoff around the open pit, waste dumps and tailings
disposal facility. Both ditches will originate above the pit area and run generally parallel to Ruby
Creek on the flanks of the surrounding hillsides, rejoining Ruby Creek downstream of the
seepage recovery dam. The channels are sized to convey the 200-year return period peak flow
with suitable freeboard to account for variations in hydraulic performance. Riprap armouring is
provided on steep channel sections where peak flood velocities may be sufficiently high to cause
erosion of the native till soils.
The diversion channels will be constructed on variable till, colluvium and talus. Where required,
the base of the channels will be lined with compacted till and/or an 80 mil LLDPE to restrict
leakage.
A net total of 86 Mm3 of undiverted water from the project site will be sent to the tailings facility
over the 22 year operating life. The vast majority of the inflow will be retained in the voids of the
tailings and submerged PAG waste rock (66%), and as free water in the tailings pond (8%). The
remaining losses are evaporation (13%), groundwater seepage (10%) and controlled releases to
Ruby Creek (3%).
A total of 264 Mm3 of pond water will be re-circulated to the mill over the 22 year operating life.
For comparison, only 9.5 Mm3 will be impacted by the waste dumps in the tailings catchment.
Hence, the tailings supernatant should govern the water quality in the tailings pond.
A dam freeboard allowance is provided for all dams to accommodate the 200-year return period
30-day storm plus an additional 3 m to allow for the routing of a 30-day PMF (assuming the
diversions fail) through an emergency spillway.
A permanent main spillway channel will be constructed on the right abutment of the tailings dam
to convey extreme floods from the tailings facility during operation and to convey the flows from
the tailings basin at closure. A spillway is also provided in the left abutment of the seepage
recovery dam. All spillway structures are designed for a peak flow of 35 m3/s produced by the
routing of the Maximum Probable Flood (PMF) from the entire undiverted catchment of the mine
site through the tailings facility.
During operation, an operating spillway will be excavated at the right abutment of the tailings
dam to route the flood flows to the main spillway channel. The initial operating spillway will be
December, 2007
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installed after completion of the tailings starter dam construction. For subsequent dam raises,
operating spillways would be excavated only in an emergency when a trigger water level is
reached (when the pond level reaches within 5 m of the dam crest). These spillways will be 5 m
wide and excavated 3 m below the tailings dam crest so that a 1 m minimum dam freeboard is
maintained above the peak PMF flood level. Riprap armouring for the spillways will be
permanently stockpiled at the right abutment of the tailings dam for immediate use.
At closure, the permanent spillway channel will be extended into the tailings basin along the right
abutment of the tailings dam. The base elevation of the channel will be set to provide a 700 m
minimum separation between the tailings pond and dam crest under average flow conditions, and
maintain a 3 m minimum dam freeboard above the peak PMF flood level.
18.1.5 Operating and Monitoring Controls
The tailings dam and seepage recovery dam will be monitored by piezometers and settlement pins
to record the phreatic levels in the dams and foundation soils, and to measure the deformations of
the structures.
Pond levels will be recorded monthly and used, in conjunction with pond filling curves, to plan
the operation of the facility. The performance of the diversion channels and the operation of the
spillways will be monitored and reported on a regular basis. The proposed monitoring and
operational controls include:
•
Establishment of flow gauging stations and survey of high water marks to confirm the
flow capacity of the channels and the design relationships between precipitation and
flood flows.
•
Weekly inspections of all channels and spillways to visually assess performance and
provide advance warning of excessive erosion or seepage through the base and banks of
the channels.
•
Daily inspections of the channels above the open pit which could represent a hazard if a
breach of the channels diverts flow into the pit.
•
Each spring before and during snow melt, the diversion structures will be inspected and
cleared of snow and ice as required to ensure proper operation.
The emergency spillway will be constructed using stockpiled materials, if the pond rises to within
5 m of the dam crest. Records of pumping from the seepage recovery pond to the tailings pond
will be kept to assess the seepage rate from the tailings facility.
Surface water sampling and flow quantities will be monitored at various points through the mine
area to evaluate water quality and maintain a site wide water balance. The water quality and flows
December, 2007
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will be evaluated to determine the allowable controlled discharge of tailings pond or seepage
water to Ruby Creek, primarily during spring runoff.
Groundwater quality will be monitored at wells located downstream of the seepage recovery dam.
This will be assessed in conjunction with Ruby Creek water quality to determine the need (if any)
for installing pumpback wells.
18.2
Mine Closure Plan
A preliminary mine closure plan is presented in the Environmental Assessment Certificate
(“EAC”) Application for the project. Adanac received the Project EAC on September 10, 2007.
Details of the mine closure plan are summarized below.
18.2.1 Tailings Facility
The diversion channels will be decommissioned to re-direct flow back into the tailings facility.
The seepage recovery dam will also be breached to prevent water storage. A 10 m wide ripraplined closure spillway will be constructed in the left abutment of the tailings dam. The spillway
channel will extend along the left abutment to the edge of the tailings pond. The invert elevation
of this spillway will be set to maintain a 700 m minimum separation between the tailings pond
and dam crest (under average flow conditions) and designed to safely pass the probably
maximum flood (PMF) from the entire upstream catchment.
The tailings are not anticipated to be acid-generating and exposed tailings beach and dam surfaces
will be suitably reclaimed to prevent erosion and meet post-closure land uses targets. The 3H:1V
exterior slope of the dam is flatter than the 2.5 H:1V maximum slope adopted for reclamation of
cyclone sand dams at other mine sites.
Natural runoff within the catchment basin will dilute and improve the water quality of the tailing
pond supernatant, with a full 10 times dilution achieved in about 10 years. More rapid dilution
could be achieved by pumping all or part of the pond water into the open pit at the end of mining.
The expected annual seepage through the tailings dam is ten times less than the annual volume of
water discharged through the closure spillway and, therefore, will be subjected to substantial
dilution.
18.2.2 Waste Dump/East Ruby
The East Ruby waste dump will be re-sloped and terraced to a safe long-term stable angle while
meeting closure land use objectives. The dump will be reclaimed as soon as practical as each
December, 2007
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portion of the dump becomes inactive. Final reclamation of the dump will occur in the final four
years of milling low-grade ore.
To minimize infiltration of surface water, the top of the East Ruby Waste Dump will be graded
and capped with glacial till to shed water to a perimeter surface water collection channel. The
channels will also intercept runoff from the catchment slopes above the waste dump. The water
will be routed to the open pit or the closed tailings facility.
Overburden soils from the dump will be stored for use in site reclamation of the waste dump and
tailings dam, as required. The remaining portions of the waste dump will be suitably reclaimed.
To reduce the effects of neutral leaching, the dump surface will be graded and capped with glacial
till to shed water to outer edges of the dump.
18.2.3 Open Pit
The open pit will ultimately fill to become a small lake. Pit walls will be left in a safe and stable
condition. No re-vegetation program is planned for the pit walls as they will be bare bedrock
exposures as currently exist over much of the pit area. Surface runoff from above the pit will
accumulate in the pit to a maximum elevation of about 1425 m, the approximate elevation of the
outlet at the eastern rim. A channel leading from the eastern rim to the tailings facility will be
constructed to ensure long-term stability and positive control of the outflow to the closed tailings
facility.
18.2.4 Temporary Mine Closure
Temporary mine closure for periods longer than one year will require that measures, as outlined
below, be implemented to mitigate the environmental impacts of the Project
Stockpiled low-grade material, which has a higher potential for metal leaching, will either be
processed in the plant or capped with overburden soils. Water from the tailings pond will be
pumped to the open pit to flood the pit walls. Any other materials deemed to be PAG or have high
neutral leaching potential would also be submerged in the open pit.
Seepage water quality from the waste dumps will be monitored and, if required, measures taken
to mitigate the potential impact on water quality in the tailings facility. These measures include
pumping seepage to the open pit for storage and/or capping exposed dump surfaces to reduce
infiltration.
An annual net surplus of about 3.7 Mm3 of water will report to the tailings facility during closure,
resulting in a 2 m to 33 m pond water level rise without release. The height of the pond water
December, 2007
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level rise depends on the size of the impoundment at the time of closure. For the first year of
closure, sufficient water will be pumped to the open pit to maintain a constant pond level without
discharge. If closure extends beyond 1 year, the closure spillway will be installed to route the
PMF from the tailings facility. Constant year-round release of water through the spillway will
then occur.
The diversion channels will be maintained during temporary closure periods. The diversion flows
will mix with the spillway flows to achieve an acceptable water quality in Ruby Creek.
18.3
Environmental Considerations
The environmental component of this feasibility was conducted by Klohn Crippen. The following
is an excerpt from the February 2006 Ruby Creek Molybdenum Project Executive Summary,
Environmental Assessment Certificate Application:
A two-stage approach was used to determine the environmental and socioeconomic effects of the
project. First, socioeconomic and environmental components were screened to rank their
importance. Components were ranked on whether they could interact with the project as well as
whether they are important to the Taku River Tlingit First Nation (TRTFN), regulators and other
stakeholders. The results of this screening generate Valued Environmental Components (VECs)
and Valued Socioeconomic Components (VSCs). These were defined as:
•
Fish and fish habitat
•
Moose
•
Game birds
•
Woodland Caribou
•
Grizzly bear
•
Receiving water quality
•
Stone’s sheep
•
Hoary marmot
Fish and fish habitat assessments were based on the removal of fish habitat and fish populations
as a result of the direct location of the footprint of the various project components, or the
reduction in flows in the rivers downstream of the project.
The baseline work identified an isolated population of 200 - 400 Arctic grayling that will be
moved as part of the environmental management for the fish habitat. The project will also
provide a no-net-loss fish habitat compensation plan for the loss of habitat in the Ruby Creek
drainage. The fish habitat compensation plan replaces lost spawning habitat on the lower parts of
creeks affected by placer mining where this habitat is no longer available to fish.
To protect the animals in the area, there will be a no-hunting policy for employees accommodated
at the project site, and access to and within the mining lease will be restricted.
December, 2007
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Woodland caribou habitat will be directly affected by project construction and indirectly affected
by project noise. Personnel of operating mines in the Northwest Territories report that caribou
become accustomed to mining activity and even use the mine roads to escape biting insects. The
mine will disturb some calving habitat; however, it is felt that this will not limit the stability of
the caribou population.
There may be some direct removal of Stone's sheep habitat, but regionally this is a small amount
relative to the total potential range of the sheep.
Moose generally use the lower section of Ruby Creek, with some use of the wetland in the lower
part of the project footprint. In general, high quality moose habitat is not disturbed by the project.
Moose also become used to project activities and experience at other mines shows that moose
migrate into the mine area no-hunting zone during the hunting season.
Grizzly bear ranges are extremely large and land use is subject to disturbance by human activity.
The additional work being carried out in the upper Ruby Creek drainage affects only a small part
of the grizzly bear habitat.
A hoary marmot colony will be moved to accommodate the tailings disposal facility. As marmot
colonies have been successfully moved in other areas, the long-term effect on their population,
which ranges from Alaska into the continental USA, is expected to be minimal.
Game bird habitat will be affected by the project footprint, though the birds' range and the current
densities are not felt to be limiting on their populations. Therefore, the project's effect on game
birds will be the displacement of some birds to adjacent habitat areas.
Provincial regulations provide water quality guidelines for the protection of aquatic life. The
project operations plan is to discharge water from the tailings disposal facility during the freshet
when the receiving environment has excess water. The receiving water quality will reflect the
background conditions and will not be affected where Ruby Creek enters Surprise Lake. At
closure, water from the project site may have concentrations of some metals at levels that are
similar to the elevated background concentrations.
The Environmental Assessment Application was accepted for review on August 01, 2006 and was
received from the Province of British Columbia on September 10, 2007.
December, 2007
18.4
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Capital Costs
2007 Capital Cost Estimate Update
The total initial capital cost for the development of the Ruby Creek Project is estimated to be
CDN$640 million. The cost estimate has been carried out to an accuracy of +15%. The estimate
was prepared in house by professional engineers.
Adanac realized the importance of updating the capital costs to reflect current market conditions
to support the updated feasibility study due to the escalation in project costs over the past few
years. The detailed estimate produced for the 2006 feasibility study was used as a basis and
updated with current labour rates and construction materials costs ie. electrical cable, concrete
and steel. Quantities were updated where detailed engineering supported refining the quantities.
Further detail was available from additional geotechnical work and civil design and this was
included in the updated estimate. Approximately CAN$130 million has been committed to date
for process, mining and mobile equipment, and these costs have been included in the updated
capital cost estimate. Committing to the process and mining equipment at an early stage of the
project has eliminated the highest risk component in current estimates due to the volatility of
these costs.
The resulting capital cost is a true reflection of current costs.
A summary of the major costs is shown in Table 18.1. Detailed descriptions of capital costs are in
the appropriate reports.
Table 18.1
Cost Centre
Atlin Infrastructure
Property Purchase
Fencing / Infrastructure
Secure Storage (2 Containers)
Atlin Office
Atlin Permanent Housing Included in Owners
costs
Total Atlin Infrastructure:
Ruby Creek Site Infrastructure
Temporary Roads and Grading
Total Construction Cost
(CDN$)
184,069
100,000
8,000
500,000
0
792,069
1,000,000
December, 2007
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Construction Camp
Camp Transport / Mobilization
Camp Rental
Maintenance Building
Catering and Consumables
Fuel and Propane
Water Supply
Sanitary Sewer
Solid Waste Disposal
Fuel Storage
Temporary Communications
Fisheries Compensation
Wildlife Management (2008)
Total Construction Camp
Permanent Facilities Construction
Aggregate Supply and Material Processing
Batch Plant
Access Road
Bridge
Water Supply
Bulk Earthworks Contractor Mob/Demob
Seepage Recovery Dam
Diversion Ditches and Rip Rap Production
Employee Village Access Road
Tailing and Roads and Sediment Ponds
Plantsite Civil and Piles
Concrete
Structural & Arch
Mechanical and Piping
Electrical and Instrumentation
Permanent Camp
Indirects
Total Permanent Facilities:
Engineering and Construction Management
ADANAC Consultants
All North Road Design and CM
AMEC Engineering and Procurement
Golder Mine Engineering
Klohn Crippen Tailings
Dam/Environmental/Geotechnical
Construction Management
Vendor Representatives
Expediting and Inspection
Ledcor Pre Construction Planning
Total Engineering and Construction:
7,750,000
6,927,000
300,000
13,671,000
8,937,500
100,000
50,000
108,500
220,003
310,000
400,000
120,000
38,894,003
10,825,000
1,900,000
3,100,000
500,000
2,000,000
12,000,000
14,836,000
2,500,000
24,950,959
27,342,349
31,853,901
23,035,562
55,641,467
30,745,974
14,270,161
60,489,088
316,270,422
600,000
1,224,859
19,956,666
205,674
4,584,963
14,925,000
3,250,000
3,000,000
1,001,629
48,748,790
December, 2007
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Procurement
Gyratory Crusher
Apron Feeders
Rock Breaker
Overhead Cranes
Mobile Crane
HPGR
Cone Crusher
Cyclopacs
Ball Mill
Liner Handler
Regrind Mill
Flotation Cells
Samplers
Compressors
Agitators
Pressure Filter
Bagging System
Scrubber
Conveyors
Lime System
Dust Collectors
Reclaim Barge
Pumps
Air Handling
Thickener
Boilers
Standby Gensets
Structural Steel
Gensets
Total Procurement:
4,317,234
862,500
326,633
1,443,885
500,000
15,266,782
3,214,610
295,809
19,043,104
1,078,000
2,880,000
1,241,169
88,000
157,500
84,800
488,000
683,900
222,300
4,750,000
355,000
25,000
2,231,550
245,000
185,523
402,500
375,000
3,300,000
24,000,000
21,433,220
109,497,019
Total Construction:
515,202,323
Contingency
50,628,464
Risk
20,130,004
Total:
585,960,791
Mining Pre-Strip
24,039,207
Mining Fleet
25,000,002
Contingency
5,000,000
Total:
640,000,000
2006 Capital Cost Estimate
The capital cost estimate produced for the 2006 Feasibility Study was a high level estimate
prepared in accordance with industry standards for Feasibility Studies, to an accuracy of +/_15%..
December, 2007
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The basis of the capital cost estimate was engineering designs and quantity take offs including:
mine plans and haul profiles, process metallurgical testwork, design criteria, flowsheets and mass
balance, equipment specifications, layouts and general arrangements, simplified P&ID’s, Single
line diagrams and electrical distribution design.
Owner supplied equipment costs were obtained from quotations by suppliers. Mining Equipment
was selected based on the mine plan, haul profiles and load/haul simulation software. Process
equipment selection was completed based on extensive metallurgical testwork, simulation and
design.
Construction bulk material costing is developed from vendor quotations and bills of materials that
are taken off from piping line lists, electrical cable lists, valve lists and structural, mechanical,
piping, electrical and instrument red lines produced by the discipline enginers.
Engineered
sketches were prepared to fully define piping, structural steel and concrete building services. Due
to the high level of design for the tailings and water management completed to support permitting
efforts these components were estimated to a very high level of accuracy.
Built up, burdened labour rates were developed from input from local and regional contractors.
The 2006 capital cost estimate was a build up of over ten thousand line items of detail including
units, manpower productivity, built up labour rates, bulk, tagged (owner supplied) equipment,
construction equipment and subcontract costs and included indirect and contingencies applied by
work package.
18.5
Operating Costs
2007 Operating Cost Estimate
Operating Costs for the 2007 Feasibility Study Update were developed from first principles
taking in to account the increase in plant throughput from 900 tonnes per hour to 1100 tonnes per
hour.
The mine operating costs were developed using enhanced mine plan, equipment selection and
operating hours. Process operating costs could be defined to a very high level of accuracy as flow
sheets and equipment selection is frozen allowing reagents and consumables to be accurately
defined, especially with respect to diesel consumption, a precise estimate of electrical power
developed and fuel consumption was prepared. Consumable costs were updated from budget
quotations from Vendors.
December, 2007
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Labour costs were developed for an updated staffing and operating plan using current burdened
labour rates from current collective bargening agreements in Northern British Columbia and the
Yukon.
G&A costs were updated using current labour, camp and transportation costs based on the
increased level of detail available since the original feasibility study.
After the operating costs were updated from first principles they were benchmarked against
similar operations as a check. Comparisons were within the accuracy of the estimate. The costs
were reviewed by a reliable independent reviewer, Mr R. Pendreigh P. Eng.
Process operating supply costs are based on budgetary prices from vendors of the consumables
and reagents. The costs for general and administration (G&A) includes mine management,
transport, insurance, warehouse and security personnel and general management.
Tailings dam construction costs have been estimated from other published costs from mines using
similar construction techniques.
The total operating cost for is estimated to be CDN$11.71 for the first four years of full
production and CDN$7.90 per tonne of mill feed for the remaining years of operation.
Commissioning and pre-stripping activities completed in Year 1 have been accounted for in the
capital cost estimates. The operating costs for mining, processing, power, tailing dam operation,
general administration and Adanac’s related cost were prepared by in house professional
engineers and reviewed by independent qualified persons to confirm that the work conforms to
good engineering practice. Tables 18.2 and 18.3 present the operating cost summary.
Table 18.2
(First Five Years of Full Production)
Description
Operating Cost
(CDN$/t ore milled)
Mining (average)
Processing
Power
G&A
Owner Cost
3.94
2.40
5.57
0.26
0.10
0.81
13.08
December, 2007
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Table 18.3
(After Five Years of Full Production to End of Life [Year 21])
Description
Mining (average)
Processing
Power
G&A
Owner Cost
Operating Cost
(CDN$/ tones of ore milled)
2.47
2.40
2.25
0.26
0.10
0.63
8.11
2006 Operating Cost Estimate
Operating Costs for the 2006 Feasibility Study were completed from first principles and
developed to a high level of accuracy. Operating costs for the mining fleet were based on
equipment usage and defined from currently accepted equipment performance characteristics.
Operating costs for the process and G&A were developed from detailed reagent and consumable
usage as defined from the metallurgical testwork and based on quotations from suppliers.
Operating costs for G&A were developed from first principles with manpower loading and
current burdened labour.
The mine has been considered to be owner-operated. The average unit mining cost was
determined to be $1.39 per tonne mined or $3.94 per tonne of ore milled for the first four years of
full production and $2.47 for the remaining years of operation.
18.6
Market
CPM Group’s report, “The Sustainability of Recent Molybdenum Prices” dated October 22,
2007, forms the basis for the molybdenum market assessment which follows below.
Structural shifts in the supply and demand of molybdenum have revived the molybdenum market
after nearly a decade of depressed fundamentals and prices. Many of the underlying themes that
drove prices sharply higher in 2004 are still present in the market. As of middle of October 2007,
the spot price for molybdenum stood close to US$30.50 per lb.
December, 2007
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18.6.1 Supply Fundamentals
Several aspects of the molybdenum market make it a unique industry that is difficult to
comprehend at first glance. Molybdenum originates from two distinct sources each of which
adhere to different market fundamentals.
Historically roughly 60%, of the world’s molybdenum supplies has come from copper deposits
with included molybdenum mineralization. In these cases, molybdenum is produced as a copper
by-product. Copper producers can exploit high grade molybdenum mineralized zones in copper
ore bodies or add molybdenum recovery circuits, if molybdenum prices are high enough to justify
the expense. This type of selective mining is not sustainable throughout the life a mine, but it can
lead to large swings in the available supply of molybdenum from year to year. As an added
complication, by-product producer’s output is largely influenced by copper prices. This has
added to the volatility in molybdenum production and price. In the future, by-product producers
could have a somewhat reduced effect on the molybdenum supply and prices, as their market
share is forecast to fall to an average 52% over the ten year projections prepared by CPM Group.
Primary molybdenum deposits compose the majority of the remaining sources of production.
Primary molybdenum producers, in contrast to copper producers, have historically acted as swing
producers because, in the past, operating costs per pound of molybdenum have been lower for byproduct producers than for primary producers. Depending on the price of molybdenum, even
large-scaled primary producers have brought their operations on and off line. This has shifted in
recent years, as more recently, it is the smaller by-product and primary producers which are better
characterized as swing producers due to higher operating costs. The increased number of primary
producers responding to the higher demand for molybdenum has helped to stabilize prices.
Increased demand for molybdenum, notably over the past four years, has stimulated a substantial
restructuring of the molybdenum industry. Operating costs for both primary and by-product
producers have also risen as miners seek to develop deposits that were once not economically
viable. Practically all molybdenum producers are now facing operating costs that are greater than
can be supported by the long-term average price for molybdenum of US$3.80 per lb. Based on
forecasts of ore grades mined and higher actual unit costs production costs, cash costs for
molybdenum production are expected to rise in the direction of US$12.00 per lb for some primary
and by-product producers. The growth in demand and the resulting increase in molybdenum, is
expected to offset these rising costs, however.
18.6.2 Supply Outlook
Under CPM Group’s Base Case scenario over the next three years, a deficit in the molybdenum
market is predicted. Years of low investment levels in exploration and development have
December, 2007
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contributed to an increasing number of bottlenecks throughout the molybdenum mining industry,
which has led to a lag in bringing new mines on stream. Reportedly, new equipment orders are
facing lead times as long as three years, as opposed to the normal six-month delivery period. In
addition, output from China has been a growing source of uncertainty in the global molybdenum
market, as China uses more of its molybdenum itself.
The Chinese government has been tightening regulation and control over not only molybdenum
production and trade, but other mineral resources as well. In late December 2006 the Chinese
government issued a revised version of its “Catalogue of Projects for which Land Use if
Prohibited.” Molybdenum mining and smelting projects fall on this list. Thus approvals of new
operations or expansions at existing operations for molybdenum mining or smelter projects is
prohibited, except for projects that are upgrading to meet government standards. Present policy
makes it impossible, or at the least very difficult, for any sizable new mine or expansion to obtain
the necessary mining permits to come online. Some cities may choose to ignore the national
policy and issue new permits. However, these would be smaller projects that could slip under the
government’s radar. China’s production of high quality finished steel products and its domestic
consumption provide strong internal demand for molybdenum. Given recent policy changes, it is
unreasonable for the market to expect excess output from Chinese producers to meet the shortfall
of western world production.
Fresh output from primary producers, which are expected to play a large role in filling the gap
between mine production and demand, will begin to trickle into the market beginning in 2009.
As for by-product production, the continued use of the leach-solvent extraction-electrowinning
(SX/EW) process will limit the potential molybdenum recovery from some new copper projects,
as the SX/EW process does not produce any molybdenum as a by-product. Considering the
demand projections outlined in CPM Group’s base case scenario, a shortfall in mine production is
forecast to remain until 2011. A ramp up at several large-scaled operations and additional
supplies from secondary output should move the market into a market surplus in 2011. A
moderate supply surplus, averaging less than 20 million pounds per year or a little over 3% of
globally demand, is forecast from 2011 through 2016.
At this preliminary stage of planning it is difficult to gauge whether or not some of the expected
molybdenum projects, which are forecast to contribute to the supply surplus in the later years of
the projection period, will be able to overcome the obstacles inhibiting these projects from
coming on stream. Permitting, economic, technical, and political hurdles plague a few of the
proposed projects. If these issues are not properly addressed, these projects could face multiple
setbacks and cost overruns, or may even be slashed before the mine is ever brought to production.
December, 2007
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18.6.3 Demand Fundamentals
Molybdenum is primarily used as an alloying agent in a wide range of steels and alloys. The
grayish non-toxic metal is employed in various steels, including many stainless steel grades
because of its durability, strength, and robust qualities. Molybdenum alloys are resistant to
extremely high temperatures as molybdenum has both a very low thermal expansion coefficient
and one of the highest melting points of all elements. These qualities in conjunction with
molybdenum’s other properties limit consumers’ ability to substitute for other metals in its
numerous applications. Demand has been not only growing in its principal end uses, but demand
for molybdenum has been evolving as many industries have sought to develop new materials that
benefit from its alloying properties.
18.6.4 Molybdenum in Steel
Steel is an alloy of iron and carbon. Other alloying elements such as chromium, nickel,
manganese, tungsten, vanadium, cobalt, silicon and molybdenum can be utilized in order to meet
desired properties. Steel types can be broadly classified into three categories: Carbon steels, lowalloy steels, and high-alloy steels. Within each of these categories there are different types of
steels, which come in various grades. This report looks at individual steel types. These include
stainless steel, full alloy steel, carbon steel, tool and high speed steel, and high strength low alloy
(HSLA) steel. Demand for molybdenum by the steel industry totalled 293 million pounds, or
69.9% of total demand in 2006.
Stainless steel is an iron - carbon alloy with a minimum of 10.5% chromium content.
Molybdenum is used as an alloying element in stainless steels to enhance heat resistance,
strength, malleability, corrosion resistance and to achieve thinness. Production of stainless steel
is projected in the base case scenario to grow at a CAGR of 5.5% though 2016, after growing at a
CAGR of 6.0% from 1995 to 2006. Demand for molybdenum by the stainless steel industry is
projected to grow at the same pace as stainless steel production. In 2006, stainless steel
accounted for 27% of molybdenum demand.
Full alloy steels are difficult to classify. They include structural and alloy engineering steel,
Hatfield steel, maraging steel, carburizing steel, rail steel, spring steel, nitriding steel, bearing
steel, pressure vessel steel, casting steel, weathering steel, and high strength steels. The
molybdenum content of full alloy steels varies widely and is dependent on the steel’s end use.
Some room for substitution from molybdenum to other alloying elements does exist. Demand for
molybdenum is expected to grow at a slower rate than global GDP and is projected to grow at a
3.9% CAGR through 2016.
December, 2007
- 100 -
Molybdenum is used in tool steels to provide resistance to cracking, hardness, toughness,
machinability and corrosion resistance. The amount of molybdenum in tool steel by weight can
vary from 3% to 8% by weight, with higher molybdenum content steels produced primarily in the
United States, European Union and Japan. Demand for tool steel tends to follow global economic
conditions. As such, demand for molybdenum in tool steel could grow roughly at the same pace
projected for global GDP growth, 4.8% per annum.
HSLA steel is used by the transportation, construction and energy industries in a wide range of
applications from oil pipelines to automotive engine supports and bridge construction. Pipeline
construction growth is expected to remain robust throughout the projection period, growing at
approximately 4.0% per annum. This growth is projected to be led by Russia, China and India,
with significant expansions in the United States as well. Molybdenum use in HSLA steels is
projected to grow at a CAGR of 4.6% as a result of strong expected growth in the primary end
uses of HSLA steel.
In carbon steels, carbon is the primary alloying element, adding strength and hardness to iron.
Molybdenum is not used intensively in carbon steel production. When it is used, it enhances
strength and hardness in addition to providing heat and corrosion resistance. Demand for
molybdenum in the carbon steel industry is expected to grow at a 5.0% per annum throughout the
projection period. This reflects strong demand for carbon steel, led by the emerging markets of
China and India.
18.6.5 Other Molybdenum Applications
Molybdenum is used substantially in the petroleum refining and plastic industries. It is used in
various steps of the refining process, especially hydroprocessing – the removal of sulfur, nitrogen,
and other impurities from crude oil with the aid of hydrogen. The results of these processes are
energy products such as gasoline, jet fuel, diesel, kerosene, and others. The forecasted 7% CAGR
through 2016 for molybdenum demand in the catalyst industry is driven by the following factors:
the construction of new refineries required to meet increased demand for energy products by
emerging economies, the increased use of heavier crudes, environmental regulations, and
molybdenum’s use in the direct liquefaction process in transforming coal to liquid energy
products. Molybdenum’s use in the catalyst sector is expected to be strong.
A superalloy is an alloy that has high temperature qualities, high creep strength, and oxidation
resistance. The term was first coined after the Second World War to describe alloys used in super
chargers and aircraft engines. Although superalloys are still dominantly used in their traditional
markets, usage has expanded to gas turbines, chemical plants, and petroleum plants. Superalloys
are estimated to comprise 50% of an aircraft engine’s weight. They are usually made to order and
are primarily composed of nickel, an estimated 80% of superalloy content. CPM Group’s
December, 2007
- 101 -
forecast of 5.5% CAGR through 2016 is heavily influenced by activities in the nickel and airline
industry. The aircraft manufacturing industry is expected to grow at 5.5% through 2025.
Molybdenum use as a pure metal or molybdenum-based alloy is dominated by its use in the
lighting, coatings, and glass industries. It is often alloyed with nickel and chromium to produce
highly resistant coatings. These coatings improve the wear and friction properties of automotive
parts such as gears, synchronizers, and piston rings. In lighting, its oldest application as a metal,
molybdenum’s elevated temperatures strength, creep resistance and chemical compatibility with
glass provide it with a significant advantage over substitutes. Recent price increases have caused
substitution to nickel and titanium in emerging applications such as liquid crystal displays.
Molybdenum demand in this sector is expected to grow at not more than 3% per annum through
2016.
Cast iron, also known as an iron-carbon-silicon alloy, refers to a family of multi-component
ferrous alloys containing primarily iron, carbon, silicon, as well as other major and minor
alloying elements. Molybdenum increases cast iron’s strength, heat resistance, and creep
resistance. Due to molybdenum’s sparing use in fabrication, its demand here is relatively
inelastic. Molybdenum demand growth in this sector is expected to almost match the pace of
global economic growth. Cast iron growth is forecasted at 4.6% through 2016, driven by growth
in China, India, and Asia-Pacific.
Molybdenum as a disulfide is used in lubricants due to a lower coefficient of friction than other
lubricants, durability to withstand high temperatures, high yield strength, and a strong affinity for
metallic surfaces. The global forecast for lubricant demand growth is 2.3% per annum through
2016, dominated by growth in lubricant use as engine oils as a result of growing manufacturing
and automotive markets in India, China, and Russia.
Molybdenum is used in molybdate base pigments because of its stable color formation and
corrosion inhibition properties. Molybdate pigments could be bright colors based on
molybdenum oranges prepared by co-precipitating lead chromate, lead molybdate, and lead
sulfate. These light and heat-stable pigments have a wide range of colors from bright red-orange
to red-yellow, and they are used in paints, inks, plastics, rubber products, and ceramics. The use
of lead and the toxicity of molybdenum present challenges in this market segment. Molybdenum
demand in the pigment industry is expected to grow at 3% CAGR through 2016.
Molybdenum’s use in chemicals and other applications is dominated by its use as a smoke
suppressant in PVC cabling. Molybdenum stabilizes char in combustible situations and thus
prevents the formation of smoke particles. PVC is made through the polymerization of a vinyl
choride monomer. PVC is used in construction, housing, automobiles, airplanes, and medical
December, 2007
- 102 -
devices. Molybdenum demand growth is expected to be 4.0% – 4.8% through the projected
period, closely following GDP growth and influenced by the housing and construction industries.
18.6.6 Substitutes
Possible substitutes for molybdenum include chromium, tantalum, vanadium, columbium
(niobium), tungsten and boron. The metals most frequently used as molybdenum substitutes in
steel are columbium and vanadium. Substitution away from molybdenum typically comes at an
economic or performance loss in most situations. The last published price for Brazilian
polychlore, the columbium ore mineral, is from 1981 and in 2006 realized costs per pound were
around US$45 per pound. Approximately 82% of mine production is from Brazil. Substitution
from molybdenum to vanadium typically results in a lower melting point and lower corrosion
resistance. In many applications the impulse may be to substitute other metals with molybdenum,
despite higher molybdenum prices, due to increases in the prices of these metals as well. Cases in
which there is little room for substitution to molybdenum include the use of chromium in
stainless steel or tantalum in capacitors.
18.6.7 End-Use Industry Analysis
Approximately 38% of molybdenum demand comes from the energy industry. Molybdenum is
used extensively in nuclear power plants, coal to liquids facilities, and petroleum refining,
pipelines, and drilling equipment. The end uses of molybdenum that relate to energy include high
strength low alloy steel, stainless steel, carbon steel, tool steel, full alloy steel, and catalysts.
Worldwide energy demand has grown at a 2.6% per annum rate between 2001 and 2006, double
the 1.3% per annum growth rate in energy requirements in the previous five year period.
Molybdenum demand grew at a CAGR of 7.1% per annum from 2001 to 2006, after growing at a
CAGR of 3.7% per annum in the previous five year period. Demand for molybdenum by the
energy industry is expected to remain robust, accounting for 38% to 40% of total demand
throughout the projection period.
Residential and commercial construction each utilize molybdenum. In the United States growth
in the construction industry has traditionally been led by residential construction, which has been
trending lower since May 2006. This decline has been offset in part by growth in the commercial
construction sector, however. The value of the construction industry exceeded US$4 trillion in
2004, with significantly higher costs in the European Union and the United States. Spending is
the most accurate way to gauge the pace of construction growth globally. Engineering News
reported costs of US$550.00 per square meter in New York compared to US$70.00 per square
meter in Shanghai. China, India and other emerging markets with lower construction costs are
projected to drive the construction industry and sustain demand for molybdenum in the
construction industry.
December, 2007
- 103 -
Molybdenum’s use in the transportation industry is a testament to its versatility as a metal and its
role in contemporary industrial technology. It is used in all modes of transportation whether by
land, air, sea, or rail. In the railway sector, it is used in wheel seats, brake pads, and locomotives
engines. Molybdenum’s use in the aerospace sector is due to its heat and creep resistance
properties in constructing airplane engines. These engines are made with superalloys, of which
molybdenum is an important component. Shipbuilders use molybdenum in duplex steels due to
its corrosion resistance properties, enabling them to transport a wide variety of materials on sea.
Finally in automotives, molybdenum is used in lighting caps, turbocharger housings, automotive
engines, automobile bodies, and suspension springs. Growth of molybdenum use in the
transportation industry is expected to almost match GDP growth at 4.6% per annum, resulting in
96.9 million pounds of molybdenum demand in 2016.
18.6.8 Price Outlook
CPM Group forecasts the price of molybdenum to remain strong over the next ten years, above
previous historical averages, due to tight demand and supply conditions covered in CPM Group’s
Base Case scenario. Prices are unlikely to revert to their levels in the 1990s as operating costs
for both primary and by-product producers have risen sharply. Prices may reach US$34.00 per lb
in 2008, due to the predicted shortage in supply. However, prices are expected to hold above
US$28.00 per lb through 2010. By 2016 prices may fall further. This is demonstrated in
Table 18.4.
December, 2007
- 104 -
Table 18.4
Year
Price (US$/lb Mo)
2009
32.25
2010
28.00
2011
23.00
2012
21.75
2013
19.50
2014
16.00
2015
15.00
2016
14.75
Given the potential for the molybdenum supply outlook to be bleaker than projected in
CPM Group ’s main scenario, toward the end of the projection period, prices could continue to be
supported at higher levels. This is presented graphically in Figure 18-1.
Figure 18-1
Real Molybdenum Prices and Word Supply and Demand Balance
CPM Group Base Case: Real Molybdenum Prices and World S upply and Demand Balance
Annual, Projected through 2016p
Million Pounds
780
$US /Lb.
42
World S upply (LHS )
36
700
World Demand (LHS )
30
620
Molybdenum Prices (RHS )
540
24
460
18
380
12
300
6
Actual
Projections
0
220
1995
1998
2001
2004
2007p
2010p
2013p
2016p
December, 2007
18.7
- 105 -
Economic Model
CPM Group prepared an Economic Model for the Project based on the following assumptions for
the base case:
•
Project construction starts in 2007 with commissioning in early 2009;
•
Production commences in the second quarter of 2009;
•
A US$/CDN$ exchange rate based on a consensus view of the leading CDN$ market
makers and their outlook for the US$/CDN$ exchange rate sourced from Bloomberg;
•
A roasting charge of US$1.77/kg of Mo contained in concentrate to convert MoS2 to
MoO3 and a conversion loss of 1.0%;
•
Concentrate shipped to a roasting facility in North America (to be determined);
•
Declining molybdenum metal price from US$32.25 per pound in 2009 to US$14.75 per
pound in 2016 and remaining flat at that level thereafter in the base case;
•
The model was prepared on a pre-tax basis.
18.7.1 Net Present Value and Internal Rate of Return Summary
CPM Group developed a spreadsheet that could be used by Adanac to evaluate the sensitivity of
the Project to various inputs. The ability to vary these inputs allows the impact of updated field
results as well as changes to capital or operating costs, metallurgical test work results, exchange
rates or the outlook for molybdenum prices to be examined quickly and the effects on the Project
determined.
Capital, initial or sustaining, was considered in the year that it was spent. The contingency
reserve was based on firm purchase orders or quotes, or refreshed estimates from detailed
engineering. Consequently the contingency amount varies from 5% to a maximum of 15%,
depending on the confidence of the design and the design category.
Revenue was calculated using the determined grade and molybdenum price assumption, and
adjusted for marketing, transportation costs and metallurgical recovery.
Molybdenum price assumptions were based on the price outlook developed by the report “The
Sustainability of Current Molybdenum Prices” dated October 22, 2007 prepared by the
Commodity Market Research Department at CPM Group. The base case outlook for
molybdenum prices contained in that report are in the Table 18-4.
December, 2007
- 106 -
The metallurgical recovery rate of 90% was selected based on metallurgical test work completed
for the Ruby Creek Molybdenum Feasibility, April 2006, by Wardrop Engineering Inc. The
gross revenue was calculated based on the recovered molybdenum. The net revenue was
calculated from gross revenue less roasting, transportation and marketing charges.
In the base case the Project has an IRR of 18.9% and an NPV of $295 million at an 8.00%
discount rate. Payback is 3.2 years.
18.7.2 Sensitivity Analysis
Sensitivities to molybdenum prices, the US$/CDN$ exchange rate, capital and operating costs on
the IRR and NPV were considered. The most significant variable influencing the economic
model is molybdenum prices. In addition, specific economic sensitivities were also run for the
conditions described below.
3-Year Historical Average Molybdenum Price – The average monthly molybdenum price over
the past 3 years is US$28.46 per lb. Using this price flat throughout the life of the Project, yields
an IRR of 30.3% and a NPV of CDN$1.014 billion at an 8.00% discount rate.
Low Case Molybdenum Prices – The low case outlook for molybdenum prices contained in the
CPM Group report reflected lower growth in demand for molybdenum, primarily in the stainless
steel and pigments industries. Lower molybdenum prices would curtail some early stage or
small-scale, high cost molybdenum projects, categorized by CPM Group as “possible” project,
from coming on stream. This stems from the fact that lower prices would likely lead to a reduced
influx of fresh supply. The Project IRR decreased to 12.3% and the NPV decreased to CDN$121
million at an 8.00% discount rate
High Case Molybdenum Prices - The high case outlook for molybdenum prices contained in the
CPM Group report reflected higher growth in demand for molybdenum in the stainless and HSLA
steel industries, lubricants and catalysts. Higher prices would also spur some additional supply
from projects, now deemed possible. However, the supply response is muted due to delays in
bringing production on-stream. Under the high case molybdenum price scenario, Project IRR
was 24.6% and the NPV increased to CDN$444 million at an 8.00% discount rate.
Mine Operations with Grid Power Starting Earlier (From Year 3 Onwards) - The base case
considers a conversion from diesel-fired electricity generation at the Project site over to accessing
grid power in Year 5. This sensitivity considered introducing grid power to the Project 2 years’
earlier than in the base case, in Year 3. The Project IRR increased to 19.5% and the NPV
increases to CDN$325 million at an 8.00% discount rate, from the base case.
December, 2007
- 107 -
Increase in In-Situ Grade by 15% - A sensitivity was carried out to determine the effect of a
potential 15% increase in ore head grade. The case for a potential increase in ore grade can be
made based on detailed comparative drill hole/bulk sample studies in the “Feasibility Study of the
Adanac Molybdenum Project” prepared by Kaiser Engineers for Kerr Addison Mines Limited,
Report 71-1, dated January 1971. While this work is not NI 43-101 compliant, the results from
processing an approximate 10,000 tonne bulk sample at a 100 TPD pilot plant on site indicated
that realized grades were higher than those predicted by drilling results by as much as 20%. For
the purpose of a sensitivity study only a grade of increase of 15% was assumed. With the
increase in grade in this scenario, the Project IRR increases to 27.6% and the NPV increases to
CDN$ 533 million at an 8.00% discount rate.
18.7.3 Summary of Results
A summary of the results of the sensitivity analyses is presented in Table 18.5.
Table 18.5
Case Description
IRR
NPV @ 8%
(CDN$
millions)
Payback
Period
(Years)
18.9%
295.0
3.2
Historical Average Mo Price (Last 3 years’)
30.3%
1,014.7
2.9
Low Case Mo Price Scenario
12.3%
120.7
5.9
High Case Mo Price Scenario
24.6%
444.1
2.6
Capital Cost +15%
14.8%
213.0
3.8
Capital Cost -15%
24.8%
377.2
2.6
Operating Cost +15%
15.2%
185.6
3.5
Operating Cost -15%
22.4%
404.5
2.9
Specific Economic Sensitivities
Mine operation with hydroelectric power starting
inj year 3
19.5%
325.0
2.6
Increase in in-situ grade by 15%
27.6%
533.4
2.4
Base Case
Sensitivities
Long Term Mo Price 2016 onward
$ US / lb
NPV @ 8%
(CDN$
millions)
$10.75
156.7
$12.75
225.9
$16.75
364.3
$18.75
433.4
December, 2007
Long Term Mo Price 2016 Onward
Long Term Mo Price 2016 Onward
- 108 -
$ US / lb
NPV @ 6%
(CDN
Millions)
$10.75
222.9
$12.75
315.2
$14.75
407.5
$16.75
499.8
$18.75
$ US / lb
592.1
NPV @
10% (CDN
Millions)
$10.75
104.2
$12.75
156.6
$14.75
209.0
$16.75
261.4
$18.75
313.8
Given that molybdenum prices are quoted in US$ and hence Project revenue, and operating costs
are largely denominated in CDN$, the US$/CDN$ exchange rate is important. Indeed recent
strength of the CDN$ has benefited Adanac by lowering the cost of long-lead capital equipment
procured to date for the Project.
Full details of the Economic Model are contained as Appendix A.
December, 2007
19.0
- 109 -
INTERPRETATIONS AND CONCLUSIONS
The Project contains a valuable molybdenum-bearing mineral resource that can be economically
extracted using proven mining methods and processing technologies, at current labour, equipment
and material costs and also based on the projected prices of molybdenum in the future.
20.0
RECOMMENDATIONS
On September 19, 2007 Adanac announced that it would proceed with the development and
construction of the Project and that it is proceeding to arrange the equity and debt financing to
build and operate the mine. Adanac has received an Environmental Assessment Certificate on
September 10, 2007 from the province of British Columbia. This Feasibility Study reaffirms the
economic viability and financial sustainability of the Project.
It is recommended that Adanac continue to develop the Project through detailed engineering and
construction.
December, 2007
21.0
- 110 -
REFERENCES
Adams Metals Ltd., Molybdenum Market Study, July7, 2005.
Adanac
Molydenum Corp, Ruby Creek Molybdenum Project Executive
Environmental Assessment Certificate Application, February 2006.
Summary,
BC Mining Research Ltd., Preliminary Evaluation of Ultra-Fine Grinding for the Ruby Creek
Molybdenum Project, February 10, 2005.
Blower, S. (2005): Technical Report – Mineral Resource Estimate Ruby Creek Molybdenum
Project, report dated April 11, 2005. (AMEC Americas Limited).
Chapman, Wood & Griswold Ltd., 1971. Feasibility Study Kerr Addison Mines Limited Adanac
Project (Volume II of II – Engineering and Economic Detail).
G&T Metallurgical Services Ltd., An Assessment of Metallurgical Response – Ruby Creek
Project, December 27, 2006.
Golder Associates Ltd., Pit Slope Stability Considerations for the Ruby Creek Project, Adanac
Moly Corp, Atlin, BC, February 10, 2006.
Golder Associates Ltd., Ruby Creek Molybdenum Project Mining Feasibility Study, March,
2006.
Golder Associates Ltd., Technical Report — Mineral Resource Estimate, Ruby Creek
Molybdenum Project, February 17, 2006.
International Molybdenum plc website: www.internationalmolybdenum.com
Janes, R.H. (1971): The Geology of the Ruby Creek Molybdenum Deposit; in Chapman, Wood
and Griswold, Economic Feasibility Study, Volume VII, pp 1-14 (unpublished).
Klohn Crippen Consultants Ltd., Ruby Creek Project — Feasibility Design of Tailings Facility,
Waste Dumps and Site Water Management, February 8, 2006.
Monger, J.W.H. (1975): Upper Paleozoic Rocks of the Atlin Terrane, Northwestern British
Columbia and South Central Yukon; Geological Survey of Canada, Paper 74 – 47,
pp 1-63.
December, 2007
- 111 -
Neil S. Seldon & Associates Ltd., Marketing and Commercial Input for the Ruby Creek
Molybdenum Project, August 2005.
Palmer, P., (2006): Technical Report Mineral Resource Estimate Ruby Creek Molybdenum
Project, report dated February 17, 2006, Golder Associates Ltd.
Pinsent, R.H. (1980): Diamond Drilling Report on the Adanac Property, Adera 1, 4-8, Hobo 8,
19-20, 47 and Key 27 Mineral Claims, Atlin Mining Division; British Columbia Ministry
of Energy, Mines and Petroleum Resources, Assessment Report #8861, 3 pages plus
appendices.
Pinsent, R.H. (2005). Diamond Drilling Report on the Adanac (Ruby Creek) Property, British
Columbia Ministry of Energy, Mines and Petroleum Resources, Assessment Report,
dated January 31, 2005.
Pinsent, R.H. and Christopher, P.A. (1995): Adanac (Ruby Creek) Molybdenum Deposit,
Northwestern British Columbia; Canadian Institute of Mining and Metallurgy Special
Volume No. 46, pp 712-717.
SGS-MinnovEX Technologies Inc., Flotation Testwork Report for Plant Design, January 2005.
SGS-MinnovEX Technologies Inc., QEM SCAN Investigation of ADA NA C Molybdenum
Products 0510-AD-547, December 19, 2005.
SGS-MinnovEX Technologies Inc., Ruby Creek Semi-autogenous Grinding Circuit Design, June
10, 2005.
Sinclair, A.J. (2005) Quality Control of the 2004 Adanac Moly Corp. Drilling Program, Ruby
Creek Deposit. Unpublished consultant’s report to Adanac Moly Corp.
Sinclair, W.D. (1995): Porphyry Mo (Low-F-type), in Selected British Columbia Mineral Deposit
Profiles, Volume 1 - Metallics and Coal, Lefebvre, D.V. and Ray, G.E., Editors; British
Columbia Ministry of Energy of Employment and Investment, Open File 1995-20, pp
93-96.
Sutherland Brown, A. (1970): Adera, in Geology, Exploration and Mining in British Columbia,
1969; British Columbia Ministry of Energy, Mines and Petroleum Resources, pp 29-35.
Tennant, S. (1979): Adanac Drill Programme; British Columbia Ministry of Energy, Mines and
Petroleum Resources, Assessment Report #7727, 5 pages plus appendices.
December, 2007
- 112 -
Wardrop Engineering Inc., Ruby Creek Feasibility, Process and Infrastructure Design and Cost
Estimate, March 2006.
White, W.H., Stewart, D.R. and Ganster, M.W. (1976): Adanac (Ruby Creek) in Porphyry
Deposits of the Canadian Cordillera, Edited by A. Sutherland Brown; Canadian Institute
of Mining and Metallurgy Special Volume No. 15, pp 476-483.
December, 2007
22.0
- 113 -
DATE AND SIGNATURE PAGE
The Technical Report titled “Ruby Creek Project, Feasibility Study Update” was prepared and
signed by the following authors dated December ●, 2007
Original signed by:
Rick Alexander, P.Eng
APPENDIX A
ECONOMIC MODEL
December, 2007
- A1 -
APPENDIX B
CERTIFICATE OF QUALIFICATION
•
Rick Alexander Certificate

Adanac Molybdenum Corporation #200-2055

Transcription

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