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Lassonde Mineral Engineering Program
University of Toronto
Capstone Final Report: MIN467
Submitted to: David Eden
From: Giancarlo Volpe, Pearl Barrett, Tsun Yu Lam, Faraz Chattha
Date: Thursday April 7, 2015
Subject: Grum Project - Faro
2
Executive Summary
Northwestern Canada is home to the Grum Deposit, located in central Yukon. Approximately 200 km
north of its capital, Whitehorse, the deposit makes up one of 7 deposits in the 35 kilometer long Anvil
Range. In previous work a preliminary pit design was constructed using basic economic assumptions.
This was complemented through a detailed investigation of the geotechnical properties of the rock
which were used to assess the stability of the pit slopes.
At this stage of the design, more realistic parameters, including costs and a detailed ramp design, have
allowed for the construction of a detailed pit design. The ramp was based on a Cat 785C haulage truck,
selected at this stage of design, with a grade of 10%. Switchbacks weren’t incorporated to promote
safety and prevent significant changes in the pit economics.
A detailed preliminary design of the site’s waste rock dump (WRD) and tailings storage facility (TSF) have
been constructed. Both designs have assumed a conservative slope geometry and knowing this, a
numerical model was developed to design both facilities. The acidic properties of the waste and slurry
material draw concerns for the possibility of acid mine drainage (AMD). A wet cover on the TSF was
therefore decided to limit this generation in the generally humid climate of the Faro area.
Additionally, a basic water balance was conducted for both waste facilities. The results suggest the
tailings facility may require additional pumping to provide adequate water for the wet cover.
Consequently, the water balance also suggests the possibility of further optimization to the TSF design.
Leading to the start of production, Benny Resource Group (BRG) will obtain all required permits, licenses
and approvals. The primary stakeholders consist of the Faro community and the Kaska people, both
affected environmental changes. As such, a preliminary Impact Benefit Agreement is also included to
outline the positive impacts the project may have on the community, while a risk matrix was used to
assess various negative impacts. It is important for BRG to prevent another Faro Mine disaster and
foster mutual respect with the communities. The site layout is designed to reflect such considerations
BRG will implement progressive reclamation and obtain all permits required for mine closure, in
compliance with the government of Yukon. Furthermore, consultation with First Nations and community
stakeholders on all phases of mine closure will be essential. The main environmental concern for closure
will be the occurrence of AMD, and as a result engineered covers will be employed on the WRD, a water
cover for the TSF, while the pit will be flooded to limit AMD. The estimated reclamation cost is between
$7 and $15 Million.
The current economic study of the design suggests a Net Present Value of $156.1 Million is attainable
with a 4.8 year payback period, a mine life of 20 years, and 2 additional years of pre-stripping.
Additionally specific smelters have been considered to begin a preliminary look into appropriate metal
markets, and the associated costs have been weighed. The current state of this study suggests that the
project should be brought to the next stage. In this case, baselines studies, further site investigation and
detailed metallurgical testing should be considered as next steps.
3
Signatures of Authors
The following signatures verify the group of graduating personnel known as “Benny Resource
Group,” have written and reviewed the contents of this document.
Pearl Barrett
Giancarlo Volpe
Faraz Chattha
Tsun Yu Lam
4
Table of Contents
Executive Summary ................................................................................................................................................................ 2
Signatures of Authors ............................................................................................................................................................ 3
1 Background .................................................................................................................................................................... 13
2 Previous Analyses ........................................................................................................................................................ 13
2.1 Rock Mass Properties ....................................................................................................................................... 13
2.2 Geotechnical Domains ...................................................................................................................................... 14
2.3 Slope Stability Analysis .................................................................................................................................... 15
3 Detailed Pit Design ...................................................................................................................................................... 16
3.1 Ramp Design ........................................................................................................................................................ 16
3.1.1 Ramp Width ................................................................................................................................................ 17
3.1.2 Ramp Section Design ............................................................................................................................... 18
3.1.3 Ramp Maintenance .................................................................................................................................. 19
3.2 Pit Slope Geometry ............................................................................................................................................ 19
4 Production Scheduling ............................................................................................................................................... 20
5 Preliminary Processing Design .............................................................................................................................. 24
6 Tailings Storage Facility Design ............................................................................................................................. 25
6.1 Selection of an Appropriate Cover System .............................................................................................. 25
6.2 Design of the Dam Geometry ......................................................................................................................... 26
6.3 Considerations for Dam Construction ....................................................................................................... 28
7 Design of the Waste Rock Dump ............................................................................................................................ 29
8 Site Layout ...................................................................................................................................................................... 31
8.1 Background ........................................................................................................................................................... 31
8.2 Placement Methodology .................................................................................................................................. 31
8.3 Tailings Storage Facility .................................................................................................................................. 32
8.4 Additional Site Requirements ....................................................................................................................... 34
8.4.1 Processing Mill ........................................................................................................................................... 34
8.4.2 Explosives Storage and Handling ....................................................................................................... 34
8.4.3 Technical Departments .......................................................................................................................... 34
5
8.4.4 Environmental Systems ......................................................................................................................... 35
9 Water Balance of the Mine Site .............................................................................................................................. 35
9.1 Water Balance of the Waste Rock Dump .................................................................................................. 35
9.2 Water Balance of the Tailings Storage Facility ...................................................................................... 37
10 Operations Planning .............................................................................................................................................. 39
10.1 Equipment Selection and Pricing Model ................................................................................................... 39
10.1.1 Daily Ore and Waste Production ........................................................................................................ 40
10.1.2 Daily Productive Hours .......................................................................................................................... 41
10.1.3 Required Hourly Production Rate ..................................................................................................... 41
10.1.4 Potential Truck and Shovel Models ................................................................................................... 42
10.1.5 Properties of Trucking Routes ............................................................................................................ 42
10.1.6 Time Spent on Travelling to and from Dump and Mill .............................................................. 43
10.1.7 Loading Time .............................................................................................................................................. 43
10.1.8 Truck Cycle Time ...................................................................................................................................... 44
10.1.9 Number of Required Shovels ............................................................................................................... 44
10.1.10 Additional Equipment and Support Fleet .................................................................................. 46
10.2 Benchmarking...................................................................................................................................................... 46
10.2.1 ARCTIC (NovaCopper Inc.) ................................................................................................................... 46
10.3 Meadowbank (Agnico-Eagle Mines Ltd.) .................................................................................................. 47
11 Environmental and Social Impact Assessment ........................................................................................... 47
11.1 Required Legal Documents ............................................................................................................................ 47
11.2 Valued Ecosystem Components ................................................................................................................... 48
11.2.1 Atmospheric Systems.............................................................................................................................. 48
11.2.2 Water Systems ........................................................................................................................................... 49
11.2.3 Terrestrial Environment ....................................................................................................................... 50
11.2.4 Natural Heritage System ........................................................................................................................ 51
11.2.5 Socio-Economic Factors ......................................................................................................................... 51
11.3 Assessment of Impacts ..................................................................................................................................... 52
11.4 Impact Benefit Agreement .............................................................................................................................. 53
12 Mine Closure ............................................................................................................................................................. 54
12.1 Introduction ......................................................................................................................................................... 54
6
12.2 Regulatory Requirements ............................................................................................................................... 55
12.2.1 Permits .......................................................................................................................................................... 55
12.3 Environmental Studies ..................................................................................................................................... 55
12.3.1 Environmental Baseline Studies ........................................................................................................ 56
12.4 Objectives and Environmental Issues ........................................................................................................ 56
12.4.1 Acid Mine Generation ............................................................................................................................. 57
12.5 Environmental Management ......................................................................................................................... 57
12.5.1 Waste Rock Dump .................................................................................................................................... 57
12.5.2 Tailings Dam ............................................................................................................................................... 58
12.5.3 Pit Lake ......................................................................................................................................................... 58
12.6 Site Monitoring .................................................................................................................................................... 59
12.6.1 Water ............................................................................................................................................................. 59
12.6.2 Air .................................................................................................................................................................... 59
12.6.3 Acid Mine Drainage .................................................................................................................................. 59
12.7 Community Relations ....................................................................................................................................... 60
12.8 Closure Costs ........................................................................................................................................................ 60
13 Detailed Economic Analysis ................................................................................................................................ 61
13.1 Revenues: $6,163,000,000 ............................................................................................................................. 64
13.1.1 Price ............................................................................................................................................................... 64
13.1.2 Variable Grades and Contained Metal over LOM ......................................................................... 65
13.1.3 Variable Rock Type, Recoveries, and Recoverable Metal over LOM ................................... 66
13.1.4 Smelter Terms ............................................................................................................................................ 67
13.2 Operating Costs: $2,902,000,000 ................................................................................................................ 67
13.2.1 Mining Operating Cost: $1,171,500,000 from $3.01/tonne mined ..................................... 67
13.2.2 Processing Cost: $741,200,000 from $14.05/tonne milled .................................................... 68
13.2.3 Freight Cost: $326,700,000 from $74.50/dmt ............................................................................. 68
13.3 Capital Cost: $534,200,000 ............................................................................................................................ 71
13.3.1 Processing Equipment Capital Cost: $108,400,000.................................................................... 71
13.3.2 Mining Equipment Capital Cost: $86,900,000 .............................................................................. 71
13.3.3 Capital Pre-strip Cost: $115,500,000 ............................................................................................... 72
13.3.4 Closure Cost: $15,000,000 .................................................................................................................... 72
13.3.5 Sustaining Capital: $145,300,000 ...................................................................................................... 72
7
13.4 Taxes: $911,700,000 at a 30% tax rate ..................................................................................................... 72
14 Conclusions & Recommendations .................................................................................................................... 72
15 References .................................................................................................................................................................. 73
16 Appendices ................................................................................................................................................................ 77
16.1 Ramp Design Considerations ........................................................................................................................ 77
16.2 Re-sloped Pit Calculations .............................................................................................................................. 79
16.3 Equipment Unit Costs ....................................................................................................................................... 80
16.4 Provided Metallurgical Recovery Data ...................................................................................................... 81
16.5 Initial Tailings Volumes ................................................................................................................................... 82
16.6 TSF Volume Calculations: Volume of a Truncated Pyramid ............................................................. 83
16.7 Summary of the Annual Rate of Rise of Tailings Deposition ............................................................ 84
16.8 Soil Classification of the Overburden Material ...................................................................................... 85
16.9 TSF Option Comparison ................................................................................................................................... 86
16.10 WRD Option Comparison ........................................................................................................................... 88
16.11 Detailed Water Balance ............................................................................................................................... 90
16.12 Suggested Water Pumping Schedule to Maintain the Water Cover .......................................... 92
16.13 Measurement of Truck Routes ................................................................................................................. 93
16.14 Rimpull and Retardation Curves ............................................................................................................. 96
16.15 Travel Times ................................................................................................................................................. 102
16.16 Loading Times.............................................................................................................................................. 103
16.17 Number of Trucks Required per Shovel............................................................................................ 105
16.18 Environmental and Social Impact Assessment .............................................................................. 108
16.19 Closure Costs ................................................................................................................................................ 121
16.20 Contained Process Metals ....................................................................................................................... 122
16.21 Mill Recoveries Used for Economics ................................................................................................... 123
16.22 Sustaining Capital ....................................................................................................................................... 124
16.23 Depreciation and Tax Calculations ...................................................................................................... 125
8
List of Tables
Table 2-1 Recommended Bench Face Angles for slopes governed by Wedge failure ................................ 16
Table 2-2 Overall pit slope safety factors for each sector, at different water saturations ........................ 16
Table 3-1 Purpose of each layer in designing a ramp ................................................................................. 19
Table 3-2 Adjusted Pit Slope Parameters ................................................................................................... 20
Table 5-1 Summary of Contained Metals before processing, Recovered Metals and Average Metal
Grades ......................................................................................................................................................... 24
Table 5-2 Results of the Preliminary Mass Balance for Froth Floatation ................................................... 25
Table 6-1 Summary of Key Parameters of the Final TSF Design ................................................................. 28
Table 6-2 Estimates for required Material needed to construct the Final TSF design ............................... 28
Table 7-1 Summary of Total Waste Rock Volume Determination with Suggested Volume Adjustment
Factors [8] ................................................................................................................................................... 30
Table 7-2 Summary of Final WRD Design Parameters ................................................................................ 30
Table 9-1 Key Coefficients used in Conducting the Mine Water Balance [9] ............................................. 35
Table 9-2 Summary of the WRD Water Balance ......................................................................................... 36
Table 9-3 Summary of the water movement contributions for water movement of each stream in the
TSF water balance ....................................................................................................................................... 38
Table 9-4 Summary of water contributions for water movement of each stream after incorporating
additional pumping ..................................................................................................................................... 39
Table 10-1 Summary of chosen loading and haulage fleet ......................................................................... 39
Table 10-2 A summary of mining rates near the end of mine life. ............................................................. 41
Table 10-3 A summary of net productive hours calculation. ...................................................................... 41
Table 10-4 The distances, grades and rolling resistances involved in the haulage routes for ore and
waste. .......................................................................................................................................................... 43
Table 10-5 Number of trucks and shovel s expected throughout the mine life ........................................ 46
Table 10-6 Number of additional and support equipment expected ........................................................ 46
Table 10-7 A comparison of preliminary equipment fleets of Grum and NovaCopper’s ARCTIC .............. 46
Table 10-8 A comparison of loading and haulage fleets between Grum and Agnico Eagle’s Meadowbank
.................................................................................................................................................................... 47
Table 11-1 Permits for various Mine Activities ........................................................................................... 48
Table 11-2 Summary of Key Impacts, Causes, and Mitigation Strategies ................................................... 52
9
Table 12-1 Permits Required for Mine Closure .......................................................................................... 55
Table 12-2 Environmental Baseline Studies ................................................................................................ 56
Table 12-3 Estimated Closure Costs ............................................................................................................ 61
Table 13-1 – Performance metrics .............................................................................................................. 61
Table 13-2 - Summary of financial results .................................................................................................. 61
Table 13-3 - The forecast prices used for the model .................................................................................. 64
Table 13-4 - The long term price forecasts and the average, consensus price from three banks .............. 65
Table 13-5 - The recoveries of each metal for each rock type.................................................................... 66
Table 13-6 - Smelter terms used, adapted from Prices and Revenues [40] ............................................... 67
Table 13-7 – The total capital costs associated with the total mining equipment fleet ............................ 71
Table 16-1 Summary of Associated Unit Costs for Selected Machinery ..................................................... 80
Table 16-2 Preliminary Recovery Data Provided for the Grum Deposit ..................................................... 81
Table 16-3 Table Showing the process in Calculating Annual Tailings Volumes......................................... 82
Table 16-4 Table Showing Summary of Tailings Rate of Rise for the final TSF design. Notice the given
Storage Length and Width used in the design. ........................................................................................... 84
Table 16-5 Summary of the Soil Classification of the Grum Overburden Material, including Key Findings
.................................................................................................................................................................... 85
Table 16-6 Economic Indicators TSF Option Comparison ........................................................................... 86
Table 16-7Environmental Indicators TSF Option Comparison .................................................................... 87
Table 16-8 Social Indicators TSF Option Comparison ................................................................................. 87
Table 16-9 Economic Indicators WRD Option Comparison ........................................................................ 88
Table 16-10 Environmental Indicators WRD Option Comparison .............................................................. 89
Table 16-11 Social Indicators WRD Option Comparison ............................................................................. 89
Table 16-12 Summary of the reported Detailed Water Balance ................................................................ 90
Table 16-13 Summary of the Recommended Pumping Schedule and resulting Water Balance (note the
negative values require pumping of water out of the dam)....................................................................... 92
Table 16-14 - Travel times for various road segments on the route of a CAT 777G ................................ 102
Table 16-15 -Travel times for various road segments on the route of a CAT 785D ................................. 102
Table 16-16 - Travel times for various road segments on the route of a CAT 789D ................................ 102
Table 16-17 -The time involved in a load, haul, dump, return cycle of a CAT 777D................................. 103
Table 16-18 - The time involved in a load, haul, dump, return cycle of a CAT 785D ................................ 104
10
Table 16-19 -The time involved in a load, haul, dump, return cycle of a CAT 789D................................. 105
Table 16-20 - The number of CAT 777G trucks required for each type of shovel .................................... 105
Table 16-21 - The number of CAT 785D trucks required for each type of shovel .................................... 106
Table 16-22 - The number of CAT 789D trucks required for each type of shovel .................................... 107
Table 16-23 Yukon Air Quality and Particulate Matter Standards ............................................................ 108
Table 16-24: Yukon water quality standards to monitor and follow, the bolded items are pertinent to the
Grum Site. ................................................................................................................................................. 110
Table 16-25: Risk assessment criteria for event severity .......................................................................... 111
Table 16-26: Risk assessment criteria for event probability ..................................................................... 112
Table 16-27: Risk Matrix ........................................................................................................................... 112
Table 16-28: Impact assessment: Pit development and mining ............................................................... 113
Table 16-29: Impact assessment: Waste rock dump ................................................................................ 116
Table 16-30: Impact assessment: Tailings storage facility ........................................................................ 116
Table 16-31: Impact assessment: Waste Management ............................................................................ 117
Table 16-32: Impact assessment: General operational ............................................................................ 118
Table 16-33: Impact assessment: Closure and remediation ..................................................................... 118
Table 16-34: Impact assessment: Natural disasters ................................................................................. 120
Table 16-35 Unit Costs of Items Needed for Closure ................................................................................ 121
Table 16-36 - The effective recoveries and recoverable metal of ore sent to the mill for each year of mine
production................................................................................................................................................. 123
Table 16-37 - The effective recoveries and recoverable metal of ore sent from the stockpile ............... 124
Table 16-38 - The calculated sustaining capital to be allotted annually over the LOM ........................... 124
Table 16-39 - Depreciation (at 20%) and tax (at 30%) calculations .......................................................... 125
List of Figures
Figure 2-1 Conservative Mohr Coulomb Criterion for Joint Strength, assuming no cohesion ................... 14
Figure 2-2 Simplified outline of the proposed Grum Pit, divided into 10 sectors with 8 unique
orientations ................................................................................................................................................. 15
Figure 2-3 Visualization of the Pit's Geotechnical Domains ....................................................................... 15
Figure 3-1 Two-Way Traffic Ramp Design ................................................................................................... 17
11
Figure 3-2 Ramp design for pushback 27. Note that the ramp exit is towards the southeast part of the
pit, making the haulage distance to WRD shorter. Thus increasing productivity. ..................................... 18
Figure 3-3 Construction Layers of the Ramp Surface ................................................................................. 19
Figure 3-4 Pit Wall Geometry for Sectors 8, 10, 9&1 .................................................................................. 20
Figure 4-1 - The production schedule needed to meet a 3.2 Mt mill capacity. The first year could be
ramped over the preceding two years, as indicated by the arrow. ............................................................ 21
Figure 4-2 -The production schedule with an initial ramp up. Further smoothing of production can be
achieved by distributing higher production in the end of mine life to earlier periods. ............................. 21
Figure 4-3 - A production schedule with low deviation; note that production is not divided into “Ore
Mined” and “Waste Mined”, but “Processed Ore” and “Waste Dump or Stockpile”. “Ore Mined” could
be processed in the mill or stored in stockpile, and “Processed Ore” could from the mine or stockpile. . 22
Figure 4-4 -The tonnage of waste associated with every 80, 000 tonnes of ore, over 766 intervals. ........ 23
Figure 4-5 - A schedule with balanced milling and production rates, using stockpiles. ............................. 23
Figure 6-1 Simplified Cross Section through the Final TSF Dam Design ..................................................... 26
Figure 6-2 Simplified Representation (in Plan View) of the Final TSF Dam Design (not to scale) .............. 27
Figure 8-1 Site layout with main geographically significant structures ...................................................... 31
Figure 8-2 Tailings Facility Site Options ...................................................................................................... 33
Figure 8-3 Waste Rock Dump Site Options ................................................................................................. 34
Figure 9-1 Visual Interpretation of the Yearly WRD Water Balance ........................................................... 37
Figure 9-2 Simplified Interpretation of the TSF Water Balance including Annual Average Volumes of
Water contributing to each Stream ............................................................................................................ 37
Figure 10-1 Toromont pass match chart for determining truck model based on milling rate and
recommended shovel models based on truck model [12] ......................................................................... 42
Figure 10-2 Capital cost associated with each shovel truck pairing ........................................................... 45
Figure 10-3 Efficiency of each shovel truck pairing .................................................................................... 45
Figure 11-1 Yukon Drainage Basins [20] ..................................................................................................... 49
Figure 12-1 A schematic cross-section of the cover over WRD .................................................................. 58
Figure 13-1 - The production schedule and resulting cash flow model for the current pit design and
operation..................................................................................................................................................... 63
Figure 13-2 - Sensitivities of prices and operating costs............................................................................. 64
Figure 13-3 - The average annual Pb and Zn grades over the LOM ............................................................ 65
Figure 13-4 - The average annual Au and Ag grades over the LOM ........................................................... 66
12
Figure 13-5 - The variation in lead and zinc recoveries over the scheduled mine life ............................... 66
Figure 13-6 - The variation in gold and silver recoveries over the scheduled mine life ............................. 67
Figure 13-7 - The interpolated unit cost of Grum, at 8800 tpd and a strip ratio of 6. ................................ 68
Figure 13-8 - Interpolated processing unit cost for two concentrates at a milling rate of 8800 tpd ......... 68
Figure 13-9 - The route and distance from Faro to Trail [41] ..................................................................... 69
Figure 13-10 - The Korea Zinc Onsan smelter, located close to a port [44] ............................................... 70
Figure 13-11 - An aerial photograph of the port town Skagway is shown on the left and the shortest
route from Faro to Skagway is shown on the right [45] ............................................................................. 70
Figure 13-12 - Interpolated processing capital cost for two concentrates at a milling rate of 8800 tpd ... 71
Figure 16-1 Haulage Truck Specifications- Cat 785C [49] ........................................................................... 77
Figure 16-2 Ramp Design for the first push back at Whittle Pit 6 .............................................................. 78
Figure 16-3 Ramp Design for the second push back at Whittle Pit 9 ......................................................... 78
Figure 16-4 Ramp Design for the third push back at Whittle Pit 18 ........................................................... 79
Figure 16-5 Diagram showing the Meanings of each constant in the Truncated Pyramid Volume
Calculation .................................................................................................................................................. 83
Figure 16-6 – An overview of the mine site layout for context, with dimensions of paths superimposed.
For a clearer depiction of measurements, refer to subsequent figures. .................................................... 93
Figure 16-7 - View of horizontal projection distances of equipment travel paths; due to the high degree
of segmentation in the pit, dimensions are overlapping and difficult to read. A magnified view could be
found in Figure 16-6. ................................................................................................................................... 94
Figure 16-8 - A magnified view of the horizontal projection lengths of the pit ramp. ............................... 95
Figure 16-9 - Rimpull curve of the CAT 777G, with appropriate speeds determined for loaded travel on
effective grades of 3%, 4%, and 13%. ......................................................................................................... 96
Figure 16-10 - Retardation curve of an empty CAT 777G on effective grades of 0% and 7%. ................... 97
Figure 16-11 - Rimpull curve of the CAT 785D, with appropriate speeds determined for loaded travel on
effective grades of 3%, 4%, and 13%. ......................................................................................................... 98
Figure 16-12 - Retardation curve of an empty CAT 785D on effective grades of 0% and 7%. ................... 99
Figure 16-13 - Rimpull curve of the CAT 789D, with appropriate speeds determined for loaded travel on
effective grades of 3%, 4%, and 13%. ....................................................................................................... 100
Figure 16-14 - Retardation curve of an empty CAT 789D on effective grades of 0% and 7%. ................ 101
Figure 16-15 - Annual contained lead and zinc processed ....................................................................... 122
Figure 16-16 - Annual contained silver and gold processed ..................................................................... 122
13
1 Background
Northwestern Canada is home to the Grum Deposit, located in central Yukon and 200 km northeast of
the capital, Whitehorse. In addition, the site is 15 km from the town of Faro. It is understood that the
deposit is host to rich lead and zinc bearing minerals, such as galena and sphalerite, while trace amounts
of lead and silver are also expected to provide economic benefit. A basic look at the processing of these
metals is given in Section 5.
The Anvil Range district, of which the deposit is part of, contains a string of 7 deposits distributed over a
strike interval of 35 km, roughly parallel to, and 3 to 6 km to the north‐east of the major Vangorda fault
zone. The galena and sphalerite bearing massive sulfide ore includes pyritic, barytic, carbonatic and
pyrrhotitic variants, with common post depositional breccia textures. The massive sulfides are fringed
laterally and below by quartzose and graphitic disseminated sulfide mineralization, which may be
banded and/or spectacularly brecciated. The ore lenses are typically elongated. Tills in this area are from
the McConnell glaciation, and are believed to be good construction material at this stage.
2 Previous Analyses
The Grum deposit has been intercepted by two exploration drill holes reaching 218.5 and 132.2 meters
in length. These boreholes struck the orebody at 250/70 and 300/70 (trend/plunge) at UTM coordinates
of 5910.87 East, 2467.40 North and 6754.40 East, 2941.30 North. The resulting drill logs yielded both
geotechnical and qualitative geological information that can be used to get an early assessment of the
ground conditions of the Grum area. This data was complemented by a 205 meter exploration tunnel in
which fractures were mapped from its entrance, of which the exact location was unknown.
Analysis on the Grum pit design had been done previously using this data, including an attempt to
quantify the site’s rock mass properties. Following from this the potential pit area was divided into
several geotechnical domains, from which starting pit slope angles were calculated using various
numerical modeling tools. These 3 aspects will be summarized in the following Section.
2.1 Rock Mass Properties
Generally the Grum site can be divided into two main rock types, quartzite and phyllite, for which
laboratory test results were provided. From this the data provided from the boreholes allowed for the
calculation and determination a distribution rock mass quality (RQD) values where it was found that 70%
of the borehole lengths were of a value of 70 or greater. This suggested a moderate to strong rock mass.
As a result the use of both the Q and RMR76 systems were warranted, and a list of known joint sets was
compiled. Examining the joint sets present, it was found that phyllite contained 2 minor sets and 2 major
sets, while quartzite contained 2 major sets and 1 minor set (labeled Minor Set 1). The following 4 sets
were discovered:
14
● Major Set 1 – Dip: 79 Dip Direction: 043 ● Major Set 2 – Dip: 44 Dip Direction: 317 ● Minor Set 1 – Dip: 72 Dip Direction: 149 ● Minor Set 2 – Dip: 20 Dip Direction: 206
Typical RMR values of 63.5 and 67.5 for phyllite and quartz respectively and typical Q values of 0.24 and 0.31, suggested a similar quality of rock mass for each rock type. However it was clear that phyllite is the weaker of the two.
Lab test data on discontinuities for shear and normal stresses, a Mohr Coulomb strength criterion was
generated for joints in each rock type. The results of this concluded that the joint in the phyllite rock
mass is much weaker, as displayed in Figure 2-1 Conservative Mohr Coulomb Criterion for Joint
Strength, assuming no cohesion. For this reason, and its overall dominance at the mine site, all rock
mass analyses utilized the strength properties of the phyllite.
τ= σn tan(40) for Quartzite
τ= σn tan(29) for Phyllite
Figure 2-1 Conservative Mohr Coulomb Criterion for Joint Strength, assuming no cohesion
A similar procedure was carried out using the Generalized Hoek-Brown failure criterion, and similarly the
phyllite was found to be weaker, however it was evident that the controlling factor for failure was due
to joint properties. Additionally, the properties of the overburden material was analysed and a Mohr
Coulomb failure criterion was generated and appeared as such:
Evidently the overburden material is much weaker and is shown to reduce the slope angles of the pit in
early years of development.
2.2 Geotechnical Domains
Using the data from geotechnical analysis, a preliminary pit was produced, with assumed 45 degree
slopes, to determine the shape of the pit. This pit was discretized based on the orientation of each
slope. This resulted in 10 sectors with 8 distinct orientations, as shown in Figure 2-2.
15
Figure 2-2 Simplified outline of the proposed Grum Pit, divided into 10 sectors with 8 unique orientations
The pit was then divided into two geotechnical domains: rock and overburden. As seen in Figure 2-3, the
rock is composed primarily phyllite, with lesser amounts of quartzite and other minerals. Thus the rock
mass was modelled as one domain with the properties of phyllite, with properties previously discussed
in Section 2.1. As previously discussed overburden is a glacial till consisting of weaker, weathered
material and therefore its strength would govern its failure and is dominant in the southern portion of
the pit.
Figure 2-3 Visualization of the Pit's Geotechnical Domains
2.3 Slope Stability Analysis
A bench height 12 meters was chosen for the convenience of re-blocking the model from 6 m x 7.6 m x
7.6 m to 12 m x 7.6 m x7.6 m. This height corresponds to the shovel reach. Bench width was determined
to be 6.9 m, based on the relation proposed by K. Esmaeili [1]:
Bench width = 0.2*bench height + 4.5m
16
When ensuring the stability of the pit it was found that the majority of cases resulted in possible wedge
failures. Using Swedge, the probability of failure (PoF) was determined for each sector, dictated by
wedge failure at different pit slopes ranging from 65 to 85 degrees. A sensitivity analysis was also
performed with water filling 50% to 100% of the discontinuities. The resulting chosen bench face angles
are displayed in Table 2-1.
Table 2-1 Recommended Bench Face Angles for slopes governed by Wedge failure
Alternatively toppling failure was the driving factor for two faces on the northern side of the pit oriented
at 225° and 335°. Bench face angles of 80° can be acceptable, with safety factors close to or above one
at 50% saturation. It is recommended that the water pressure in these slopes is closely monitored with
pumping programs in place to control the water level.
Lastly the overall slopes used in the preliminary design were generated and checked using Rocscience
Slide software. The result is shown in Table 2-2, differentiating between host rock and overburden (OVB)
overall slope angles (OSA).
Table 2-2 Overall pit slope safety factors for each sector, at different water saturations
3 Detailed Pit Design
Following from previous work, a detailed pit could be constructed. In open pit planning, roads play a
crucial role and therefore will be incorporated early in the planning process as they can significantly
alter pit slope angles. They can also affect the economics of reserves. The overall slope angles
determined in the scoping study had not accounted for roads, therefore ignored unplanned stripping
and reserve sterilization. The next section will outline ramp specifications and its effect on the pit.
3.1 Ramp Design
The ramp will consist of two lanes; one lane for uphill traffic carrying material and the other lane for
empty downhill traffic. The two-way traffic system will be efficient and will eliminate costs for designing
two separate one-way traffic ramps. According to Couzens, 1979, the roadway of a two-way traffic ramp
should have a width greater than four times the truck width. For safety purposes, a berm, with a repose
angle of 35 and height equal to truck’s tire radius, will also be added along the sides of the ramp to
17
enhance road safety and will be added to the total roadway width. The grade of the ramp will be 10%.
The ramp curve radius is 150 m, widening the curves enough to ensure safety and reduce difficulties in
turning.
3.1.1 Ramp Width
As mentioned previously, the ramp width has to be greater than four times the width of the operating
haulage truck. Since the bench width is 6.9 m, and the Grum Pit is a small open pit (small pits normally
have bench heights of 12 m) [2], Benny Resource Group (BRG) ensured that there was enough space for
efficient and safe haulage operations. Therefore, BRG has selected the CAT 785C haulage trucks.
According to the 1965 AASHO Manual for Rural Highway Design-Mine Haulage Road [3] the space
adjacent to each lane, both right and left, should equal to one-half the width of the haulage truck. The
ramp design is shown in Figure 3-1 below. The full specification of the CAT 785C is shown in Appendix
Section 16.1.
Figure 3-1 Two-Way Traffic Ramp Design
Once the dimensions of the ramp were finalized, they were inputted into GEOVIA GEMS (GEMS) to
generate a ramp design for each pushback: 6, 9, 18 and 27. Figure 3-2 displays ramp design for pushback
27. The ramp design for pushbacks 6, 9, and 18 can be found in Appendix Section 16.1.
18
Figure 3-2 Ramp design for pushback 27. Note that the ramp exit is towards the southeast part of the pit, making the haulage distance to Waste Rock Dump shorter. Thus increasing productivity.
Since the Grum Pit is located in Yukon, the roads can expect to become icy and wet, therefore,
switchbacks were avoided during designing. As a result, a spiral ramp was designed because of the
following reasons:
Safe to operate on, especially in weather conditions like rain, ice etc.
Reduce tire wear
Unlike the switchback, the overall slope of the pit changes within a small degree (discussed in
the subsequent section)
Enhance visibility for drivers
Efficient fleet operations and increased productivity
BRG created the ramp, with iterations, to exit towards the west side of the pit, for all pushbacks, where
the dump sites are located for optimum productivity.
3.1.2 Ramp Section Design
One of our main targets is to maintain low costs during the life of the mine. Poorly constructed and
maintained roads incur extra and large haulage costs and can become a safety hazard. Therefore, a good
ramp design is necessary. The ramp will be comprised of four different layers discussed in Table 3-1
(occurring in the order presented, from top to bottom).
19
Table 3-1 Purpose of each layer in designing a ramp
Figure 3-3 shows the section of the ramp. The material for each layer is dependent on both economic
and operating factors. Operating factors, for instance, are contingent on material’s ability to distribute
estimated loads from haulage trucks.
Figure 3-3 Construction Layers of the Ramp Surface
3.1.3 Ramp Maintenance
Deterioration of the roads can generate extra costs, which can place a dent in the economics of the
operation. A damaged road can reduce the life of equipment significantly, thus incurring extra capital
costs. Therefore to ensure the operation runs as planned, the following objectives will be met:
Drivers will be recommended to drive on different areas of the lane to prevent formation of ruts on roads due load concentration
Snow and ice will need to be immediately removed using a motor grader
Spillage of material from loaded trucks will be prevented as they will cause unnecessary bumps, causing tire wear
Maintain ramp grade and slope and smooth depressed surfaces
3.2 Pit Slope Geometry
By adding the ramp, the overall slope angle of the pit changes. When constructing the ramp, the aim
was to ensure that ramp was designed as intended without significantly changing the economics. The
ramp changed the overall slope angle of the walls on the west side of the pit to an insignificant degree
20
and therefore the change was neglected. The walls on the east side of the pit, however, had their overall
slopes change significantly after the construction of the ramp. These changes are summarized in Table
3-2 and are visualized in Figure 3-4 with calculations shown in Appendix Section 16.2. After calculating
the new overall slope angles, they were re-entered into Whittle to determine the new economics of the
operation (discussed in Section 13).
Table 3-2 Adjusted Pit Slope Parameters
Figure 3-4 Pit Wall Geometry for Sectors 8, 10, 9&1
4 Production Scheduling
Using the pit design, as described in Section 3, the production schedule produced is shown in Figure 4-1.
21
Figure 4-1 - The production schedule needed to meet a 3.2 Mt mill capacity. The first year could be ramped over the preceding two years, as indicated by the arrow.
Following this exact schedule would be unreasonable due to high fluctuation in mining rates, especially
in the first year. Assuming the first year could have prestripping over earlier years, the resulting
production would yield Figure 4-2.
Figure 4-2 -The production schedule with an initial ramp up. Further smoothing of production can be achieved by distributing higher production in the end of mine life to earlier periods.
Although the deviation of production has been reduced, there is still a significant difference between
the higher beginning and ending rates, with the lower rates at the middle of the mine life. To reduce
variation of production rates further, the production of years 12 to 17 could be distributed to the years 4
to 11. The resulting production theoretically has a balanced production rate of 27 Mt per year, as shown
in Figure 4-3.
0
10000000
20000000
30000000
40000000
50000000
60000000
70000000
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Production Schedule
Ore Waste
0
10000000
20000000
30000000
40000000
50000000
60000000
70000000
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Production Schedule with Ramp Up
Ore Waste
22
Figure 4-3 - A production schedule with low deviation; note that production is not divided into “Ore Mined” and “Waste Mined”, but “Processed Ore” and “Waste Dump or Stockpile”. “Ore Mined” could be processed in the mill or stored in stockpile, and
“Processed Ore” could come from the mine or stockpile.
However, forwarding production earlier does not mean only waste is forwarded, but ore associated with
that waste. For this reason, stockpiles would be required as more ore would be mined than the mill
would be capable of handling during early mine life. Later in the mine life, ore extraction would not
meet the milling capacity, so stockpiles would be consumed to do so.
To determine the tonnage and grade of the stockpiles, the ore that follows the forwarded production
needs to be determined. A Whittle schedule was made with a smaller milling limit, to determine how
the amount of waste and the grade changes per unit of ore over the mine life. This was accomplished by
producing a schedule with a smaller milling limit, which would show how much waste needed to be
extracted for a certain tonnage of ore.
Due to Whittle’s hardcoded limits of 999 periods and seven minutes per iteration, the smallest unit of
ore used was one fortieth of the target milling rate, at 0.08 Mt/period. The resulting schedule
represented how much waste is required to extract every 0.08 Mt of ore. The resulting breakdown is
shown in Figure 4-4.
0
5000000
10000000
15000000
20000000
25000000
30000000
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Theorectical Balanced Production Schedule
Processed Ore Waste Dump or Stockpile
23
Figure 4-4 -The tonnage of waste associated with every 80, 000 tonnes of ore, over 766 intervals.
The appropriate tonnages of ore and waste mined, as shown in Figure 4-4, can be matched with the
target production, as in Figure 4-3. The intervals of waste and ore were integrated to best match the
target production of each year. In the years which the tonnage of ore mined exceeds mill capacity, ore
would be stockpiled. Meanwhile, in years which ore production does not meet mill capacity, the
stockpile would be processed. The resulting schedule is shown in Figure 4-5, in terms of:
Stockpiled ore: Ore that has been mined and is stockpiled due to exceeding mill capacity.
Processed mined ore: Ore that is processed after extraction
Processed stockpile ore: Ore sent to the mill from stockpiles
Mined Waste: Waste rock without economic value, sent to waste rock dump (WRD)
Figure 4-5 - A schedule with balanced milling and production rates, using stockpiles.
0
500000
1000000
1500000
2000000
2500000
3000000
3500000
4000000
2
33
64
95
12
6
15
7
18
8
21
9
25
0
28
1
31
2
34
3
37
4
40
5
43
6
46
7
49
8
52
9
56
0
59
1
62
2
65
3
68
4
71
5
74
6
Waste Associated with every 80000 Mt of Ore
Ore Waste
0
5000000
10000000
15000000
20000000
25000000
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Balanced Production Schedule with Stockpiles
Stockpile Processed Ore Mined and Processed
Ore Stockpiled Waste
24
5 Preliminary Processing Design
Before an appropriate Tailings Storage Facility (TSF) design could be constructed it was essential that a
preliminary design of the ore processing was considered. For the purposes of this study a high level
approach was taken due to a lack of geochemical data and laboratory testing which could more
accurately represent the results of processing.
To obtain a good sense of the required processing method the amounts of metal contained in the
extracted ore was compared. From these results, seen in Table 5-1, it is clear that the focus will be
placed on the concentration of lead and zinc.
Table 5-1 Summary of Contained Metals before processing, Recovered Metals and Average Metal Grades
One such method includes the use of lead – zinc froth floatation, which would produce two separate
concentrates, one lead and one zinc, with the gold and silver reporting as pollutants in both streams.
From here the concentrates would be sold to the smelter company. It has been suggested that such a
process has the potential to generate a concentrate containing a lead grade of 60%, while zinc could
reach a grade of 56% [4].
Some preliminary metallurgical lab data was provided for the site (see Appendices, Section 16.4). This
data appeared to match the recovery range of 80 to 90%, common for lead and zinc floatation [4]. As a
result it was decided that this data would be sufficient for use in a preliminary processing mass balance.
However note that it is recommended that future lab tests are carried out in the future for more
accurate results.
Using the recovered metal data produced from these assumptions, and the material data generated for
the pit using Whittle, average tailings grades were found using the following equation:
𝑀𝑒𝑡𝑎𝑙 𝐺𝑟𝑎𝑑𝑒 𝑜𝑓 𝑇𝑎𝑖𝑙𝑖𝑛𝑔𝑠 = (𝑀𝑒𝑡𝑎𝑙 𝐼𝑛𝑝𝑢𝑡 − 𝑀𝑒𝑡𝑎𝑙 𝑅𝑒𝑐𝑜𝑣𝑒𝑟𝑒𝑑)
100 ×(𝑂𝑟𝑒 𝐼𝑛𝑝𝑢𝑡 − 𝑀𝑒𝑡𝑎𝑙 𝑅𝑒𝑐𝑜𝑣𝑒𝑟𝑒𝑑)
From this it was found that tailings will have an estimated grade of 0.41% lead and 0.26% zinc. A
preliminary mass balance was then completed assuming 1 tonne of feed, and the results of which can be
seen in Table 5-2 below.
Metal Total Input (Metric Tonnes) Recovered (Metric Tonnes) Input Grade
Lead (%) 113557353 99940258 2.056
Zinc (%) 180277921 158668979 3.264
Gold (g) 35424963 21251519 0.641
Silver (g) 1933521098 995616297 35.003
25
Table 5-2 Results of the Preliminary Mass Balance for Froth Floatation
The results of Table 5-2 were calculated assuming the mass balance for each stream follows the
processing mass balance equation written as:
𝐹𝑒𝑒𝑑(𝑖𝑛𝑝𝑢𝑡 𝑔𝑟𝑎𝑑𝑒) = 𝐶𝑜𝑛𝑐𝑒𝑛𝑡𝑟𝑎𝑡𝑒(𝑐𝑜𝑛𝑐𝑒𝑛𝑡𝑟𝑎𝑡𝑒 𝑔𝑟𝑎𝑑𝑒) + 𝑇𝑎𝑖𝑙𝑖𝑛𝑔𝑠(𝑡𝑎𝑖𝑙𝑖𝑛𝑔𝑠 𝑔𝑟𝑎𝑑𝑒)
In addition it has been assumed that if the overall processing is considered an average mass balance can
be taken between the two streams. This was done to gain a sense of the overall amount of materials
reporting to the TSF, which is around 96% of every ton of ore processed, as seen in Table 5-2. It is
important to note that this method of estimation represents a very rough estimate of the overall
processing mass balance. As such, careful metallurgical testing should be conducted in order to produce
an accurate processing mass balance which accounts for the 2 separate concentrate streams and other
factors, such as the mass balances of individual crushers, grinders, and floatation cells required in the
circuit. However, for this level of study the current analysis is sufficient to conduct further estimates for
tailings management purposes.
6 Tailings Storage Facility Design
At this stage it has been suggested that the specifics regarding the stability of the impoundment are not
essential, and can be determined in later design stages. Instead this level of design will focus on the
appropriate geometry necessary to store the tailings material. In doing this, it allows for the estimation
of a possible design footprint and therefore an appropriate site layout. This document will cover the
technical details involved in finding a preliminary dam geometry while the process of site layout and
selection will be covered in its own document.
6.1 Selection of an Appropriate Cover System
The site has been marked as a massive sulfide deposit, which is capable of producing acidic effluent, and
therefore appropriate measures must be taken to inhibit acid mine drainage (AMD). Due to this a
proposed tailings storage design should be able to keep acid generation to a minimum, and mitigate the
release of potentially harmful effluent to the environment.
Given that the Faro area sees a regular amount of precipitation (approximately 316 mm annually), and it
can considered a humid climate, prevention of AMD using dry tailings throughout the mine life could
prove difficult [5]. As a result, the abundant amount of nearby water sources suggests that a designed
water cover could provide an effective strategy to combat AMD throughout the mine life. Therefore the
preferred method of tailings impoundment in humid climates, a wet cover system, will be employed [6].
Concentrate Grade Amount Reporting to Conc. Tailings Grade Amount Reporting to Tailings Input Grade
Pb 60% 5% 0.41% 95% 3.305%
Zn 56% 3% 0.26% 97% 2.075%
Avg Mass Balance 4% 96%
26
Typical water covers provide protection against AMD using a relatively thin layer of water that prohibits
oxygen ingress to the acid generating tailings [6]. A water cover thickness of 2 m has been selected for a
conservative approach. This has been done in response to the heightened social sensitivity to the
spillage of effluent as a result of the nearby Faro site; the Faro mine is currently a major remediation
project for contamination due to old mine workings. By using a thicker water cover this should
significantly reduce the possibility for acid generation from the tailings. More details on the community
and the effects of the Faro site are covered in Sections 8 and 11, Site Layout and Environmental and
Social Impact Assessment, respectively.
6.2 Design of the Dam Geometry
After the cover system was selected a numerical model was generated to determine the overall
geometry of the required tailings dam. In doing this, the first fundamental assumption was that the
generated tailings, when first deposited as a slurry, would have a moisture content of 40%, by weight of
solids, which is within the range suggested by McPhail – 30 to 50% – for freshly placed tailings [5]. Also
as part of the preliminary design stage a conservative dam geometry has been suggested in advance,
utilizing a crest width of 8 m, a berm width of 15 m and slope of 1:2.5, height to width, on the
downstream face. The beach of the impoundment will also assume a gradient of 1:2.5. This produced
the final design geometry presented in Figure 6-1, below.
Figure 6-1 Simplified Cross Section through the Final TSF Dam Design
In order to reach this final design the numerical model took into account the previous geometrical
assumptions along with the water content of the tailings to attempt to find an appropriate dam
configuration to accommodate the tailings. For this to work an initial estimate of the amount of tailings
volume (including water) produced per year was generated. This was done by using the ore tonnages
sent to the mill, obtained from Whittle Analyses, and applying the assumed 40% water content and
average ore density of 2.64 ton/m3, found from earlier lab testing. The results of this can be seen in
Appendix Section 16.5. Note that the values are presented in yearly amounts, which is important for
determining the mine’s water balance, covered in Section 9.
Knowing the volume of material going into the TSF each year, the geometry can be used to predict the
annual height of the tailings. This was done by utilizing the expression for the volume of a truncated
27
pyramid (explained in Appendix Section 16.6), presented by Bronstein et al. [6]. The truncated pyramid
shape could be used to represent the geometry of capacity of the TSF. In this case it is assumed that the
shape of the tailings as it fills the dam will be that of the truncated pyramid when it is inverted, or
flipped on its head.
With the tailings volume accounted for, the numerical model uses this equation in determining the
height of the tailings, and its annual rate of rise, shown in Appendix Section 16.7. The model does this by
taking the storage width and length, graphically shown in Figure 16-5, as well as the desired dam height
as inputs. Geometry is then used to calculate the overall length and width of the TSF, assuming a
rectangular shape. Furthermore, the model is able to determine the number of slopes and berms the
downstream slope will require, as visually shown in Figure 6-1.
After initially constructing the model it was found that the mountainous landscape in the vicinity of the
Grum deposit provided significant challenges for the previous assumptions. An additional model was
created to account for the change in gradient of the area the TSF was placed. However results showed
that this change would cause large losses in dam capacity, requiring larger amounts of space than the
prior model. As a compromise the first model was adjusted by assuming a natural slope can take the
place of one of the downstream slopes, as shown at the top of Figure 6-2. This eliminated the need for a
downstream slope on one end of the dam, reducing its overall length, and assumes that the natural
slope could be re-graded to the necessary 1:2.5 height to width ratio.
Figure 6-2 Simplified Representation (in plan view) of the Final TSF Dam Design (not to scale)
The downside of this assumption is that it would require that the base of the TSF is leveled, which may
require a large amount of material. Therefore for a preliminary phase this design should represent a
28
conservative approach and different strategies may be used to reduce the cost and size of this design.
The final design parameters are summarized in
Table 6-1.
Table 6-1 Summary of Key Parameters of the Final TSF Design
6.3 Considerations for Dam Construction
A final estimate of the volume of construction material necessary to construct the final dam design was
calculated on a yearly basis. These values can be seen in Table 6-2. This was estimated by using the
product of the estimated final volume of building material and the ratio of yearly tailings volume to the
final tailings volume; the latter is shown as the approximate dam completion. The purpose of this
exercise was to get a “ball-park” estimate of how much material will be needed to construct it. This
result could then be used to see if additional material will be required for construction, and can have
ramifications on the final cost estimates, however this was done as a point to move on from for future
studies.
Table 6-2 Estimates for required Material needed to construct the Final TSF design
Due to the foreseen high level of public scrutiny and the large consequences of failure, a downstream
method of deposition and dam creation will be used. This appears to be most conservative as the new
materials are placed on older dam materials, rather than on top of the tailings. Downstream deposition
Dam Area 2.2 km2
Length 1733 m
Width 1266 m
Final Dam Capacity 5.31E+07 m3
Total Tailings Held 3.75E+07 m3
Free Board 11.87 m
Summary of Final TSF Dimensions
End of Production Year Tailings Capacity Needed (m3) Approx. Dam Completion Additional Dam Material Needed (m3/year)
1 1.20E+05 0% 1.38E+05
2 7.77E+05 2% 7.58E+05
3 2.57E+06 6% 2.07E+06
4 4.90E+06 12% 2.69E+06
5 6.99E+06 17% 2.41E+06
6 8.85E+06 21% 2.14E+06
7 1.12E+07 27% 2.76E+06
8 1.36E+07 33% 2.76E+06
9 1.60E+07 39% 2.76E+06
10 1.84E+07 45% 2.76E+06
11 2.08E+07 50% 2.76E+06
12 2.32E+07 56% 2.76E+06
13 2.56E+07 62% 2.76E+06
14 2.80E+07 68% 2.76E+06
15 3.04E+07 74% 2.76E+06
16 3.28E+07 79% 2.76E+06
17 3.51E+07 85% 2.76E+06
18 3.75E+07 91% 2.76E+06
19 3.99E+07 97% 2.76E+06
20 4.13E+07 100% 1.59E+06
Total (m3) 3.98E+08 4.76E+07
29
allows for better control over the engineering properties of the dam structure and as such should
produce a more stable design.
Additionally, note that there is a need for an impervious core material, likely clay or compacted local till
material, which is not shown in Figure 6-1. This would be done in order to manage the amount of flow
out of the toe of the dam, which could lead to potential instability in the design.
It was also previously stated that the area sees regular precipitation throughout the year and as a result
designed spillways should be placed on the abutments of the dam, where the dam makes contact with
the natural slope. These spillways should reduce the chances of overtopping if a flood event occurs,
which is critical in ensuring the continued stability of the design. Additionally, should discharge through
the spillways be necessary, a form of water diversion, should be created around the dam so that water
can be lead to the water treatment facility. From here any excess water can be released safely to the
environment, however the specifics of the design of these diversions is left to later studies.
Lastly preliminary data was collected for the Grum area’s overburden material and was analysed; this
data and results are tabulated in Appendix Section 16.8. The findings of this analysis found that
according to the ASTM soil classification scheme the material is an SC-Clayey Sand. This represents good
quality building material, characteristic of glacial tills, however some uncertainties in the lab test results
suggests more detailed testing is required; this is further explained in Appendix Section 16.8. For
seepage purposes this material has a permeability ranging from 5.5x10-9 to 5.5x10-6 m/s [7]. This data
therefore suggests that natural liner material obtainable from the local area will have a permeability of
5.5x10-9 m/s at best. For this reason the water balance, discussed in Section 9, will utilize this value.
7 Design of the Waste Rock Dump
Following the TSF design, the disposal of unprocessed material will also be an important factor in the
mine design of this location. Just like the tailings, the waste rock can also be considered as Potentially
Acid Generating (PAG), and as a result a low permeability mat material will need to be placed on the
selected site of the WRD. Additionally it was decided that only one dump would be necessary as any
Non-PAG material will be assumed to be used immediately for dam construction at this stage of design.
Considering this, a similar approach was used to design the WRD as the TSF design. In this case the
overall waste rock generated over the life of mine was considered from the whittle model. This was
done because it allows for the overall footprint of the design to determined using a numerical model;
yearly waste values are also not sensitive to the yearly water balance.
The numerical model used takes on the assumption that the slopes of the WRD will take on the same
geometry as the downstream face of the TSF, as suggested prior to starting the design. This conservative
assumption will allow for a focus on the selection of an appropriate site rather than its overall stability.
Just as the TSF, the selection of an appropriate site is covered in the Section 8.
30
From here the amount of waste volume was estimated by applying both a bulking factor, due to the
mechanical handling of material, and a compaction factor, assuming efforts will be made to
mechanically compact the waste [8]. The calculation of the Final waste rock volume, using the previously
assumed density of 2.64 ton/m3, can be seen in .
Table 7-1.
Table 7-1 Summary of Total Waste Rock Volume Determination with Suggested Volume Adjustment Factors [8]
By specifying the length and width of the rectangular WRD, the numerical model finds the height of the
dump required to accommodate the volume of waste. By testing different variations of the WRDs, a final
design was found, and its geometry is summarized in Table 7-2. The method by which these geometries
were chosen are further discussed in Section 8.
Table 7-2 Summary of Final WRD Design Parameters
Similar to the TSF, water runoff from the WRD should also be diverted to a water treatment facility from
which water can be safely released to the environment. As a result of this the diversion of runoff water
would also be done through the use of appropriate ditches following the perimeter of the facility and
would direct it to the site’s water treatment facility. This process would occur until the end of
production, where an appropriate dry covering system will be used; this is further described in Section
12.
Total Waste Rock 3.5E+08 Metric Tonnes
Avg Feed Density 2.64 Ton/m^3
Bulking Factor 1.15
Compaction Factor 0.95
Volume of Waste 1.5E+08 m3
Overall Dimensions Value Units
Length 1500 m
Width 1500 m
Dump Height 109 m
Slope Parameters Value Units
# of Berms 10 Berms
# of Slopes 11 Slopes
Top Dimensions Value Units
Length 654 m
Width 654 m
31
8 Site Layout
Figure 8-1 Site layout with main geographically significant structures
8.1 Background
The mine site evolves around the pit and the material excavated from it. The tailings pond and the waste
rock dump are the most significant components of the mine site next to the open pit. Both require a
large amount of space and are permanent installations on the landscape. The tailings storage facility
(TSF) and waste rock dump (WRD) generate acid mine drainage due to the presence of sulphides in the
ore. This presents certain requirements for site choice for these structures. Both the TSF and WRD
require an impermeable liner to ensure a layer of water remains on the tailings to slow acid generation
and so the bleed water running off the waste rock does not flow into the nearby streams. Design of TSF
and WRD were seen in Section 6 and Section 7, respectively, and an environmental risk matrix in Section
11.3. Emphasis was given to impacts of placement on water systems and the community.
8.2 Placement Methodology
When determining placement, a minimum distance of 150 m from streams and public roads is used as a
buffer zone and stream diversion is considered if necessary. The TSF and WRD are designed to hold the
waste produced from the mine and mill. An iterative process of selecting the site and calculating the
height, length and width to meet capacity is the main methodology to physically determine the best
sites. For these sites, economic, environmental and social effects of the design are used to compare
each alternative to find the most acceptable solution. Appendix Section 16.18 shows the economic,
environmental and social considerations and indicators when comparing the options for the site of the
tailings facility and waste rock dump.
32
Other less geographically significant features present on the mine site include:
Ore mill including ore stockpile
Water treatment plant
Topsoil stockpile
Site admin office, metallurgical testing lab and parking
Septic field and waste management facility
Garden nursery, operation beginning within last 5 years of life
Maintenance garage
Access roads and power corridors
Explosives magazine
The site for each of the above features depends on the structure they cater to. The mill will be located
between the pit and the TSF, the site office and parking will be located at the entrance of the mine site,
roads will go where needed, the maintenance garage near the exit of the ultimate pit ramp, etc. The
explosives magazine will also be located away from the buildings, pit and waste facilities; the blast
radius of a fully stocked magazine will determine the distance. Figure 8-1 shows the complete site
layout.
8.3 Tailings Storage Facility
The TSF was placed first to ensure it was away from homes, infrastructure and streams with the use of
the natural landscape to confine at least part of the structure. The options were chosen based on
capacity and then compared against the other options for economic, environmental effects outlined in
Appendix Section 16.9. The chosen site uses a south dipping mountain side to create a confining slope.
The facility is placed within one watershed with potential to expand without diverting the streams
leading to the productive Vangorda Creek. Due to the slurry nature of the Grum tailings, the tails will be
piped to the site from the mill. The site selection considers pipe and access road crossings over streams.
Figure 8-2 shows the three site options for the TSF. TSF one was the chosen option and it is located
North-East of the pit.
33
Figure 8-2 Tailings Facility Site Options
After the TSF site was determined, three potential sites were compared for the waste rock dump. With
similar constraints as the tailings facility but solid rather than slurry, three geometries were determined.
Again due to the Acid Mine Drainage caused by the sulphides in the waste, water was a concern. The
chosen site avoids stream diversion has the possibility to expand. The dimensions of the dump also bring
down the height which is a concern for the tourism community trying to show off beautiful terrain.
Figure 8-3 details the three potential sites and Appendix Section 16.10 outlines the economic,
environmental and social comparison of the potential WRD sites. After placing the WRD on the chosen
site, geometry and distance from open pit allowed the dump to move inward, away from the road and
closer to the pit. Specifics associated with the WRD design were found in Section 7.
34
Figure 8-3 Waste Rock Dump Site Options
8.4 Additional Site Requirements
8.4.1 Processing Mill
The processing mill will contain crushers, grinders and two flotation circuits for zinc and lead. The mill is
located just north of the pit. The placement avoids truck and pipe crossings, with each other and/or
streams. The mill was also placed directly upstream of the pump pond where, if a mill breach occurred,
the effluent would travel.
8.4.2 Explosives Storage and Handling
A contract will be entered into with a recognized supplier of mining explosives, to establish his own
facilities in the south west of the waste rock facility, well away from the local population and mine
activities, and to supply emulsion as needed.
8.4.3 Technical Departments
The site admin office, engineering department, metallurgical testing lab, revegetation nursery, septic
field and human waste treatment facility will be located at the entrance of the mine site surrounded by
existing vegetation. These buildings will be surrounded with parking to provide easy access and distance
from haul trucks.
35
8.4.4 Environmental Systems
The water treatment plant and topsoil stockpile are located east of the pit between the small pump
pond and tailings facility. The pipe leading from the mill to the water treatment plant must travel below
the road surface to bring the reusable water to the plant. A pipe runs from the tailings facility to the
water treatment plant providing a safe discharge of extra water. The topsoil pile will be covered during
operation and used for progressive remediation efforts around the mine site.
9 Water Balance of the Mine Site
Having looked at the major causes for concern when dealing with the contamination of water, the water
balance can provide a key tool for managing the water flow around the mine site. As previously
mentioned the TSF alone can account for up to 80% of all water movements at a mine site, and as a
result it, and the WRD, will be the focus of this exercise [9]. Table 9-1 tabulates the key coefficients, as
suggested by McPhail, which were used in estimating the mine water balance for both the TSF and WRD.
Table 9-1 Key Coefficients used in Conducting the Mine Water Balance [9]
9.1 Water Balance of the Waste Rock Dump
Starting with the simpler of the two designs, the PAG materials in the WRD provides a challenge for
maintaining good water quality in the nearby environment. This balance then aims at determining the
appropriate amount of water a water treatment plant can expect to process on a yearly basis due to the
WRD.
The key source of water that will reach the WRD is assumed to be due to precipitation. Before
continuing note that in this area of the Yukon around a third of the annual precipitation is received as
snow. However for the purposes of this preliminary analysis it will be treated as rain in all cases.
Factor Low High Comments
Pond Area 10% 30% of Beach Area
Pond 100% 100%
Dry Tailings & Beach 50% 60% Average used for WRD
Pond Rate 80% 100%Low is in the Summer; High in Winter Months.
Assumed 100% for the TSF.
Wet Beach Rate 60% 80% of Pond Evap Rate
Damp Beach Rate 40% 60% of Pond Evap Rate
Dry Beach Rate 0% 20%of Pond Rate (Depends on Rate of Rise of Pond).
Average used for WRD.
Seepage Rate
Moisture Content 30% 50% Recommended Range for Newly Placed Tails
Interstitial Water AllowanceSubtract from m, above (will reduce over time
due to desication; does not affect seepage)
Remaining Water Change 50:50 between evaporation & seepage
Amount 30% 50% of the water pumped onto the dam (including
50%
Underdrainage & Decant Water
15%
Infiltration
Equals permiability of Tailings or the Foundation (whichever is lower) and can incorporate
representative pond depth.
Seepage
Runoff
Evaporation
36
Knowing the amount of annual rainfall in the area is 316 mm per year, and that the WRD will be 1500 m
by 1500 m (even at the end of the first year of production) a quick estimate of annual volume can be [5].
From here an average runoff coefficient of 55% for dry tailings and beaches can be used to determine
how much of the precipitation will stay in the tailings [9]. Additionally an annual amount of evaporation
can be estimated by applying an average evaporation coefficient of 10% for dry beaches alongside the
300 mm mean annual evaporation rate for bodies of water in this area of Canada [10]. The result of this
is 252 thousand m3 of net retained water (shown as Net Water Balance in Table 9-2) within the WRD
annually. In addition the result shows that 391 thousand m3 of runoff water is produced, which must be
treated each year. Furthermore the annual average results are summarized and visually depicted in
Figure 9-1.
Table 9-2 Summary of the WRD Water Balance
When examining these results the constant values across all years can be attributed to the fact that the
facility is expected to reach its maximum outer dimensions after the first year of production.
Additionally the basic nature of this study does not account for the variable wetness of the WRD, which
could affect the evaporation rate, as suggested by McPhail [9].
End of Production Year Precipitation (m3) Runoff (m3) Evaporation (m3) Net Water Balance
1 7.11E+05 3.91E+05 6.75E+04 2.52E+05
2 7.11E+05 3.91E+05 6.75E+04 2.52E+05
3 7.11E+05 3.91E+05 6.75E+04 2.52E+05
4 7.11E+05 3.91E+05 6.75E+04 2.52E+05
5 7.11E+05 3.91E+05 6.75E+04 2.52E+05
6 7.11E+05 3.91E+05 6.75E+04 2.52E+05
7 7.11E+05 3.91E+05 6.75E+04 2.52E+05
8 7.11E+05 3.91E+05 6.75E+04 2.52E+05
9 7.11E+05 3.91E+05 6.75E+04 2.52E+05
10 7.11E+05 3.91E+05 6.75E+04 2.52E+05
11 7.11E+05 3.91E+05 6.75E+04 2.52E+05
12 7.11E+05 3.91E+05 6.75E+04 2.52E+05
13 7.11E+05 3.91E+05 6.75E+04 2.52E+05
14 7.11E+05 3.91E+05 6.75E+04 2.52E+05
15 7.11E+05 3.91E+05 6.75E+04 2.52E+05
16 7.11E+05 3.91E+05 6.75E+04 2.52E+05
17 7.11E+05 3.91E+05 6.75E+04 2.52E+05
18 7.11E+05 3.91E+05 6.75E+04 2.52E+05
19 7.11E+05 3.91E+05 6.75E+04 2.52E+05
20 7.11E+05 3.91E+05 6.75E+04 2.52E+05
Total Over LOM 1.42E+07 7.82E+06 1.35E+06 5.05E+06
37
Figure 9-1 Visual Interpretation of the Yearly WRD Water Balance
9.2 Water Balance of the Tailings Storage Facility
Continuing from the WRD water balance the TSF balance uses the water added through the tailings as a
starting point. These tailings (40% water by mass) are then deposited, and approximately 15%
(subtracted from the 40%) by mass of the tailings becomes trapped in the voids. The remaining 25% is
free as bleed water and floats above the tailings contributing to the required water cover. The water
cover is then susceptible to losses, due to seepage and evaporation, and further gains from precipitation
[9]. This process is visually depicted in Figure 9-2 below.
Figure 9-2 Simplified Interpretation of the TSF Water Balance including Annual Average Volumes of Water contributing to each Stream
When considering the tailings water balance the net water balance will be considered as the amount of
water contributing to the 2 m thick water cover each year; this is represented by the light blue in Figure
9-2. The initial tailings water can be easily calculated, and was mentioned previously in Appendix Section
16.5 as the total water added. In addition an interstitial, or trapped water volume can be calculated
from the tailings using 15% water by mass of tailings [9]. Precipitation is then calculated using the rate of
38
316 mm per year, but using the pond area and a factor of 5, as the catchment area is cited as being
upwards of 5 times the pond area in valley locations in many cases [9]. For evaporation the lake
evaporation rate of 300 mm per year was used with the pond area and an evaporation coefficient of
100% [10].
Seepage was estimated by using the assumed minimum permeability of nearby materials equal to
5.5x10-9 m/s (or 0.17 m/year), as explained in Section 6.3. The amount of seepage water per year was
then solved by the product of the catchment area and the yearly permeability. This and other values can
be seen in the full water balance in Appendix Section 16.11. The final water balance is then found using
the following equation:
𝑁𝑒𝑡 𝑊𝑎𝑡𝑒𝑟 𝐵𝑎𝑙𝑎𝑛𝑐𝑒 = 𝑇𝑎𝑖𝑙𝑖𝑛𝑔𝑠 𝑊𝑎𝑡𝑒𝑟 − 𝐼𝑛𝑡𝑒𝑟𝑠𝑡𝑖𝑡𝑖𝑎𝑙 𝑊𝑎𝑡𝑒𝑟 + 𝑅𝑎𝑖𝑛𝑓𝑎𝑙𝑙 − 𝑆𝑒𝑒𝑝𝑎𝑔𝑒 − 𝐸𝑣𝑎𝑝𝑜𝑟𝑎𝑡𝑖𝑜𝑛
Going through the water balance it is seen that the total water movements across the life of mine sum
to 84.2 million m3 of water. In order to obtain a better picture of where this water is going the
contributions of each stream was calculated and was tabulated in Table 9-3. Also average values for
each stream were calculated and presented graphically in Figure 9-2.
Table 9-3 Summary of the water movement contributions for water movement of each stream in the TSF water balance
As seen here it is seen that the largest contributor to water losses over the mine life is due to seepage,
accounting for 16%. Due to this it is likely that this water will have to be drained to the water treatment
facility, contributing an average value of 0.745 Million m3 of water annually. Combining this value with
that of the WRD amounts to 1.132 Million m3 of water that must be processed, and released to the
environment, by the water treatment plant every year. As a result some form of water holding pond
may be needed to accommodate the rate of processing and a similar dam geometry can be assumed for
it at this stage, however the specifics of this will be left to future studies.
In addition if the amount of water needed to ensure the water cover remains 2 m thick is considered it is
found that there is a deficit of water after the first year of production. This was found by working the
computed water balance back into the TSF model described in Section 6.2. By doing this it was found
that a pumping schedule, tabulated in Appendix Section 16.12, could be added into the water balance to
ensure a 2 m cover is maintained. Table 9-4, akin to Table 9-3, was created in order to fully realize the
Precipitation 30%
Evaporation 6%
Tailings Water 24%
Seepage 16%
Bleed Water 15%
Trapped (Interstitial) Water 9%
Total Water Balance 100%
Water Balance Contributions
39
impact of supplementary pumping, shown below. Note that this is now a breakdown of 104 Million m3
in total water volume movement.
Table 9-4 Summary of water contributions for water movement of each stream after incorporating additional pumping
The pumping schedule sees that an additional 0.855 Mm3 of water is added over top of the tailings in the
first year, while all subsequent years require water to be pumped out. Now accounting for 19% of total
water movements across the mine site this can be seen as a large cost to this design. However by
analysing the new TSF design the incorporation of the water balance increases the final freeboard to just
shy of 22 m. As such this presents the possibility for future modifications to the TSF, or the possibility of
allowing for excess water to accumulate in later years to reduce the need for pumping.
10 Operations Planning
10.1 Equipment Selection and Pricing Model
The mining equipment fleet selected and its change over the mine life is shown in Table 10-1. Details on
selection methodology are detailed in the following sub-sections.
Table 10-1 Summary of chosen loading and haulage fleet
Years into Production -2 -1 1 to 17 18
Haul Trucks CAT 785D 150 ton 7 13 21 8 Shovels CAT 6040 22 m3 1 1 1 1 Front End Loaders CAT 994F 7.7 m3 1 1 1 1 Track Dozer CAT D9T 13.5 m3 1 2 2 1 Wheel Dozer CAT 854K 7.9 m3 1 1 1 1 Motor Grader CAT 24M 16’ blade 1 2 2 1 Articulated Truck CAT 735B 24 m3 1 1 1 1 Vibratory Compactor CAT CS-64 112 kW 1 1 1 1 Tool Carrier CAT IT 38H 2.5 m3 1 1 1 1 Diesel Drill --- 4.5’’ to 8.5’’ 2 4 6 1 Secondary Drill --- 4.5’’ to 5.5’’ 1 1 1 1
Precipitation 24%
Evaporation 5%
Tailings Water 19%
Seepage 13%
Bleed Water 12%
Trapped (Interstitial) Water 7%
Supplementary Pumping/Drainage 19%
Total Water Balance 100%
Wate Balance Contributions with Supplementary Pumping
40
The equipment selection model used selects the model and quantity of equipment best suited to the
geometry of the mine site, available work hours, the target milling rate, and expected strip ratio at a
certain point in the mine’s production life. The process of selection is listed as follows:
1. Calculate daily production
𝐷𝑎𝑖𝑙𝑦 𝑊𝑎𝑠𝑡𝑒 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 = 𝐷𝑎𝑖𝑙𝑦 𝑂𝑟𝑒 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 ∗ 𝑆𝑡𝑟𝑖𝑝 𝑅𝑎𝑡𝑖𝑜
2. Determine effective number of working hours per day
𝐷𝑎𝑖𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑣𝑒 𝐻𝑜𝑢𝑟𝑠 = (𝐷𝑎𝑖𝑙𝑦 𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒𝑑 𝐻𝑜𝑢𝑟𝑠 − 𝐵𝑟𝑒𝑎𝑘𝑠) ∗ 𝐸𝑓𝑓𝑖𝑐𝑖𝑒𝑛𝑐𝑦
3. Calculate the effective hourly production
𝐻𝑜𝑢𝑟𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 = 𝐷𝑎𝑖𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 ∗𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑣𝑒 𝐻𝑜𝑢𝑟𝑠
𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒𝑑 𝐻𝑜𝑢𝑟𝑠
4. Select potential trucks and shovels from pass match chart
5. Determine the lengths and grades of truck routes (bassed on section on Site Layout)
6. Determine the load time for a truck and shovel pairing
𝐿𝑜𝑎𝑑 𝑇𝑖𝑚𝑒 = 𝐹𝑖𝑟𝑠𝑡 𝑃𝑎𝑠𝑠 + 𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑃𝑎𝑠𝑠𝑒𝑠 ∗ 𝑃𝑎𝑠𝑠 𝑇𝑖𝑚𝑒 + 𝑆𝑝𝑜𝑡 𝑇𝑖𝑚𝑒
7. Determine the cycle time for a truck
𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒 = 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 + 𝑇𝑟𝑎𝑣𝑒𝑙 𝑡𝑜 𝐷𝑢𝑚𝑝 𝑆𝑖𝑡𝑒 + 𝐷𝑢𝑚𝑝𝑖𝑛𝑔 + 𝑇𝑟𝑎𝑣𝑒𝑙 𝑡𝑜 𝑃𝑖𝑡
8. Determine the number of shovels required
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑆ℎ𝑜𝑣𝑒𝑙𝑠 = 𝑅𝑂𝑈𝑁𝐷𝑈𝑃 (𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑇𝑟𝑢𝑐𝑘𝑠 ∗ 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 𝑇𝑖𝑚𝑒)
9. Compare different truck and shovel pairings by cost and efficiency
𝑆ℎ𝑜𝑣𝑒𝑙 𝐸𝑓𝑓𝑖𝑐𝑖𝑒𝑛𝑐𝑦 =𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑇𝑟𝑢𝑐𝑘𝑠 ∗ 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 𝑇𝑖𝑚𝑒
𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒 ∗ 𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑆ℎ𝑜𝑣𝑒𝑙𝑠
This model was used with recommendations from Andrew Moebus, sales support staff of Toromont.
10.1.1 Daily Ore and Waste Production
The product of a daily milling rate and expected strip ratio is the expected daily waste production rate as
shown:
𝐷𝑎𝑖𝑙𝑦 𝑊𝑎𝑠𝑡𝑒 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 = 𝐷𝑎𝑖𝑙𝑦 𝑂𝑟𝑒 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 ∗ 𝑆𝑡𝑟𝑖𝑝 𝑅𝑎𝑡𝑖𝑜
For an open pit mine in the arctic, it was assumed 5 days are lost to holidays and other work
interruptions every year [11]. Targeting a yearly milling rate of 3.2 Mt of ore per year and assuming 360
41
effective working days per year, about 8889 tpd of ore can be expected per day. To determine the total
size of the fleet, the conditions with the highest production rate. Based a production-balanced mine
schedule with stockpiles (refer to section “Production Schedule”), strip ratio reaches approximately
7.43, producing an expected waste production of about 65000 tpd of waste, as in Table 10-2.
Table 10-2 A summary of mining rates near the end of mine life.
Material Movement Units
Strip Ratio Waste/ore 7.43
Ore Per Day Tonnes 8889
Waste Per Day Tonnes 65132
10.1.2 Daily Productive Hours
Assume a number of working hours scheduled per day; the daily productive hours can be estimated
based on an estimated efficiency and time used for shift changes and breaks.
It was assumed that a schedule can designed for a 24 hour day, with 4 hours lost to breaks and shift
changes [11]. Of the remaining 20 workings hours, assume 90% efficient use [11], resulting in 18 hours
of productivity per day, as summarized in Table 10-3.
Table 10-3 A summary of net productive hours calculation.
Scheduling and Availability
Daily Scheduled Hours hrs 24
Shift changes, lunches and Breaks hrs 4
Gross Scheduled hours hrs 20
Efficiency % 90
Daily Productive Hours hrs 18
𝐷𝑎𝑖𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑣𝑒 𝐻𝑜𝑢𝑟𝑠 = (𝐷𝑎𝑖𝑙𝑦 𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒𝑑 𝐻𝑜𝑢𝑟𝑠 − 𝐵𝑟𝑒𝑎𝑘𝑠) ∗ 𝐸𝑓𝑓𝑖𝑐𝑖𝑒𝑛𝑐𝑦
10.1.3 Required Hourly Production Rate
The product of targeted mining rates and the fraction of the working day available is the required daily
productivity. The product of the previously calculated mining rates and daily productive hours results in
a required production rate of 494 t/hr of ore and 3160 t/hr of waste.
𝐻𝑜𝑢𝑟𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 = 𝐷𝑎𝑖𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 ∗𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑣𝑒 𝐻𝑜𝑢𝑟𝑠
𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒𝑑 𝐻𝑜𝑢𝑟𝑠
42
10.1.4 Potential Truck and Shovel Models
Out of the different types of loading vehicles, the front shovel was recommended for its greater
versatility than excavators and rope shovels [11]. Variety in the haulage fleet would be limited to only
one model of shovel and one model of truck, to avoid issues with maintenance and inventory of spare
parts.
Using the Toromont Pass Match chart shown in Figure 10-1, for a mine between 8000 to 10000 tpd
milled, the CAT 785 truck is recommended, in conjunction with shovel models from 6030FS to 6050FS.
To account for variations in mine geometry and strip ratios, trucks from 777G, 785D, and 789D and each
model of shovel would be considered for analysis.
Figure 10-1 Toromont pass match chart for determining truck model based on milling rate and recommended shovel models based on truck model [12]
10.1.5 Properties of Trucking Routes
The properties of the trucking routes (the distance, rolling resistance, slopes of roads) site was
determined from measurements of the of the site layout map shown in Appendix Section 16.13. Sloped
distances were determined from the horizontal distances on the map and slope grades using
trigonometry.
𝑆𝑙𝑜𝑝𝑒 𝐷𝑖𝑠𝑡𝑎𝑛𝑐𝑒 =𝑀𝑎𝑝 𝐷𝑖𝑠𝑡𝑎𝑛𝑐𝑒
cos(𝐺𝑟𝑎𝑑𝑒)
43
The grades of the dump and pit ramps were determined in the Final Pit Design Section. Rolling
resistance (RR) values were adopted from the Toromont’s sample values [12]. Compiled properties of
road segments are in shown in Table 10-4.
Table 10-4 Distances, grades and rolling resistances involved in the haulage routes for ore and waste.
Segment Length Grade RR
Units m % %
Load area 100 0 4
Ramp 3298 10 3
Crusher Road 1713 0 3
Crusher Ramp 100 10 3
Crusher Pad 50 0 3
Dump Road 702 0 3
Dump Ramp 3224 10 3
Dump Run 423 0 3
10.1.6 Time Spent on Travelling to and from Dump and Mill
From the geometry of the site layout (the distance, rolling resistance, slopes of roads), assumed speed
limits, and the truck models chosen for analysis, appropriate truck speeds for each road segment could
be determined using the appropriate rimpull and retarding curves (Appendix Section 16.3). Rimpull
curves were used for uphill travel, retardation curves were used for downhill travel.
After taking appropriate speed estimates from the charts, a speed limit of 35 km/h was applied to
sections in the pit or on a slope and a 50 km/h limit was applied mine wide [12]. Knowing the speed and
distance of each segment, the time required to traverse each segment could also be calculated. The
calculated speeds and time to traverse each segment are shown in Appendix Section 16.4.
10.1.7 Loading Time
Loading time depends on the number of passes required by a certain shovel to fill a truck. The number
of passes required would depend on the ratio of bucket capacity to truck capacity.
The capacity of a truck depends on its model. Considering trucks 777, 785, and 789, bucket capacities
range from 100 to 200 tons.
The capacity of the bucket and the number of passes required can be calculated from its volume using
typical fill factors and loose densities of ore and waste [12]. The total time spent loading would be the
sum of every pass and spotting times. Typical pass and spot times were recommended by Toromont
Industries Ltd., 2016. Calculated values are shown in Appendix Section 16.13.
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑃𝑎𝑠𝑠𝑒𝑠 =𝑇𝑟𝑢𝑐𝑘 𝐶𝑎𝑝𝑎𝑐𝑖𝑡𝑦
𝑆ℎ𝑜𝑣𝑒𝑙 𝑉𝑜𝑙𝑢𝑚𝑒 ∗ 𝐿𝑜𝑜𝑠𝑒 𝐷𝑒𝑛𝑠𝑖𝑡𝑦 ∗ 𝐹𝑖𝑙𝑙 𝐹𝑎𝑐𝑡𝑜𝑟
44
𝐿𝑜𝑎𝑑 𝑇𝑖𝑚𝑒 = 𝐹𝑖𝑟𝑠𝑡 𝑃𝑎𝑠𝑠 + 𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑃𝑎𝑠𝑠𝑒𝑠 ∗ 𝑃𝑎𝑠𝑠 𝑇𝑖𝑚𝑒 + 𝑆𝑝𝑜𝑡 𝑇𝑖𝑚𝑒
10.1.8 Truck Cycle Time
Having previously determined the time spent in travel, the total cycle time required for a truck to have
hauled and dumped its load and returned for another load is the sum of loading time, travel time to the
dump/crusher, dumping time, and return time to the bottom of the pit.
𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒 = 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 + 𝑇𝑟𝑎𝑣𝑒𝑙 𝑡𝑜 𝐷𝑢𝑚𝑝 𝑆𝑖𝑡𝑒 + 𝐷𝑢𝑚𝑝𝑖𝑛𝑔 + 𝑇𝑟𝑎𝑣𝑒𝑙 𝑡𝑜 𝑃𝑖𝑡
10.1.8.1 Number of Required Trucks
The number of required trucks is the ratio of the hourly production rate to the productivity of a single
truck. The productivity of a single truck is a ratio of its actual capacity, which is the product of the
number of shovel passes used for loading and the capacity of the shovel bucket, to its total cycle time.
The results are shown in Appendix Section 16.17.
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑇𝑟𝑢𝑐𝑘𝑠 =𝐻𝑜𝑢𝑟𝑙𝑦 𝑃𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑃𝑎𝑠𝑠𝑒𝑠 ∗ 𝑆ℎ𝑜𝑣𝑒𝑙 𝐶𝑎𝑝𝑎𝑐𝑖𝑡𝑦
10.1.9 Number of Required Shovels
Number of required shovels can be calculated from rounding up the ratio of the cycle time for a truck to
the sum of loading times for each truck in the fleet. The results are shown in Appendix Section 16.17.
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑆ℎ𝑜𝑣𝑒𝑙𝑠 = 𝑅𝑂𝑈𝑁𝐷𝑈𝑃 (𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒
𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑇𝑟𝑢𝑐𝑘𝑠 ∗ 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 𝑇𝑖𝑚𝑒)
10.1.9.1 Comparison Metrics
The combinations between trucks and shovels can be compared by their capital cost and shovel
efficiencies. Typical capital costs were taken from the Relative Pricing chart [13]. Shovel efficiencies
were calculated as the ratio of time the shovel spends loading to the productive hours available for the
shovel to work. Graphs depicting capital cost and efficiencies of each truck and shovel pairing are shown
in Figure 10-2 and Figure 10-3.
𝑆ℎ𝑜𝑣𝑒𝑙 𝐸𝑓𝑓𝑖𝑐𝑖𝑒𝑛𝑐𝑦 =𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑇𝑟𝑢𝑐𝑘𝑠 ∗ 𝐿𝑜𝑎𝑑𝑖𝑛𝑔 𝑇𝑖𝑚𝑒
𝑇𝑟𝑢𝑐𝑘 𝐶𝑦𝑐𝑙𝑒 𝑇𝑖𝑚𝑒 ∗ 𝑁𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑆ℎ𝑜𝑣𝑒𝑙𝑠
45
Figure 10-2 Capital cost associated with each shovel truck pairing
Figure 10-3 Efficiency of each shovel truck pairing
Based on these performance metrics, the pairing of the CAT 785G truck and the 6040 shovel were
selected due to the high efficiency and low cost of this pairing, with the required number of each
detailed in Table 10-5. Of concern is the reliance on only one shovel, but capital would be allocated to
ensure proper maintenance and replacement parts if needed.
$60,000,000
$65,000,000
$70,000,000
$75,000,000
$80,000,000
$85,000,000
$90,000,000
777G 785D 789D
Truck Model
Capital Cost of Shovel-Truck Pairings
6015 6018 6030 6040 6050 6060
30%
40%
50%
60%
70%
80%
90%
100%
777G 785D 789D
Truck Model
Efficiencies of Shovel-Truck Pairings
6015 6018 6030 6040 6050 6060
46
Table 10-5 Number of trucks and shovel s expected throughout the mine life
-2 -1 1 to 17 18
Trucks 7 13 21 8
Shovels 1 1 1 1
10.1.10 Additional Equipment and Support Fleet
Based on recommendations from Moebus and reference to other technical reports, additional
equipment requirements are detailed in Table 10-6. CAT models are used as reference for cost at this
time and unspecified models would be determined in later studies.
Table 10-6 Number of additional and support equipment expected
2 Front End Loaders CAT 994F 18 m3
4 Track Dozer CAT D9T 13.5 m3
2 Wheel Dozer CAT 854K 7.9 m3
3 Motor Grader CAT 24M 24’ blade
1 Water Truck --- 10000 gal
1 Articulated Truck CAT 735B 19.7 m3
1 Vibratory Compactor CAT CS-64 112 kW
1 Tool Carrier CAT IT 38H 2.5 m3
6 Diesel Drill --- 4.5’’ to 8.5’’
1 Secondary Drill --- 4.5’’ to 5.5’’
10.2 Benchmarking
10.2.1 ARCTIC (NovaCopper Inc.)
NovaCopper Inc. is a company operating in Alaska, USA, with several polymetallic properties. Of note is
their planned ARCTIC mine in Alaska [14], which like Grum is a polymetallic open pit mine with a similar
milling rate of 10000 tpd, about 12% larger than 8900 tpd for the Grum project. A comparison of the
chosen fleet provides reasonable benchmarking. A comparison of equipment is listed in Table 10-7
below.
Table 10-7 A comparison of preliminary equipment fleets of Grum and NovaCopper’s ARCTIC
ARCTIC Grum
Capacity Number Capacity Number
Shovels 11.0 m3 4 22.0 m3 1
Haul Trucks 100 t 24 150 t 21
47
NovaCopper’s ARCTIC uses 4 times to number of shovels, with a total shovel capacity twice that of
Grum, but only 12% higher milling rate. Also, the trucks of ARCTIC have only two thirds the capacity of
the trucks selected for Grum. This suggests that ARCTIC has shorter haulage routes, allowing trucks to
return quicker and more often to utilize more shovels. This suggests that Grum may have longer haul
lengths and could not achieve the greater efficiency and redundancy from using smaller equipment.
These advantages more be possible for the Grum pit if the site layout and ramp exit are better
optimized in the future; currently, the ramp exits in a direction opposite to that of the mill, requiring an
extra 3 km of travel.
10.3 Meadowbank (Agnico-Eagle Mines Ltd.)
Meadowbank [15] is a gold mine located in Nunavut, Canada, which has a milling rate of 10100 tpd,
which is 13% larger than the 8900 tpd of Grum. A comparison of equipment is shown in Table 10-8.
Table 10-8 A comparison of loading and haulage fleets between Grum and Agnico Eagle’s Meadowbank
Meadowbank Grum
Capacity Number Capacity Number
Shovels 15.0 m3 3 22.0 m3 1
Haul Trucks 150 t 11 150 t 21
100 t 8
Similar conclusions from ARCTIC can be drawn for the differences between Grum and Meadowbank;
Meadowbank likely has shorter haulage routes, which would be an objective for later Grum designs.
11 Environmental and Social Impact Assessment
11.1 Required Legal Documents
Along with this very brief and preliminary summary, many documents are required to determine and
describe the environmental and social impacts of the project. A full federal Environmental Impact
Assessment (EIA) or Environmental and Social Impact Assessment (ESIA) is required by law by the
Canadian Environmental Assessment Agency as well as the Yukon Environmental and Socio-economic
Assessment required by the Yukon Environmental and Socio-economic Assessment Board. These
documents must be filed and approved before construction commences. Table 11-1 describes the legal
steps that must be followed to obtain approval to mine in the area [16]. It outlines the authorization,
which act it is from and who requires it for each activity. Section 11.4 also outlines a preliminary
potential Impact Benefit Agreement (IBA) with the local Aboriginal Communities.
48
Table 11-1 Permits for various Mine Activities
11.2 Valued Ecosystem Components
Valued Ecosystem Components (VECs) are defined as broad components of the biophysical and human
environments, which, if altered by the project, would be of concern to regulators, participating
Aboriginal Groups, resource managers, scientists, and the public. The purpose of identifying the VECs is
not to be all inclusive, recognizing the practical impossibility of analyzing everything, but to look at
potential project effects on representative components.
11.2.1 Atmospheric Systems
11.2.1.1 Air Quality
Air quality can be subdivided into three key indicators: ambient air quality concentrations, particulate
matter deposition, and greenhouse gases in the atmosphere. Ambient air quality deals with the
potential effects of air emissions on the environment. Mining activities generate air emissions due to
fuel consumption, erosion, and material transfer caused by ore processing and traffic on unpaved roads.
Table 16-23, in Appendix Section 16.18, outlines the ambient air quality and particulate matter
standards for Yukon Territory [17]. The standards must be met to ensure the safety and sustainability of
the area and its components. Major air pollutants such as sulphur dioxide, nitrogen dioxide, carbon
monoxide and particulate matter in the study area must be monitored.
49
11.2.1.2 Noise
Noise levels are important to individuals and wildlife for several reasons such as sleep disturbance and
annoyance. Both the Fannin’s Sheep and Boreal Woodland Caribou are sensitive to noise, so special
consideration is required for the eastern side of the site. More noise is expected during the construction
phase; the site will be run on diesel power generators and more vehicle traffic is expected. Mitigation
strategies are outlined in Section 11.3. Baseline studies on the A-weighted noise levels (dBA) are
required. Yukon does not currently have any published noise guidelines or regulation [18].
11.2.2 Water Systems
The mine site is located within the Yukon River drainage basin, Figure 11-1. The site is upstream of
Vangorda Creek, which empties into the Pelly River approximately twelve kilometers from the mine site.
Within four kilometers of the Pelly River, the Vangorda Creek provides some seasonal recreational
fishing. However, there are better choices in the area, including the Pelly [19]. A water treatment plant
will be located on site to treat the effluent water from pumping and processing. By managing the quality
of released water, the nearby water systems and aquatic habitat will be preserved. An adaptive
treatment plant that allows for fluctuating discharge rates will ensure the effluent quality does not
exceed limits outlined in Table 16-24 in Appendix Section 16.18 [17].
Figure 11-1 Yukon Drainage Basins [20]
11.2.2.1 Pelly River
Pelly River contains various types of fish including:
Rainbow trout
Kokanee salmon
Arctic char
Jackfish or pike
Lake trout
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Arctic Grayling [19]
The most likely catch is the Salmon and Arctic Grayling. Most of the above fish are found in much
greater abundance in the surrounding lakes leading to the Pelly River. The indicator species in the area is
the Rainbow Trout. The Pelly River fish are not considered at risk, endangered or of special concern. The
river is also used as canoe and kayak routes for tourists [21].
11.2.2.2 Vangorda Creek
The sections of Vangorda Creek affected by the project are not fish bearing. Several kilometers
downstream, there is a natural waterfall that prevents fish passage to this section of the creek. Two
testing points on Vangorda Creek are planned to monitor for possible effluent caused toxins [22].
11.2.3 Terrestrial Environment
In 1969, a forest fire cleared most of the old growth boreal forest in the faro area and Cyprus Anvil’s
Faro Mine cleared the adjacent property by 1998. A baseline study is necessary to determine if any rare
plant species exist on or close to the mine site. Various lichen species should be mapped to determine
wintering grounds of the woodland caribou.
11.2.3.1 Vegetation Communities
In 1969, a forest fire cleared most of the old growth boreal forest in the faro area and Cyprus Anvil’s
Faro Mine cleared the adjacent property by 1998. A baseline study is necessary to determine if any rare
plant species exist on or close to the mine site. Various lichen species should be mapped to determine
wintering grounds of the woodland caribou.
11.2.3.2 Wildlife
The mine is not located within a game management subzone. Hunting by non-aboriginals in this
area is limited to special guide license [21].
Three large prey species live within the Grum project area; moose, woodland caribou and
Fannin’s sheep. The largest population of wildlife in the area consists of moose. There are 1 to 10 moose
per square kilometer on the immediate Grum site and adjacent property. Moose are not considered of
special concern and can adapt easily to noises and habitat loss. Moose are most susceptible to vehicle
collisions and increased non-aboriginal hunting [23]. Boreal populations of woodland caribou in the
Yukon are not considered of special concern. Measures will be taken to avoid the habitat of the
woodland caribou because they are a sensitive species. Revegetation during remediation will focus on
jack pine forests which are lichen bearing. Lichen is the caribou’s primary food source in the winter.
Every fall sheep gather in the mountains and cliff sides around Faro to spend the winter. The cliff sides
provide grasses that remain from the summer and a secure home for the sheep [24]. Fannin’s sheep are
an important tourist attraction for the Faro Community. The sheep, like caribou are sensitive to habitat
changes.
51
Wolves, coyotes, wolverines, lynx and bears prey on the moose, caribou and sheep in the Faro
Area. Coyotes and bears could be an issue on the future mine site if drawn by food waste scents. The
predators tend to follow the big game and prey mostly on the smaller sheep and young moose and
caribou. More information is required regarding predator habitat in the Faro region.
11.2.4 Natural Heritage System
Heritage resource sites commonly found in the Yukon include cabins, tent frames, brush camps, caches,
traps and snares, fire-cracked rock, fish camps, watercraft, stone adze-cut stumps, game drives and
surrounds, trails, and graves. The Faro Interpretive Centre is open May to September and accounts for
0.4% of Faro’s economy. The center highlights the natural heritage in the Faro region and organizes
various activities [25]. The Dena Cho Trail retraces the 80-km route between Ross River and Faro
originally used by gold prospectors [26].
11.2.5 Socio-Economic Factors
A full Social Assessment is required, and will outline local economy, traditional pursuits, aboriginal
community, health, heritage resources and physical infrastructure, regional economy and mining
industry. The two groups to focus on are the Ross River Dena Council of the Kaska traditional territory,
and the community of Faro.
11.2.5.1 Faro
Faro is a town with a population of 367, located 12 km downstream of the potential Grum Pit.
Although Faro was created as a community to house workers and provide services to the Cyprus Anvil
Mine, mining is no longer the base of the community's economy. In 1996, when the Anvil Range mine
was still producing, well over 50 percent of the community's workforce was employed in mining. Other
industries provided services to the mine, such as transportation of minerals or delivery of goods. The
people of Faro are now focused on practical economic development opportunities to guide the
community. Important issues include residential development, community facilities, recreation, the
environment, infrastructure and social well-being. Because Faro established a wide variety of
community and recreational facilities when the Faro Mine was in operation, there is a high level of
services and well-developed community and recreational infrastructure. A good housing stock, along
with extensive community services, can help support economic growth. The people of Faro are wary of
new mining projects within close proximity to the mine due to the abandoned Faro Mine and the
environmental issues it has caused. It is important to assess potential effects on employment and
income as it is a prominent factor in determining community benefits arising from the project. Changes
in employment and income can affect the wellbeing of individuals, families, and communities in the
area. Public consultations will occur during each stage of the design and the Environmental Assessment.
Benny Resource Group will also invest a percentage of monthly profits in the community to support post
mining prosperity.
52
11.2.5.2 Ross River Dena Council
The land in which the Grum project is located is within the Kaska Nation and is governed by the
Ross River Dena Council. From an Aboriginal perspective, maintaining cultural identity requires the use
of land for harvesting traditional resources, opportunities to transfer traditional knowledge and skills,
and speaking the local language. The Kaska people are settled in Ross River, upstream of the Grum site.
Since the closure and lack of remediation of the Faro Mine, traditional hunting and trapping has ceased
in the project area [27]. The Ross River Dena will be appropriately compensated for lost land and
employed in the Grum Mine. It is important not to infringe on aboriginal rights. A preliminary Impact
Benefit Agreement is presented in Section 11.4.
11.3 Assessment of Impacts
It is important to identify the potential impacts the project may have on the stakeholders and VECs.
Understanding the consequences helps to manage and mitigate the causes. Table 16-28 through Table
16-34 in Appendix Section 16.18 establish the preliminary impacts, associated risks and preliminary
mitigation measures. The significance of environmental impacts were determined for effects after
application of appropriate mitigation measures, and was evaluated on the basis of the criteria described
in Table 16-25 through Table 16-27, Appendix Section 16.18 [28].
Table 11-2 below, provides a summary of the key impacts, the cause of the impact, and the mitigation
strategies in place to reduce the risk negative impacts.
Table 11-2 Summary of Key Impacts, Causes, and Mitigation Strategies
Valuable Ecosystem Component System
Cause Effect Mitigation
Socioeconomic
Lack of transparency, dishonesty, cumulative and/or combined causes stated below.
Inability to obtain permits, degradation of public reputation, strikes, blockades and damage to property.
Say what you mean and do what you say. Implement mitigation measures stated below. Consult with the community and remain transparent, honest and approachable.
Mine site accidents, vehicle accidents, natural disasters
Injury to employees Dry mine site, proper preventative and reactive training.
Visual impact of mine structures, degradation of water quality,
Decreased tourism and recreation leading to reduced economic prosperity
Compensation and contribution to community amenities. Vegetated barrier, natural slopes at closure.
See all below See all below See all below
Water Stream diversion/area loss Contaminated surface water-
sediment build up Divert minimum, consider different site layout. Monitor water quality.
53
Acid mine drainage Contaminated groundwater/surface water
Divert water to catchment pond and treat, line TSF and WRD. Monitor water quality upstream and downstream. Place pipelines from Mill to TSF and TSF to water treatment plant so they do not cross water systems. Remain within one watershed (Vangorda Creek/Pelly River).
Pumping Reduced local and regional groundwater affecting aquatic species.
Release treated water as required. Recycle tailings water for processing.
Cultural Heritage Restriction of project area and clearing area for the site
Interference to traditional pursuits, damage to local heritage
Report found artefacts, consult with community to determine options, if any.
Terrestrial
Clearing area for site Habitat loss Provide 150 m buffer around streams. Condense site as much as possible. Reclaim area as a productive ecosystem. Monitor wildlife
Acid mine drainage Contaminated habitat See above Water Systems mitigation measures. Monitor wildlife
Atmospheric
Vehicles, blasting, mill, generators
Noise: Negative effects on animals and people.
Increased greenhouse gas emissions
Vegetation barrier, use electrical power after construction, blast during afternoon shift change
Vehicles, blasting Increased dust in the air: reduced air quality
Water sprayers to reduce airborne particulate matter
Monitor air quality.
The focus of the summary is on the impact on the socio-economical system. The Faro community and
the Kaska people will be affected in changes to the environment. It is important to Benny Resource
Group to prevent another Faro Mine disaster and foster mutual respect with the communities. Many of
the mitigation strategies involve site characteristics, such as footprint considerations and inclusion of a
water treatment facility. The site layout, as discussed in Section 8, was designed to reflect such
considerations. Consultation to present and discuss designs is also a pertinent way to diffuse potential
social issues, and obtain valuable input from the community. It is also important to provide
compensation when negative effects are foreseen that cannot be avoided. Monitoring is planned to take
place throughout all phases of mine life including post-closure. Proactively monitoring and having
procedures in place, to adapt and prevent significant negative events is incorporated into the design of
the project.
11.4 Impact Benefit Agreement
The Kaska and the natural resource development sector is committed to the reclamation of impacted
sites. Consideration must be given to compensation for impacts incurred. An important issue is how to
determine fair compensation and how to implement a compensation policy. The Kaska expect
compensation plans to be applied to themselves, wildlife and habitat, Kaska Land Stewards, and
54
Trappers. In addition, there is expected resource development and federal regulatory compensation
plans [29].
In order to properly compensate the Kaska Nation for the future mine impacts, payment, in addition to
jobs and training, should be provided. The agreement includes a payment plan, training and
employment opportunities, and environmental commitments for the Kaska First Nations. Until the mine
goes into production, the Kaska would receive fixed bonus payments at different points during the
permitting process. Once the mine is in operation, the First Nations would continue to receive fixed
annual payments until Benny Resource Group has paid off the cost of building the mine. After that, the
Kaska would receive a percentage of the mine’s monthly profits [30]. Those payments would be divided
among the Nations according to the Kaska Collaboration Agreement, which sets out guidelines for
sharing the benefits of resource development among the communities [31]. Money will also be invested
in first nation’s communities and shares will be offered to the community to directly involve the band so
positive outcomes for the company mean positive outcomes or the band.
Benny Resource Group will provide Ross River Dena a voice in the remediation planning, and
consultation will occur at each stage of design. Careers in waste, food service, custodial services, and
mine labour will be made available for the Native people first with few restrictions and thorough
training. There will also be career development workshops and information sessions with regional high
schools. Contracts will be made with aboriginal construction companies to make certain that local
communities gain experience and thrive. The target for Aboriginal workforce is 20%. Post-secondary
equivalent training for more technical positions in the mine will also be available to members of the Faro
and Kaska communities. During the early phases, Traditional Ecological Knowledge consultants can aid in
baseline studies, wildlife patterns, obtaining information on traditional cultural heritage and take part in
monitoring programs [32]. Remediation strategies will be greatly affected by the First Nation
community.
12 Mine Closure
12.1 Introduction
This section will address potential environmental issues, regulations, required permits, social
involvement, and information in regards to environmental sustainability during operation, and closure.
This section will also cover the following, as in compliance with the Yukon Mine Reclamation and
Closure Policy:
Some baseline studies for important environmental components with their environmental monitoring programs that plan to be implemented
Reclamation objectives
Stabilization of structures and workings
Areas of re-vegetation where practicable
Cost estimate for the reclamation process
55
12.2 Regulatory Requirements
BRG will be responsible for the reclamation process in accordance with the established legislative
framework set by the Yukon government. The reclamation process will be integrated during the
planning, development and operating phases of the mine. Before advancing with the development, the
Yukon government will need to approve the reclamation and closure plan. A Certificate of Closure, as
stated in section 137 of the Quartz Mining Act, will be issued by the Yukon government, given it is in
compliance with the prescribed conditions set by the government [33]. It is mandated that BRG seek
consultation of government, First Nation, and local community member during the reclamation and
closure planning, which is also part of the provision of the Certificate of Closure.
12.2.1 Permits
In order to operate the Grum Pit successfully, obtaining permits and licenses will be necessary. The
process of receiving permits can be between 100-130 days after submission. This time frame can
include: appropriately addressing concerns of all parties affected by the operation such as the
government agencies; the First Nations; and local stakeholders. Table 12-1 displays the permits required
for all closure activities.
Table 12-1 Permits Required for Mine Closure
12.3 Environmental Studies
For successful mine closure, the Yukon Environmental and Socio-economic Assessment (YESA) will be
required, which outlines and identifies the potential environmental and socio-economic effects. YESA
will be necessary to obtain the permits needed to launch development. The process will entail a
systematic approach to identifying potential issues by conducting series of baseline studies of the
environment, with the main focus on the Valued Ecosystem Components (VECs).
56
12.3.1 Environmental Baseline Studies
Collecting the environmental baseline data is a key element of the YESA for the Grum Pit. The data will
include: climate; air quality; hydrology; sediment and water quality; and flora and fauna. The data will
be collected for the whole year, which will aid in observing the seasonal variation of these different
environmental disciplines. YESA will be submitted to the Yukon Environmental and Socio-economic
Assessment Board (YESAB), and will be important in evaluating the potential impacts from the project so
alternative actions and mitigation measures be constructed. A rough baseline study will also be
conducted for site layout, but it will be based be on information already available, and not carried out by
BRG. Table 12-2 summarizes the environmental baseline data BRG need to collect and study.
Table 12-2 Environmental Baseline Studies
Category Program Elements
Climate Meteorology stations, providing up-to-date information
Air Quality/Noise Total Particulate (Dust fall) samples
Hydrology Water quality measurements and evaluation
Fisheries and flora Monitoring population of distinct fish species and flora communities
Fauna Monitoring wildlife communities and population
Hydrogeological studies
Determining the distribution of groundwater and its quality, depth and other properties. Study of watersheds.
Geophysical studies Monitoring seismic activities
Socioeconomic studies
Background studies on First Nations, communities and other stakeholders
Archeological studies Determining scientific interest archeological sites and First Nation heritage sites.
Sediments and Water quality
Observing existing chemicals and pH level
12.4 Objectives and Environmental Issues
The closure plans and design process will incorporate opinions of many stakeholders including the
indigenous communities. The main targets of the closure phase are:
1. To safeguard the health and welfare of all humans, plants and animals 2. To safeguard and return the environment to as close as possible to its pre-disturbed state 3. To restore and re-contour the mine site for both aesthetic and land use purposes, especially the
tourist sites 4. To assist in socio-economic benefits and opportunities 5. To implementing cost-effective measures to mitigate long-term risks
In order to achieve these targets, potential environmental will be assessed and analyzed. The main
concern is the acid rock drainage. Therefore, appropriate measures will have to be taken to address, and
prevent this concern from occurring.
57
12.4.1 Acid Mine Generation
Acid Mine Drainage (AMD) forms when the sulphide minerals, (e.g. in lead and zinc ores), oxidize. These
oxidized sulphide minerals are later discharged into the environment. The leaching of these materials
can occur in various ways like along the pit slopes, on the base of the waste rock pile and tailings, and
often occurs during rainfall, where water becomes the medium of transport. Both WRD and the tailings
outputted from the mill and stored in the tailings impoundment will have PAG waste. The occurrence of
AMD will have adverse effect on the immediate environment particularly the aquatic ecosystems, those
near the Vangorda Creek and the watersheds. Acid contact with water bodies will reduce the pH level
and increase the toxicity, posing a threat to the biotic life. The oxidized sulphide minerals may seep into
the ground, contaminating groundwater and negatively affect those that rely on it as a source of
drinking water. AMD will require a long-term remediation plan and it will have adverse effects that last
many years if not dealt appropriately and will be discussed in the following section.
12.5 Environmental Management
Environmental Management Systems (EMS) will be developed for both corporate and specific to the
operation throughout the entire mine life and will be consistent with both ISO 14001 EMS standards and
the standardized framework established by the Yukon government. Upon the commencement of the
operation, a more elaborate and complete environmental management plan will be made to better
reflect the practicality of the operation. The subsection below provides ways in which the waste will be
managed. This is the initial, ostensible closure plan and therefore is subjected to change once mine
begins to operate and if unforeseen circumstances abound. The mine site is expected to have one Waste
Rock Dump (WRD) and one Tailings Storage Facility, seen in Section 8.
12.5.1 Waste Rock Dump
The main aim will be to return the dump site to close as possible to the original state, identical to
surrounding lands, while ensuring that both physical (slope) and chemical (prevention of metal
transport) stability remains robust for long-term basis. The following plans will be implemented for
waste rock dump during closure:
Re-contouring of WRD terraced slopes to appear less steep and to naturally blend in with surrounding landform
Implementing a drainage system to better cope with heavy rainfalls and prevent soil erosion. If stream diversion results due to expansion of WRD, that would be considered when implementing the drainage system
Re-sloping of terraced slopes to establish long-term stability with cover on top
Restoring watersheds affected by both WRD and TSF
A multi-layer, engineered cover system will be placed on the acid-generating tailings impoundment. The
first layer placed directly on top of the tailings will comprise of sand. The sand layer will then be overlain
by coarser material. On top of the coarser material will lay a layer comprised of stockpiled fine-soil
mixed with gravel. The interface between the fine and course layer will be the capillary barrier for
58
controlling percolating water and water migration. The main aim for the fine-soil layer will be to retain
the infiltrated water until the water is recycled back to the atmosphere through evaporation. Another
way to recycle water back to the atmosphere is through transpiration; therefore the top layer will be
fertilized and seeded to re-vegetate. Vegetation on the cover will consists of a diverse mixture of native
plants, which will not only maximize the removal of retained water through evapotranspiration process,
but will also remain stable and resilient to unforeseen alteration in the environment and climatic
fluctuations. These plant species have the inherent ability to quickly adapt to environmental changes.
The gravel mixed with fine-soil, will help in preventing wind erosion by reducing tractive shear stresses.
Figure 12-1 shows the schematic of a cross-section of the cover to be used on the WRD.
Figure 12-1 Schematic cross-section of the cover over WRD
12.5.2 Tailings Dam
A two meter water cover will be used to prevent acid generation in the tailings dam. This cost-effective
method will mitigate AMD on long-term basis because oxygen has a very low diffusion and solubility rate
in water [34]. The main aim will be to prevent cover from drying up or spilling out. Therefore, a
catchment area will be made to ensure that there is adequate inflow to sustain water balance. There will
be a sufficient freeboard with emergency spillway to prevent overtopping during extreme flood events.
Lime will be added to increase the pH level of the waste before the cover is placed to further prevent
the potential of AMD from occurring. In order to maintain the integrity of the cover, the stability of the
dams on long-term basis will be important. This is further discussed under monitoring.
12.5.3 Pit Lake
Since the PAG rock associated with mineralization will be mined out and placed in the tailings pond, and
because no excess material will be left for backfilling; the open pit will be flooded to form an end-pit
lake. The water cover will slow acid generation, until BRG finds other means to permanently prevent it.
The bottom of the pit will contain a layer of highly permeable gravel and crushed limestone to neutralize
acid and prevent leaching. BRG will promote an ice-cover over the lake. The ice-cover will prevent
diffusion of oxygen with column of water below it and therefore prevent ARD [35]. The pit lake will be
integrated into the landscape aesthetically. The pit lake will neither be used for recreational purposes
59
nor will it have any aquatic life until it is deemed safe. Placards and fences will be placed around areas of
the pit lake to prevent human and animal intrusions.
12.6 Site Monitoring
Monitoring various environmental components will be largely contingent to specific requirements
outlined in the permits issued by the government agencies, as was discussed in Section 11, as well as
assessing it relative to the baseline data collected. The following section will briefly cover some of the
main environmental components be monitored. The full monitoring program will be outlined in the
YESA.
12.6.1 Water
Monitoring both surface water and groundwater along different locations to determine if any impacts
have occurred due to the operation. Wells will be drilled for monitoring groundwater. Water quality
parameter, such as pH level, and as well as hydrometric parameters of surface water will be monitored
to observe for any concerning and significant changes.
12.6.2 Air
Air quality will constantly be monitored during the operation. It will also be monitored during
reclamation phase to assess metal-bearing particles that may be airborne. A monitoring plan will be
placed in compliance with Yukon air quality guidelines and standards. This will entail monitoring the
total suspended particulates, metal concentrations, and emissions related to the mining operation.
12.6.3 Acid Mine Drainage
Long-term monitoring of the WRD will include monitoring the covers used and to ensure performance is
acceptable in mitigating infiltration and diffusion of water and oxygen. The monitoring program will
include the following:
Ensuring the covers are preventing the mobilization and release of contaminates
Monitoring water infiltration
Checking signs for potential desiccation and frost cracking (results in high permeability) within compacted soil layers
Monitoring for frost heaving that may result into layers mixture
The following would need to be monitored for the tailings dam:
Ensure that the water cover is no less than two meters
Prevent ice lenses from forming on the surface of tailings. This will require the monitoring of water temperature on top of the tailings to ensure it’s well above freezing point
Monitor abrupt climate change such as heavy rainfall or intense drop in temperature
Monitoring potential piping in dam structures and take appropriate measures to prevent it
The following would need to be monitored for the pit lake:
60
Promoting ice-cover and ensuring that it remains
Physical-chemical and biological monitoring
Monitoring of long-term water balance and water quality
Monitoring for potential leaching
12.7 Community Relations
Along with taking the initiative to conduct baseline studies and undertaking YESA, BRG will display
authentic corporate responsibility by consulting with the local community stakeholders and the First
Nations during mine closure. BRG plans to develop and maintain a strong rapport with First Nations and
other stakeholders by proactively consulting with them throughout the entire mine life and reclamation
process. Under the Yukon First Nation Final Agreement, BRG are obligated to consult and create
transparent relationships with the First Nations, and will respectfully uphold such requirements.
Another part of BRG corporate responsibility includes helping the local community members on socio-
economic grounds, by helping in the following ways:
Create employment opportunities for local community members
Find local community contractors for needed materials during mine closure
Financially assist the education system
Fund various community organizations and hospitals
12.8 Closure Costs
The costs of decommission and reclamation is estimated based on the costs incurred by two zinc-lead
mining companies, Teck and Yukon Zinc Corp., both which are operating in Yukon [36] [37]. The costs for
both companies varied partly due to the terms set by independent contractors. It is important to note
that these costs are estimated and are prone to changes during the operation. The costs summarized
below entail costs related to project closure, the decommissioning of facilities and structures,
reclamation undertakings, and post monitoring. The approximated costs are based on the following
assumptions and premises:
Exclusion of salvage value
Non-discounted estimate
Costs relating to reclamation and decommissioning are based on terms set by third party contractor
Unit rates were obtained from Government of Yukon Third Party Equipment Rental Rates. Shown in Appendix Section 16.19.
Contingencies costs, ranging from 10% to 20%, will be added in addition due to degree of potential risk and level of uncertainty
Costs based on assuming closure phase for over 10 years after end-of-mine-life
Costs for water treatment is based on 3 year period
Note that the costs are subjected to change to better reflect the Grum Pit conditions. The total closure
cost accounting for contingency (average of 15%) and monitoring, along with necessary things like liners
and covers can be anywhere between $7 M to $15 M as summarized in Table 12-3.
61
Table 12-3 Estimated Closure Costs
13 Detailed Economic Analysis
The economic analysis is preliminary in nature and depends on a block model of unknown origin. It is
not known if inferred resources are included in the block model definition. 2012 InfoMine cost models
[38] were used for estimating operating costs, the capital cost of the processing equipment, and freight.
Performance metrics are shown in Table 13-1 and a summary of financial results are shown in Table
13-2. A 15% discount rate was used, based on standard industry practice of 14% discount in the scoping
stage as well as a 1% risk adjustment for arctic mining [39].
Table 13-1 – Performance metrics
At a 15 % discount rate and 30% tax rate
NPV $156,072,026
IRR 21%
Payback Period 4.8 years
Life of Mine 22 years
Highest Sensitivities of NPV
Zinc price
Mining operating cost
+$4.74M/% increase of price
-$3.45M/% increase of costs
Table 13-2 - Summary of financial results
Revenues $ 6,163,300,000
Lead
Zinc
Gold
Silver
$ 1,863,200,000
$ 2,769,200,000
$ 909,600,000
$ 621,300,000
62
Operating Costs $ 2,901,900,000
Mining $ 1,171,500,000
Milling $ 714,200,000
Freight $ 364,700,000
Capital Cost $ 534,200,000
Pre-strip $ 115,500,000
Mining Equipment $ 86,900,000
Processing Equipment $ 108,400,000
Sustaining Capital $ 145,300,000
Closure $ 15,000,000
Taxes $ 911,700,000
Net Cash Flow $ 1,815,500,000
NPV $ 156,100,000
The cash flow model and the production schedule it models, are shown in Figure 13-1. The variation in
revenues are largely due to variation in grades and the recoveries of different host rocks over the mine
life. The largest costs in this design are operating costs and taxes.
63
Figure 13-1 - The production schedule and resulting cash flow model for the current pit design and operation
The sensitivity analysis for this design is shown in Figure 13-2. The current design has its greatest
sensitivities in zinc price and mining operating costs. Every percent increase in zinc price could raise NPV
by $4.74 M, is due to zinc contributing to 45% of total revenues. Every percent increase in mining
operating costs could reduce NPV by $3.45 M; reductions in this area would greatly increase profitability
of the mine and would be explored in future studies.
0
5000000
10000000
15000000
20000000
25000000
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Balanced Production Schedule with Stockpiles
Stockpile Processed Ore Mined and Processed Ore Stockpiled Waste
-$400
-$200
$-
$200
$400
$600
-2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Free
Cas
h F
low
s ($
M)
Years into Production
Cash Flow Model
Revenue Operating Costs Initial Capital Taxes
Pre Strip Sustaining Capital Closure Bond
64
Figure 13-2 - Sensitivities of prices and operating costs.
Estimation methodology of revenues, operating and capital costs, and taxes are detailed in the following
subsections.
13.1 Revenues: $6,163,000,000
Revenue was determined from the payable metal contained in the ore and the price it could be sold for
over a long period of time. Payable metal depends on the grade, the recovery and smelter terms on the
sale of the concentrate.
13.1.1 Price
The price used in this model were the consensus price forecasts by Scotiabank, Bank of America, and
National Bank Financial. The consensus prices are shown in Table 13-3 below. Long term price
predictions made by each bank for each metal are shown in
Table 13-4.
Table 13-3 - The forecast prices used for the model
zinc $ 1.03 /lb gold $ 1303 /toz
lead $ 0.97 /lb silver $ 18.93 /toz
$(400)M
$(200)M
$M
$200M
$400M
$600M
$800M
-150% -100% -50% 0% 50% 100% 150%Ch
ange
in N
PV
($
M)
Axis Title
NPV Sensitivity to Prices and Opex
Pb Price ($/lb) Zn Price ($/lb) Au Price ($/toz)
Ag Price ($/toz) Mining Cost ($/ton mined) Processing Cost ($/ton ore)
65
Table 13-4 - The long term price forecasts and the average, consensus price from three banks
Zn ($/lb) Pb ($/lb) Au ($/toz) Ag ($/toz)
Scotiabank $ 1.00 $ 0.95 $ 1,200 $ 17.50
Bank of America $ 1.06 $ 0.98 $ 1,358 $ 19.80
National Bank $ - $ - $ 1,350 $ 19.50
Average $ 1.03 $ 0.97 $ 1,303 $ 18.93
13.1.2 Variable Grades and Contained Metal over LOM
Contained metal was determined on an annual basis, based on the annual production tonnages and
their associated grades, in both the mill throughout and stockpile. The variability in the grades results in
variability of the contained metal for processing. The changing average annual grades are shown in
Figure 13-3 and Figure 13-4. A graph of the annual contained metal processed is shown in Appendix
16.20.
Figure 13-3 - The average annual Pb and Zn grades over the LOM
0
1
2
3
4
5
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18
Gra
de
(%)
Years into Production
Pb and Zn Grades over LOM
PB ZN
0
0.2
0.4
0.6
0.8
0
10
20
30
40
50
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18
Au
Gra
de
(g/t
)
Ag
Gra
de
(%)
Years into Production
Ag and Au Grades over LOM
AG AU
66
Figure 13-4 - The average annual Au and Ag grades over the LOM
13.1.3 Variable Rock Type, Recoveries, and Recoverable Metal over LOM
A key factor considered in the metal production model was the variation of the recoveries over the mine
life, which affects the recoverable metal sent to the mill or kept in stockpile. Recovery varies with
different rock types and the rock type mined changes over the course of production. It is assumed that
the recoveries of each ore type, provided from the Whittle model, are accurate enough for this
economic model. These recoveries are shown in Table 13-5. The annual recoveries are shown in Figure
13-5 and Figure 13-6.
Table 13-5 - The recoveries of each metal for each rock type
QRTZ MQRT PYMS BMAS PHYL BAMS GREE
Pb 87% 85% 86% 86% 89% 91% 88%
Zn 87% 85% 86% 86% 89% 91% 88%
Au 58% 61% 65% 65% 61% 63% 63%
Ag 52% 50% 55% 52% 50% 51% 55%
Figure 13-5 - The variation in lead and zinc recoveries over the scheduled mine life
85%
86%
87%
88%
89%
90%
91%
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18
Rec
ove
ry (
%)
Year of Production
PB ZN
50%
52%
54%
56%
58%
60%
62%
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18
Rec
ove
ry (
%)
Year of Production
AG AU
67
Figure 13-6 - The variation in gold and silver recoveries over the scheduled mine life
A limitation of this approach is the incomplete metallurgical data. It is unknown which processing
method produced these recoveries or what effects from mixing ore types have on recovery. Until more
data it is acquired, this model will assume that these recoveries can be achieved regardless of processing
method and that mixing ores are inconsequential. These assumptions may be incorrect and it is
recommended that more extensive metallurgical testing be performed. Detailed tables of recoverable
metal tonnages are shown in Appendix Section 16.21.
13.1.4 Smelter Terms
Appropriate smelter terms are difficult to benchmark due to the polymetallic nature and poorly
understood metallurgy of the ore. Smelter terms would depend on the grade of each metal in each
concentrate, presence of deleterious elements, and the impact of the ore chemistry on smelter
recovery.
It was assumed that 60% of the lead and zinc concentrates to metal by weight [40], but it is unknown
what proportion and grade of gold and silver would be reporting to each of these concentrates.
Deductions for gold and silver were not considered for this reason.
Without knowledge on the expected metallurgical behaviour of the ore, smelter terms used in this
model are the medians of typical smelter term values [40]. The smelter contract terms used in this
model are listed in Table 13-6.
Table 13-6 - Smelter terms used, adapted from Prices and Revenues [40]
Pb Zn Au Ag
Payables 95% 95% 95% 95%
Deductions 2% 8% --- ---
TC/RC $180/dmt $170/dmt $5.00/oz $0.30/oz
13.2 Operating Costs: $2,902,000,000
13.2.1 Mining Operating Cost: $1,171,500,000 from $3.01/tonne mined
The mining cost was determined using the 2012 InfoMine cost model [38], which models costs per ton
milled based on a milling rate and strip ratio. Having a milling rate of about 8800 tpd and a strip ratio of
around 6, the expected mining unit costs were interpolated from model estimates, as shown in Figure
13-7. This gave the estimate of $20.95/tonne milled. Translating this value into 2016 USD and to a “per
tonne mined” basis results in a value of $2.00/tonne mined. Considering a +50% factor for arctic mines,
the cost $3.01/tonne mined was used.
68
Figure 13-7 - The interpolated unit cost of Grum, at 8800 tpd and a strip ratio of 6.
13.2.2 Processing Cost: $741,200,000 from $14.05/tonne milled
Processing unit costs were determined from InfoMine cost models [38], which determines a cost per
tonne milled, based on a milling rate and number of concentrates produced. Grum has is designed for a
milling rate of 8800 tpd and two concentrates, as shown in Figure 13-8.
Figure 13-8 - Interpolated processing unit cost for two concentrates at a milling rate of 8800 tpd
13.2.3 Freight Cost: $326,700,000 from $74.50/dmt
Freight costs were determined from expected costs incurred transporting concentrate to a chosen
smelter. Chosen smelters, for simplicity, were able to smelt both lead and zinc concentrates. All unit
costs are in relation to dry metric tonnes (dmt), assuming an increase of 8% in mass from water.
Two smelters were considered, one for a land freight case, with the Teck Trail Smelter, located in B.C.,
Canada, and a shipping case, for the Korea Zinc smelter, in Onsan, South Korea. The current design finds
shipping freight preferable and its associated costs were used in this model.
$0.00
$10.00
$20.00
$30.00
$40.00
$50.00
$60.00
$70.00
100 1000 10000 100000Un
it C
ost
($
/to
nn
e m
illed
)
Milling Rate
Mining Unit Cost to Milling Rate and Number of Concentrates
1
2
4
8
Grum
$-
$50.00
$100.00
$150.00
$200.00
$250.00
10 100 1000 10000 100000
Pro
cess
ing
Co
st (
$/t
on
ne
mill
ed)
Milling Rate (tpd)
Processing Cost to Milling Rate and Number of Concentrates
1 2 3 Grum
69
13.2.3.1 Land Freight to Teck Smelter
The Trail smelter is physically the closest lead-zinc smelter to Faro, allowing freight over land. However,
the shortest routes cover at least 5700 km roundtrip and requires passage through highways in the
Rocky Mountains, as shown in fig. The cost associated with such a delivery has poor economics, with an
estimate by Minecost models to cost at least $240.47/dmt by trucking. Train routes were not found at
this point in the study; further research may bring expected costs down.
Figure 13-9 - The route and distance from Faro to Trail [41]
13.2.3.2 Shipping Freight to Korea Zinc Smelter
The Korea Zinc smelter was chosen for being the world’s largest smelter [42], regularly smelting large
quantities of lead and zinc concentrates. The smelter is located near the port of the city Onsan, which
facilitates shipping, as shown in Figure 13-10.
Transport to this smelter requires transport from the mill to a port, where handling and shipping costs
would be incurred. The closest port to Faro is Skagway, which is located 331 miles from Faro, as shown
in Figure 13-11, and has a history of providing transport for metal mines in the Yukon [43]. At this
distance, the trucking cost is estimated to be about $44.69/dmt.
70
Figure 13-10 - The Korea Zinc Onsan smelter, located close to a port [44]
Figure 13-11 - An aerial photograph of the port town Skagway is shown on the left and the shortest route from Faro to Skagway is shown on the right [45]
The median InfoMine cost estimate for shipping cost from western Canada to East Asia is $23.81/dmt. It
was assumed Skagway, a southern Alaskan port, would have similar costs. The handling charge of
Anchorage has a median value of $3.00/dmt; it was assumed that this value was applicable for Skagway
and would be incurred again in the port of Onsan.
71
This totals to about $74.50/dmt, which is less than a third of the costs expected for transportation to
Trail in B.C. These costs were used for the cash flow model.
Additional costs associated with freight are insurance and losses transport. Insurance was assumed to
be 0.5% of the ore’s payable value and losses were assumed to be 0.1% per handling [40]. With three
handlings at the mill, at Skagway, and at Onsan, 0.3% would be lost.
13.3 Capital Cost: $534,200,000
13.3.1 Processing Equipment Capital Cost: $108,400,000
The capital cost of the processing equipment were determined by linear interpolation between
InfoMine cost models for a 8800 tpd milling rate and 2 concentrates, as shown in Figure 13-12.
Figure 13-12 - Interpolated processing capital cost for two concentrates at a milling rate of 8800 tpd
13.3.2 Mining Equipment Capital Cost: $86,900,000
The capital costs of the mining equipment refers to typical costs of Caterpillar vehicles [46] and the costs
of drills from similar mines in the area [14]. The list of equipment models selected and their unit costs
are listed in table below. These costs are incurred over the two years ramping up to full production as
the fleet expands to its max size. Shown in Table 13-7 are the total costs attributed to each type of
equipment.
Table 13-7 – The total capital costs associated with the total mining equipment fleet
Number of Units Capital Cost
Haul Trucks CAT 785D 150 ton 21 $ 63,680,000 Shovels CAT 6040 22 m3 1 $ 8,641,500 Front End Loaders CAT 994F 7.7 m3 1 $ 1,166,000 Track Dozer CAT D9T 13.5 m3 2 $ 2,684,000
$-
$100,000,000
$200,000,000
$300,000,000
$400,000,000
$500,000,000
$600,000,000
$700,000,000
$800,000,000
0 10000 20000 30000 40000 50000 60000 70000 80000 90000
Cap
ital
Co
st (
$)
Milling Rate (tpd)
Capital Cost of Processing Equipment to Milling Rate and Nubmer of Concentrates
1 2 3 Grum
72
Wheel Dozer CAT 834K 7.9 m3 1 $ 1,388,000 Motor Grader CAT 16M 16’ blade 2 $ 2,080,000 Articulated Truck CAT 735B 24 m3 1 $ 782,000 Vibratory Compactor CAT CS-64 112 kW 1 $ 230,000 Tool Carrier CAT IT 38H 2.5 m3 1 $ 379,000 Diesel Drill --- 4.5’’ to 8.5’’ 4 $ 5,060,000 Secondary Drill --- 4.5’’ to 5.5’’ 1 $ 806,000
Total: $ 86,896,000
13.3.3 Capital Pre-strip Cost: $115,500,000
Capital pre-stripping was determined from the mining unit cost applied to the initial waste-only
production. Larry Smith has observed that pre-stripping tends to far exceed budgets; his suggestion for
doubling pre-strip unit costs were used. For a planned 19.186 Mt of pre-strip, at a unit cost of
$6.02/tonne moved, $115,499,000 was allocated for pre-stripping.
13.3.4 Closure Cost: $15,000,000
The closure cost was determined in Section 12. Due to its low weighting compared with other capital
costs, it was treated as a capital cost incurred at the beginning of mine life. This produced a more
conservative estimation in case of an overly low valuation of closure.
13.3.5 Sustaining Capital: $145,300,000
Sustaining capital was allotted based on the rule of thumb proposed by Larry Smith, at $0.30/tonne
mined per year for mining equipment and 1% of initial capital of processing equipment. The annual
allotted sustaining capital is shown in Appendix Section 16.22.
13.4 Taxes: $911,700,000 at a 30% tax rate
The effective tax rate of 30% was the sum of the 15% tax rate in the Yukon territories and 15% corporate
tax rate to the Federal government [47]. Depreciation rate used was the standard CCA rate for large,
industrial equipment at 30% [48]. The annual tax and depreciation calculations are shown in Appendix
Section 16.23.
14 Conclusions & Recommendations
By looking at the various aspects of potential operations at the Grum site, it appears that despite the
relatively remote location and environmental concerns with the sites waste products, this project should
be advanced to the next stage. Its high internal rate of return and reasonable NPV of $156.1 Million
suggests further studies are warranted.
In order to conduct a successful next round of study current findings suggests that a detailed look into
the metallurgical behaviour and processing scheme is necessary. Additionally further site investigation
for potential building material, such as the overburdens, would be useful to confirm availability. Lastly
73
baseline studies should be considered to ensure appropriate time is given to allow necessary permits
time to be processed. If these steps are taken the next stage of study should be successful.
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74
[13] Toromont Industries Ltd., "Relative Pricing," 2016.
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2016].
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75
[27] R. Stasyszyn, "Yukon News: Faro mine's remediation mess," 3 July 2012. [Online]. [Accessed
February 2016].
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Guide," Government of Canada, 2014.
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[Online]. [Accessed 3 February 2016].
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Dena Kayeh Institute, Lower Post, 2010.
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Available: http://www.emr.gov.yk.ca/mining/pdf/mine_reclamation_policy_web_nov06.pdf .
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emphasis on northern environment," 2009. [Online]. Available: http://www.dfo-
mpo.gc.ca/Library/337077.pdf.
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76
https://www.google.ca/maps/dir/Faro,+YT/Trail,+BC/@54.5201333,-
132.8043953,4.75z/data=!4m14!4m13!1m5!1m1!1s0x5150b66882cd1707:0x18bf64e7f49b3df!2
m2!1d-
133.3531599!2d62.2285419!1m5!1m1!1s0x5362d9f2d078efbb:0xdabf948e746e9f84!2m2!1d-
117.7117301!2d49.0965676.
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http://www.skagway.org/index.asp?SEC=21FD65B5-E64D-488B-9F51-
51B650B9D6DE&Type=B_BASIC.
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gun,+Ulsan,+South+Korea/@35.432512,129.1953824,11z/data=!3m1!4b1!4m2!3m1!1s0x3567d
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b3df!2m2!1d-
133.3531599!2d62.2285419!1m5!1m1!1s0x56aaa92500de9141:0x5e251777323a536a!2m2!1d-
135.3138889!.
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785C.pdf. [Accessed March 2016].
77
16 Appendices
16.1 Ramp Design Considerations
Figure 16-1 Haulage Truck Specifications- Cat 785C [49]
78
Figure 16-2 Ramp Design for the first push back at Whittle Pit 6
Figure 16-3 Ramp Design for the second push back at Whittle Pit 9
79
Figure 16-4 Ramp Design for the third push back at Whittle Pit 18
16.2 Re-sloped Pit Calculations
Formula employed sectors 8, 9, 10 and 1:
Tan = {(n x h) / [((n-1) x w) + ((n x h)/ tan) + Ramp width]}
= Overall pit slope angle with ramp
n= Number of benches
h = Bench height
w = Bench width
= Bench face angle
Sample Calculation of sector 8
= tan-1 {(25 x 12) / [((25 -1) x 6.9) + ((25 x 12) / tan (80)) + 30]} = 49
Note: The same process for calculating was carried out for sectors 1, 9 and 10, but with different bench
dimensions.
80
16.3 Equipment Unit Costs
Table 16-1 Summary of Associated Unit Costs for Selected Machinery
81
16.4 Provided Metallurgical Recovery Data
Table 16-2 Preliminary Recovery Data Provided for the Grum Deposit
82
16.5 Initial Tailings Volumes
Table 16-3 Table Showing the process in Calculating Annual Tailings Volumes
End of Production Year Incremental Ore Processed (Tonnes) Process Dry Waste (Ton) Volume Water Added (m3)
1 1.60E+05 1.54E+05 6.14E+04
2 8.80E+05 8.44E+05 3.38E+05
3 2.40E+06 2.30E+06 9.21E+05
4 3.12E+06 2.99E+06 1.20E+06
5 2.80E+06 2.69E+06 1.07E+06
6 2.48E+06 2.38E+06 9.52E+05
7 3.20E+06 3.07E+06 1.23E+06
8 3.20E+06 3.07E+06 1.23E+06
9 3.20E+06 3.07E+06 1.23E+06
10 3.20E+06 3.07E+06 1.23E+06
11 3.20E+06 3.07E+06 1.23E+06
12 3.20E+06 3.07E+06 1.23E+06
13 3.20E+06 3.07E+06 1.23E+06
14 3.20E+06 3.07E+06 1.23E+06
15 3.20E+06 3.07E+06 1.23E+06
16 3.20E+06 3.07E+06 1.23E+06
17 3.20E+06 3.07E+06 1.23E+06
18 3.20E+06 3.07E+06 1.23E+06
19 3.20E+06 3.07E+06 1.23E+06
20 1.84E+06 1.77E+06 7.06E+05
End of Production Year Dry Volume Added (m3) Tailings Volume Added (m3) Cumm. Tailings Volume (m3)
1 5.81E+04 1.20E+05 1.20E+05
2 3.20E+05 6.58E+05 7.77E+05
3 8.72E+05 1.79E+06 2.57E+06
4 1.13E+06 2.33E+06 4.90E+06
5 1.02E+06 2.09E+06 6.99E+06
6 9.01E+05 1.85E+06 8.85E+06
7 1.16E+06 2.39E+06 1.12E+07
8 1.16E+06 2.39E+06 1.36E+07
9 1.16E+06 2.39E+06 1.60E+07
10 1.16E+06 2.39E+06 1.84E+07
11 1.16E+06 2.39E+06 2.08E+07
12 1.16E+06 2.39E+06 2.32E+07
13 1.16E+06 2.39E+06 2.56E+07
14 1.16E+06 2.39E+06 2.80E+07
15 1.16E+06 2.39E+06 3.04E+07
16 1.16E+06 2.39E+06 3.28E+07
17 1.16E+06 2.39E+06 3.51E+07
18 1.16E+06 2.39E+06 3.75E+07
19 1.16E+06 2.39E+06 3.99E+07
20 6.69E+05 1.37E+06 4.13E+07
83
16.6 TSF Volume Calculations: Volume of a Truncated Pyramid
𝑉 = ℎ
6(𝑎𝑏 + (𝑎 + 𝑐)(𝑏 + 𝑑) + 𝑐𝑑) [6]
Figure 16-5 Diagram showing the Meanings of each constant in the Truncated Pyramid Volume Calculation
84
16.7 Summary of the Annual Rate of Rise of Tailings Deposition
Table 16-4 Table Showing Summary of Tailings Rate of Rise for the final TSF design. Notice the given Storage Length and Width used in the design.
Dam Height 60 m
Storage Width 800 m
Storage Length 1500 m
Year of Production Pond Width (m) Pond Length (m) Beach Length of Tails (m)
1 501 1201 0.5
2 506 1206 3.5
3 521 1221 11.2
4 539 1239 20.8
5 554 1254 29.1
6 567 1267 36.2
7 583 1283 44.9
8 599 1299 53.4
9 614 1314 61.5
10 629 1329 69.3
11 643 1343 76.9
12 656 1356 84.3
13 670 1370 91.4
14 683 1383 98.3
15 695 1395 105.0
16 707 1407 111.6
17 719 1419 118.0
18 731 1431 124.2
19 742 1442 130.3
20 748 1448 133.7
Year of Production Pond Area(m2) Itterative Tailing Height (m) Incremental Rate of Rise (m/year)
1 6.02E+05 0.20 0.2
2 6.11E+05 1.28 1.1
3 6.36E+05 4.16 2.9
4 6.67E+05 7.74 3.6
5 6.95E+05 10.81 3.1
6 7.19E+05 13.43 2.6
7 7.49E+05 16.69 3.3
8 7.78E+05 19.82 3.1
9 8.07E+05 22.84 3.0
10 8.35E+05 25.75 2.9
11 8.63E+05 28.57 2.8
12 8.90E+05 31.29 2.7
13 9.17E+05 33.94 2.6
14 9.44E+05 36.51 2.6
15 9.70E+05 39.01 2.5
16 9.95E+05 41.44 2.4
17 1.02E+06 43.81 2.4
18 1.05E+06 46.13 2.3
19 1.07E+06 48.39 2.3
20 1.08E+06 49.67 1.3
85
16.8 Soil Classification of the Overburden Material
Table 16-5 Summary of the Soil Classification of the Grum Overburden Material, including Key Findings
Plasticity index 10%
Liquid limit 25%
15% CL or OL > Plots between "U" line and "A" line
Sand 35%
Silt 40%
Clay 25%
Gravel 0%
Calcite 3%
Dolomite 3%
Closing Remarks
>Likely to be SC-Clayey Sand
Important Note: No info provided on Oven dried LL, as a result could be classified as Sandy Organic Clay
which is a serious cause for concern for the development of long term or permanent infrastructure.
Further lab tests are required to confirm this.
>12% fines w/ CL symbol
Weak material % is low
Notice that trench data (shown on the right) reflects closer to 15%
Low amounts of Carbonate minerals are present; as a result there is low concern for their possible
decreasing of the soil strength. However there does appear to be localized anomalies containing
higher contents of dolomite which should be looked out for.
Supplimentary information on till:
Atterberg limits
Grain size distribution
Carbonate content
Comments
Plastic Limit Plasticity Chart:
ie. Quartz and feldspar, between clay and Sand
ie. Fines < Sieve #200
Assumed inorder to add to 100%
86
16.9 TSF Option Comparison
Table 16-6 Economic Indicators TSF Option Comparison
87
Table 16-7Environmental Indicators TSF Option Comparison
Table 16-8 Social Indicators TSF Option Comparison
88
16.10 WRD Option Comparison
Table 16-9 Economic Indicators WRD Option Comparison
89
Table 16-10 Environmental Indicators WRD Option Comparison
Table 16-11 Social Indicators WRD Option Comparison
90
16.11 Detailed Water Balance
Table 16-12 Summary of the reported Detailed Water Balance
End of Production Year Volume Water Added with Tailings (m3) Trapped Water in Tailings Bleed Water
1 6.14E+04 2.30E+04 3.84E+04
2 3.38E+05 1.27E+05 2.11E+05
3 9.21E+05 3.45E+05 5.76E+05
4 1.20E+06 4.49E+05 7.48E+05
5 1.07E+06 4.03E+05 6.72E+05
6 9.52E+05 3.57E+05 5.95E+05
7 1.23E+06 4.61E+05 7.68E+05
8 1.23E+06 4.61E+05 7.68E+05
9 1.23E+06 4.61E+05 7.68E+05
10 1.23E+06 4.61E+05 7.68E+05
11 1.23E+06 4.61E+05 7.68E+05
12 1.23E+06 4.61E+05 7.68E+05
13 1.23E+06 4.61E+05 7.68E+05
14 1.23E+06 4.61E+05 7.68E+05
15 1.23E+06 4.61E+05 7.68E+05
16 1.23E+06 4.61E+05 7.68E+05
17 1.23E+06 4.61E+05 7.68E+05
18 1.23E+06 4.61E+05 7.68E+05
19 1.23E+06 4.61E+05 7.68E+05
20 7.06E+05 2.65E+05 4.41E+05
Total Over LOM 2.12E+07 7.96E+06 1.33E+07
% Contribution 24% 9% 15%
End of Production Year Catchment Area (m2) Vol. of Yearly Avg Rainfall (m3) Vol. of Soil Seepage (m3) Vol. of Evaporation(m3)
1 3.01E+06 9.51E+05 5.22E+05 1.81E+05
2 3.05E+06 9.65E+05 5.30E+05 1.83E+05
3 3.18E+06 1.00E+06 5.51E+05 1.91E+05
4 3.34E+06 1.05E+06 5.79E+05 2.00E+05
5 3.47E+06 1.10E+06 6.03E+05 2.08E+05
6 3.59E+06 1.14E+06 6.23E+05 2.16E+05
7 3.74E+06 1.18E+06 6.49E+05 2.25E+05
8 3.89E+06 1.23E+06 6.75E+05 2.33E+05
9 4.04E+06 1.28E+06 7.00E+05 2.42E+05
10 4.18E+06 1.32E+06 7.25E+05 2.51E+05
11 4.32E+06 1.36E+06 7.49E+05 2.59E+05
12 4.45E+06 1.41E+06 7.72E+05 2.67E+05
13 4.59E+06 1.45E+06 7.95E+05 2.75E+05
14 4.72E+06 1.49E+06 8.18E+05 2.83E+05
15 4.85E+06 1.53E+06 8.41E+05 2.91E+05
16 4.98E+06 1.57E+06 8.63E+05 2.99E+05
17 5.10E+06 1.61E+06 8.85E+05 3.06E+05
18 5.23E+06 1.65E+06 9.07E+05 3.14E+05
19 5.35E+06 1.69E+06 9.28E+05 3.21E+05
20 5.42E+06 1.71E+06 9.40E+05 3.25E+05
Total Over LOM 2.67E+07 1.47E+07 5.07E+06
% Contribution 3.00E-01 1.65E-01 5.71E-02
91
End of Production Year Yearly TSF Water Balance (m3) Cummulative Water Balance (m3)
1 2.87E+05 2.87E+05
2 4.63E+05 7.50E+05
3 8.38E+05 1.59E+06
4 1.02E+06 2.61E+06
5 9.58E+05 3.57E+06
6 8.91E+05 4.46E+06
7 1.08E+06 5.54E+06
8 1.09E+06 6.63E+06
9 1.10E+06 7.73E+06
10 1.11E+06 8.84E+06
11 1.12E+06 9.96E+06
12 1.14E+06 1.11E+07
13 1.15E+06 1.22E+07
14 1.16E+06 1.34E+07
15 1.17E+06 1.46E+07
16 1.18E+06 1.57E+07
17 1.19E+06 1.69E+07
18 1.20E+06 1.81E+07
19 1.21E+06 1.93E+07
20 8.89E+05 2.02E+07
Total Over LOM 2.02E+07
92
16.12 Suggested Water Pumping Schedule to Maintain the Water Cover
Table 16-13 Summary of the Recommended Pumping Schedule and resulting Water Balance (note the negative values require pumping of water out of the dam)
93
16.13 Measurement of Truck Routes
Figure 16-6 – An overview of the mine site layout for context, with dimensions of paths superimposed. For a clearer depiction of measurements, refer to subsequent figures.
94
Figure 16-7 - View of horizontal projection distances of equipment travel paths; due to the high degree of segmentation in the pit, dimensions are overlapping and difficult to read. A magnified view could be found in Figure 16-6.
95
Figure 16-8 - A magnified view of the horizontal projection lengths of the pit ramp.
96
16.14 Rimpull and Retardation Curves
Figure 16-9 - Rimpull curve of the CAT 777G, with appropriate speeds determined for loaded travel on effective grades of 3%, 4%, and 13%.
97
Figure 16-10 - Retardation curve of an empty CAT 777G on effective grades of 0% and 7%.
98
Figure 16-11 - Rimpull curve of the CAT 785D, with appropriate speeds determined for loaded travel on effective grades of 3%, 4%, and 13%.
99
Figure 16-12 - Retardation curve of an empty CAT 785D on effective grades of 0% and 7%.
100
Figure 16-13 - Rimpull curve of the CAT 789D, with appropriate speeds determined for loaded travel on effective grades of 3%, 4%, and 13%.
101
Figure 16-14 - Retardation curve of an empty CAT 789D on effective grades of 0% and 7%.
102
16.15 Travel Times
Table 16-14 - Travel times for various road segments on the route of a CAT 777G
Segment Length Grade RR Total RR
Loaded Speed
Time in seg
Total RR
Empty Speed
Time in seg
Units m % % % km/h min % km/h min
Load area 100 0 4 4 32 0.19 0 35 0.17
Ramp 3298 10 3 13 10 19.79 7 35 5.65
Crusher Road 1713 0 3 3 43 2.39 0 50 2.06
Crusher Ramp 100 10 3 13 10 0.6 7 35 0.17
Crusher Pad 50 0 3 3 15 0.2 0 35 0.09
Dump Road 702 0 3 3 43 0.98 0 50 0.84
Dump Ramp 3224 10 3 13 10 19.34 7 35 5.53
Dump Run 423 0 3 3 30 0.85 0 35 0.73
Dumping Time 1 1
Table 16-15 -Travel times for various road segments on the route of a CAT 785D
Segment Length Grade RR Total RR
Loaded Speed
Time in seg
Total RR
Empty Speed
Time in seg
Units m % % % km/h min % km/h min
Load area 100 0 4 4 33 0.18 0 35 0.17
Ramp 3298 10 3 13 10 19.79 7 35 5.65
Crusher Road 1713 0 3 3 45 2.28 0 50 2.06
Crusher Ramp 100 10 3 13 10 0.6 7 35 0.17
Crusher Pad 50 0 3 3 15 0.2 0 35 0.09
Dump Road 702 0 3 3 45 0.94 0 50 0.84
Dump Ramp 3224 10 3 13 10 19.34 7 35 5.53
Dump Run 423 0 3 3 30 0.85 0 35 0.73
Dumping Time 1 1
Table 16-16 - Travel times for various road segments on the route of a CAT 789D
Segment Length Grade RR Total RR
Loaded Speed
Time in seg
Total RR
Empty Speed
Time in seg
Units m % % % km/h min % km/h min
Load area 100 0 4 4 35 0.17 0 35 0.17
Ramp 3298 10 3 13 10 19.79 7 35 5.65
Crusher Road 1713 0 3 3 50 2.06 0 50 2.06
103
Crusher Ramp 100 10 3 13 10 0.6 7 35 0.17
Crusher Pad 50 0 3 3 20 0.15 0 35 0.09
Dump Road 702 0 3 3 50 0.84 0 50 0.84
Dump Ramp 3224 10 3 13 10 19.34 7 35 5.53
Dump Run 423 0 3 3 30 0.85 0 35 0.73
Dumping Time 1 1
16.16 Loading Times
Table 16-17 -The time involved in a load, haul, dump, return cycle of a CAT 777D
6015 6018 6030
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 7 7 10 10 16.5 16.5
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 13.30 11.97 19.00 17.10 31.35 28.22
Haul Truck payload - Target t 100.0 100.0 100.0 100.0 100.0 100.0
Passes to Fill Truck # 8 8 5 6 3 4
Haul Truck Payload - Actual t 106.4 95.8 95.0 102.6 94.1 112.9
Haul Truck Payload Utilized % 106% 96% 95% 103% 94% 113%
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 3.60 3.60 2.10 2.60 1.10 1.60
6040 6050 6060
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 22 22 26 26 34 34
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 41.80 37.62 49.40 44.46 64.60 58.14
Haul Truck payload - Target t 100.0 100.0 100.0 100.0 100.0 100.0
Passes to Fill Truck # 2 3 2 2 2 2
Haul Truck Payload - Actual t 83.6 112.9 98.8 88.9 129.2 116.3
Haul Truck Payload Utilized % 84% 113% 99% 89% 129% 116%
104
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 0.60 1.10 0.60 0.60 0.60 0.60
Table 16-18 - The time involved in a load, haul, dump, return cycle of a CAT 785D
6015 6018 6030
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 7 7 10 10 16.5 16.5
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 13.30 11.97 19.00 17.10 31.35 28.22
Haul Truck payload - Target t 150.0 150.0 150.0 150.0 150.0 150.0
Passes to Fill Truck # 11 13 8 9 5 5
Haul Truck Payload - Actual t 146.3 155.6 152.0 153.9 156.8 141.1
Haul Truck Payload Utilized % 98% 104% 101% 103% 105% 94%
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 5.10 6.10 3.60 4.10 2.10 2.10
6040 6050 6060
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 22 22 26 26 34 34
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 41.80 37.62 49.40 44.46 64.60 58.14
Haul Truck payload - Target t 150.0 150.0 150.0 150.0 150.0 150.0
Passes to Fill Truck # 4 4 3 3 2 3
Haul Truck Payload - Actual t 167.2 150.5 148.2 133.4 129.2 174.4
Haul Truck Payload Utilized % 111% 100% 99% 89% 86% 116%
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 1.60 1.60 1.10 1.10 0.60 1.10
105
Table 16-19 -The time involved in a load, haul, dump, return cycle of a CAT 789D
6015 6018 6030
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 7 7 10 10 16.5 16.5
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 13.30 11.97 19.00 17.10 31.35 28.22
Bucket payload - maximum (rated) t
Haul Truck payload - Target t 200.0 200.0 200.0 200.0 200.0 200.0
Passes to Fill Truck # 15 17 11 12 6 7
Haul Truck Payload - Actual t 199.5 203.5 209.0 205.2 188.1 197.5
Haul Truck Payload Utilized % 100% 102% 105% 103% 94% 99%
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 7.10 8.10 5.10 5.60 2.60 3.10
6040 6050 6060
Load Time Units ORE WASTE ORE WASTE ORE WASTE
Bucket size m3 22 22 26 26 34 34
Bucket fill factor % 95.00% 95.00% 95.00% 95.00% 95.00% 95.00%
Loose density t/m3 2.00 1.80 2.00 1.80 2.00 1.80
Bucket payload - Actual t 41.80 37.62 49.40 44.46 64.60 58.14
Bucket payload - maximum (rated) t
Haul Truck payload - Target t 200.0 200.0 200.0 200.0 200.0 200.0
Passes to Fill Truck # 5 5 4 4 3 3
Haul Truck Payload - Actual t 209.0 188.1 197.6 177.8 193.8 174.4
Haul Truck Payload Utilized % 105% 94% 99% 89% 97% 87%
First Bucket min 0.10 0.10 0.10 0.10 0.10 0.10
Time per pass (cycle time) min 0.5 0.5 0.5 0.5 0.5 0.5
Spot time min 0.75 0.75 0.75 0.75 0.75 0.75
Loading Time - Total min 2.10 2.10 1.60 1.60 1.10 1.10
16.17 Number of Trucks Required per Shovel
Table 16-20 - The number of CAT 777G trucks required for each type of shovel
106
6018 6018 6030
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 35.91 61.30 34.41 60.30 33.41 59.30
Productivity/hr Trucks t/hr 178 94 166 102 169 114
Required Productivity/hr t/hr 494 3618 494 3618 494 3618
Required trucks units 2.92 38.59 2.92 35.44 2.92 31.68
Total Required Trucks Units 42 39 35
6040 6050 6050
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 32.91 58.80 32.91 58.30 32.91 58.30
Productivity/hr Trucks t/hr 152 115 180 91 236 120
Required Productivity/hr t/hr 494 3618 494 3618 494 3618
Required trucks units 2.92 31.41 2.92 39.55 2.92 30.23
Total Required Trucks Units 35 43 34
Table 16-21 - The number of CAT 785D trucks required for each type of shovel
6018 6018 6030
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 37.29 63.68 35.79 61.68 34.29 59.68
Productivity/hr Trucks t/hr 235 147 255 150 274 142
Required Productivity/hr t/hr 494 3618 494 3618 494 3618
Required trucks units 1.8 24.68 1.8 24.17 1.8 25.51
Total Required Trucks Units 27 26 28
6040 6050 6050
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 33.79 59.18 33.29 58.68 32.79 58.68
Productivity/hr Trucks t/hr 297 153 267 136 236 178
Required Productivity/hr t/hr 494 3618 494 3618 494 3618
107
Required trucks units 1.8 23.71 1.8 26.53 1.8 20.29
Total Required Trucks Units 26 29 23
Table 16-22 - The number of CAT 789D trucks required for each type of shovel
6018 6018 6030
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 37.29 63.68 35.79 61.68 34.29 59.68
Productivity/hr Trucks t/hr 235 147 255 150 274 142
Required Productivity/hr t/hr 494 3160 494 3160 494 3160
Required trucks units 1.8 21.56 1.8 21.11 1.8 22.28
Total Required Trucks Units 24 23 25
6040 6050 6050
Truck Count Calculation Units Ore Waste Ore Waste Ore Waste
Total Cycle min 33.79 59.18 33.29 58.68 32.79 58.68
Productivity/hr Trucks t/hr 297 153 267 136 236 178
Required Productivity/hr t/hr 494 3160 494 3160 494 3160
Required trucks units 1.8 20.71 1.8 23.17 1.8 17.72
Total Required Trucks Units 23 25 20
108
16.18 Environmental and Social Impact Assessment
Table 16-23 Yukon Air Quality and Particulate Matter Standards
Parameter Yukon Guideline
Sulphur Dioxide (SO2) μg/m3
1-hour average 450
24-hour average 160
Annual arithmetic mean 25
Nitrogen Dioxide (NO2) μg/m3
1-hour average 400
24-hour average 200 (106 ppb)
Annual arithmetic mean 100 (53 ppb) 60 (32 ppb)
Carbon Monoxide (CO) μg/m3
1-hour average 14,300 (13 ppm)
8-hour average 5,500 (5 ppm)
Total Suspended Particulate Matter (TSP) μg/m3
24-hour average 120
Annual arithmetic mean 60
Particulate Matter (PM2.5) μg/m3
24-hour average 30
109
Excerpt from interview of Ian Dunlop, CAO Faro by Pearl Barrett:
“… The Pelly River is noted for its fishing opportunities including Arctic Grayling and Salmon. Vangorda
Creek is a small tributary that may have some seasonal fishing but there are better choices in the area.
Swimming in the Pelly is not recommended due to the swift currents, but you could wade into the
shallower areas near shore if you like. Typical water mammals in the area include beaver and muskrat.
The Town also has a stocked lake with a day-use recreation park called Fisheye Lake, with Kokanee and
Rainbow Trout. You can swim in the lake in summer…, but keep in mind the water is quite cold as it is
throughout the Yukon…”
110
Table 16-24: Yukon water quality standards to monitor and follow, the bolded items are pertinent to the Grum Site.
Parameter
Units
WUL QZ96-006 Effluent Quality Standards
Frequency Daily Limit
pH pH units Weekly 6.5-9.0
Suspended Solids mg/L Weekly 15
Aluminum mg/L Weekly 0.5
iron mg/L Weekly 1
Copper mg/L Weekly 0.01
Lead mg/L Weekly 0.002
Manganese mg/L Weekly 0.2
Nickel mg/L Weekly 0.065
Zinc mg/L Weekly 0.03
Total Ammonia mg/L Weekly 1
Oil and Grease visibility Weekly No visible oil or grease
Rainbow Trout Acute Lethality Test
<50% mortality in 100% effluent
Monthly
Pass
111
Table 16-25: Risk assessment criteria for event severity
Level Descriptor Detail Descriptions
Safety Cultural Socio-Economic Economic Environmental
1 Insignificant Minor injury, requiring first aid only.
No to minimal impact on Indigenous or historical heritage sites or values.
No or few reported complaints about the project. No socio-economic impacts.
<1% of revenue
No impact, minor breach in procedure, minor nonconformance
2 Minor Medical attention required.
Minor impact on Indigenous or historical heritage sites or values (e.g. restricted access to recreation areas).
Some inconvenience to stakeholders, minimal adverse impact on socio-economic environment, and some intervention required.
1%-2% of revenue
Minimal impact outside the local area.
3 Moderate Disability/ Lost Time Incident
Moderate impact on Indigenous or historical heritage sites, managed under normal procedures. Some negative media coverage could be expected.
Moderate disruption or inconvenience to stakeholders. Require careful management to restore trust.
2%-5% of revenue
Minimal impact outside the local area.
4 Major Permanent Disability / Fatality
Major disturbances to 1 or 2 significant Indigenous or historical heritage sites or values. Major breach of statutory obligation, access to resource denied in the medium to long-term.
Significant adverse impacts to sectors of the community and stakeholders. Long-term social disruption, diminished quality of life of large or specific sectors of the community (e.g. fishing sector)
5%-10% of revenue
Major environmental harm or breach of license conditions or obligations, discharges off site.
5 Catastrophic Multiple Fatalities
Major disturbances to a number (3 or more) of significant Indigenous or historical heritage sites or values. Major breach of statutory obligation, access to resource permanently denied.
Irreversible damage to the socio-economic environment. Potential for strike or riot and major damage to property or haul routes.
>10% of revenue
Long term, significant ecological changes, with legal implications and potential to affect community health.
112
Table 16-26: Risk assessment criteria for event probability
Level Measure Description Guide
A Almost certain Issue will or almost certainly will occur, is currently a problem or is expected to occur in most circumstances (e.g. acid generation from waste).
Weekly
B Likely to occur Issue has been a common problem in the past and there is a high probability it will occur in most circumstances.
Once per month
C Moderate Issue may have arisen in the past and there is a high probability that it should occur at some time.
Once per year
D Unlikely Issue may have occurred in the past, and there is a moderate probability that it could occur at some time.
Once per 10 years
E Rare Issue has not occurred in the past, and there is a low probability that it may occur in exceptional circumstances.
Once per 100 years
Table 16-27: Risk Matrix
Severity
Probability 1
Insignificant
2
Minor
3
Moderate
4
Major
5
Catastrophic A. Almost Certain M S H H H
B. Likely to occur M M S H H
C. Moderate L M S S H
D. Unlikely L L M S S
E. Rare L L M M S
H- High
Detailed research and management planning required at senior levels. Immediate action required
S- Significant
Senior management attention needed
M- Moderate
Management responsibility and integration into management plans required
113
L- Low
Manage by routine procedures
Table 16-28: Impact assessment: Pit development and mining
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Pit Development and Mining
Clearing Area for Pit development, and site necessities (TSF, WRD, etc)
Terrestrial System 2 A S - Create habitat buffer
areas around streams
- Loss of vegetation and habitat - Increased weed species
Cultural Heritage 3 C S - Report any found
artefacts or sites - Consult with community
- Interference/ damage/ destruction to local heritage
Water Systems
3
A
H
- Divert the minimum amount of stream
- Place all other site infrastructure away from streams
- Monitor water quality
- Stream diversion causing erosion and increased suspended solids
Atmospheric System 2 B M - Design a vegetated
buffer zone - Use water spraying to
reduce airborne dust - Noise caused by vehicle traffic - Dust generated by exposed
areas and machinery
Socioeconomic 1 A M
- Provide compensation for lost land
- Employ local community members
- Give preference to Faro members and Kaska members for jobs and bids within means
- Loss of recreational land - Increase in nuisance noise,
vibrations - Creation of jobs
Acid generation Water Systems 3 B S - Pump contaminated
water to treatment
114
from pit walls - Increased acidity of groundwater, surface water and soil
plant, clean and use for processing
Dewatering Water Systems 2 D L - Release appropriate
amount of water where needed
- Treat water before release - Decrease in regional
groundwater - Decrease in regional water
quality
Terrestrial System 2 D L - Release appropriate
amount of water where needed
- Treat water before release - Decreased groundwater leads
to vegetation impacts
Blasting Atmospheric System
2
A
S
- Blast at certain times of the day
- Use mobile water sprayers to reduce dust
- Design a vegetated buffer zone
- Dust: Increased particulate matter in the air
- Noise and vibration
Terrestrial System
1
B
M
- Monitor wildlife - Provide a barrier
- Damage by fly rock - Cumulative effect on wildlife
Greenhouse gas emissions
Atmospheric System 1 A M - Reduce number of diesel
powered equipment - Use electric power at
the site - Send only full load trucks
to smelter
- Contribution to human caused climate change
- Contribution to Canada’s overall greenhouse gas emissions
- Degradation of the environment
Explosives incident Socioeconomic
5 E S - Ensure person managing
explosives is qualified and well trained
- Place explosives magazine far from infrastructure
- Blast only during shift change
- Injury to employees - Damage to mine infrastructure
115
Heavy equipment accident
- Injury to truck driver/ other employees 4 D S
- A dry site will reduce the number of accidents, no drinking, no drugs
Pit ramp/wall failure
Socioeconomic
4
D
S
- Properly train employees
- Monitor the pit walls - Install displacement
monitors near ramp to monitor integrity of rock
- Injury to employees - Damage to mine infrastructure - Disruption of production
Flooding of open pit Water Systems 3 E M - Install pumping sites
directed to water treatment plant
- Effluent likely pumped untreated
Socioeconomic 3 D M - Train employees for
events
- Injury to employees - Disruption of production
116
Table 16-29: Impact assessment: Waste rock dump
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Waste Rock Dump
Placement and visual impact of waste rock dumps on surrounding landscape
Socioeconomic
3
A
H
- Vegetation barrier - Natural slopes at closure - Dump can be reduced in
height if increase in footprint, consult with community
- Decreased tourism and recreation
- Aesthetics
Generation of acid rock drainage
Water/Terrestrial Systems 4 A H - Impermeable barrier
below waste rock - Catchment ditch around
structure to capture runoff
- Well distributed particle size to reduce oxygen contact
- Increased acidity of soil, groundwater, and surface water
- Erosion causing sediment build up
Table 16-30: Impact assessment: Tailings storage facility
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Tailings Storage Facility
Generation of acid rock drainage
Water/Terrestrial Systems 2 A S - Impermeable barrier
below dam - Conservative dam design - Catchment ditch around
structure to capture runoff
- Increased acidity of soil, groundwater, and surface water
Complete Dam Failure
Water/ Terrestrial Systems 5 E S - Monitor water levels - Repair the dam as soon
as maintenance is required
- Recycle tailings water to reduce volumes
- Monitor pore pressures and wall displacements
- Increased acidity of soil, groundwater, and surface water
- Sediment build up - Effluent limits exceeded killing
fish populations
117
Placement and visual impact of tailings facility on surrounding landscape
Socioeconomic 2 D L - Consult community to discuss alternatives
- Decreased tourism and recreation
- Aesthetics
Table 16-31: Impact assessment: Waste Management
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Waste Management
Odour Terrestrial/
Socioeconomic
1
C
L
- On site waste incinerator
- Enclosed compost facility for revegetation nursery
- Attraction of scavengers such as bears
- Safety issues
Disposal Terrestrial/ Water Systems 1 C L - On site waste
incinerator - Enclosed compost
facility for revegetation nursery
- Septic management
- Contamination of soil, groundwater and surface water
- Vegetation Loss
Atmospheric System
1
C
L
- Incinerate waste and filter byproduct
- Landfill greenhouse gas emissions
118
Table 16-32: Impact assessment: General operational
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
General Operational
Increased large vehicle use on roads
Atmospheric System 3 A H - Send trucks in convoys
at low speeds to reduce frequency of noise
- Increased noise and air emissions
Terrestrial Systems
4
B
H
- Send trucks in convoys at low speeds to reduce chance of wildlife collisions
- Create wildlife crossing points
- Wildlife roadkill
Light vehicle accident
Socioeconomic 4 B H - Post and monitor speed
limits
- Injury to employees or
members of the public
Table 16-33: Impact assessment: Closure and remediation
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Closure and Remediation
Impact on environment over time after pit closes
Socioeconomic 3 E M - Return the site to
sustainable pre-mining landscape
- Monitor and provide proper signage - Third party use
- Ex. swimming - Visual disturbance
Water System 3 D M - Monitor water quality
and levels
- Change in groundwater quality
and levels
119
Poor seeding success
Terrestrial System 3 D M - Test multiple native flora
species - Try again using tested
vegetation - Use fertilizers or
resurface with new topsoil
- Loss of vegetation - Slow growth rates - Loss of stabilization of soil
Impact on soils and degree of erosion
Water System
3
B
S
- Progressive remediation to reduce vulnerable areas
- Sediment buildup in surface water and other drainage systems; water finding alternative route causing erosion and flooding of areas
Atmospheric System 3 C S - Progressive remediation to reduce vulnerable areas
- Monitor air quality - Cover if necessary
- Wind erosion before revegetation causing suspended particulate matter
120
Table 16-34: Impact assessment: Natural disasters
Key Issues Key Potential Impacts, VEC impacted Consequence Ranking
Likelihood Ranking
Risk Ranking Mitigation Strategy
Natural Disasters
Severe Electrical Storm
Socioeconomic 5 D S - Have backup generators
ready in case of power outage
- Make sure all employees wear proper clothing
- Disaster training for employees
- Employee struck by lightning - Lightning initiating fire or
explosion - Disruption to power/ security
systems
Cyclone or seismic event
Water Systems 4 E M - Design tailings facility
conservatively to keep the tailings behind the dam in such circumstances - Collapse of mine infrastructure
- Release of untreated effluent, tailings, acid generating waste
Socioeconomic 5 C H - Have a medical trained
staff member on site at all times
- Record events - Communication disruptions - Employee injury and inability to
evacuate
Major Forest Fire Socioeconomic
4
C
S
- Have a medical trained staff member on site at all times
- make sure the employee fighting the fire has been properly trained
Injury to employees fighting the bushfire
Atmospheric Systems 3 C S - House flammable goods
such as fuel in flame retardant containers
- Unregulated burning of fuel, waste, materials, etc.
- Decrease in air quality
121
16.19 Closure Costs Table 16-35 Unit Costs of Items Needed for Closure
Contractor Unit Rates; Misc Costs Units Cost ($)
Excavation of Soil in Stockpile m3 5
Supply & place geotextile m2 12
Load, haul and place topsoil m3 5
Load, haul and place tailings cover m3 7
Load, haul and place rock cover, m3 8 organics, granular till and clay
Drill, Blast and Haul Rip Rap m3 22
Place Rip Rap m3 14 Camp costs day/person 75 Surface water quality analyses sample set 420 Ground water quality analyses sample set 290
Water treatment costs m3 0.4 Revegetation seed mix kg 13 Fertilizer kg 1 Seed and fertilizer application ha 1,500
Concrete m3 85 Erosion Barrier /linear km 3,000
122
16.20 Contained Process Metals
Figure 16-15 - Annual contained lead and zinc processed
Figure 16-16 - Annual contained silver and gold processed
0
50000000
100000000
150000000
200000000
250000000
300000000
350000000
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Po
un
ds
of
Co
nta
ined
Met
al
Year into Production
Annual Contained Lead and Zinc Processed
Lead Zinc
0
20000
40000
60000
80000
100000
120000
140000
160000
0
1000000
2000000
3000000
4000000
5000000
6000000
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Co
nta
ined
Go
ld (
toz)
Co
nta
ined
Silv
er (
toz)
Year into Production
Annual Contained Gold and Silver Processed
Silver Gold .
123
16.21 Mill Recoveries Used for Economics
Table 16-36 - The effective recoveries and recoverable metal of ore sent to the mill for each year of mine production
Recoveries of Ore Mined sent to the Mill (from the Mine)
Year 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18
PB% 87 87 88 88 89 88 90 88 88 90 89 88 88 87 87 86 88 87
ZN% 87 87 88 88 88 88 89 88 88 90 88 88 88 87 87 85 88 88
AU% 58 58 59 60 60 60 60 61 58 60 60 59 60 59 59 55 59 59
AG% 52 52 52 52 52 52 52 52 51 52 51 52 51 51 51 51 51 51
Recoverable Metal from Ore sent to the Mill (from the Mine)
Year 1 2 3 4 5 6
PB (lbs) 4654451 35304064 108217502 150897225 133187762 115014877
ZN (lbs) 9379724 64222084 182306434 262753294 227153465 185289697
AU (toz) 1246 7679 24314 37228 35895 29772
AG (toz) 87213 540496 1542040 2143442 1915519 1681286
Year 7 8 9 10 11 12
PB (lbs) 146065776 127414654 136986276 138194609 103201042 85790397
ZN (lbs) 235351647 208004230 225962990 216882760 161840422 133118121
AU (toz) 45627 40677 39401 45552 39390 31025
AG (toz) 2152248 1886674 1948951 1984997 1513032 1268474
Year 13 14 15 16 17 18
PB (lbs) 104803958 71945331 79700163 104729101 122557357 130649169
ZN (lbs) 162766841 108687421 113090669 152516536 185219914 178477739
AU (toz) 34350 24343 27354 32936 40720 42666
AG (toz) 1581843 1047825 1190311 1562914 1796422 1753048
124
Table 16-37 - The effective recoveries and recoverable metal of ore sent from the stockpile
Recoveries of Ore sent to Mill from Stockpile
Year 11 12 13 14 15 16 17 18 19 20
PB % 88 88 88 88 88 0 0 0 88 88
ZN % 88 88 88 88 88 0 0 0 88 88
AU % 60 60 60 60 60 0 0 0 60 60
AG % 51 51 51 51 51 0 0 0 51 51
Recoverable Metal from Ore sent to Mill (from Stockpile)
Year 11 12 13 14 15 16 17 18 19 20
PB (lbs) 249158 298989 199326 548147 348821 0 0 0 1993260 1146125
ZN (lbs) 397561 477073 318049 874634 556585 0 0 0 3180487 1828780
AU (toz) 249158 298989 199326 548147 348821 0 0 0 1993260 1146125
AG (toz) 397561 477073 318049 874634 556585 0 0 0 3180487 1828780
16.22 Sustaining Capital
Table 16-38 - The calculated sustaining capital to be allotted annually over the LOM
-2 -1 1 2 3
Equipment $ 1,918,584 $ 3,837,168 $ 5,814,977 $ 6,917,247 $ 6,640,367
Processing - $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638
Total $ 1,918,584 $ 4,920,806 $ 6,898,614 $ 8,000,884 $ 7,724,005
4 5 6 7 8
Equipment $ 6,685,401 $ 6,762,947 $ 6,714,842 $ 6,689,750 $ 6,744,664
Processing $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638
Total $ 7,769,039 $ 7,846,585 $ 7,798,480 $ 7,773,387 $ 7,828,301
9 10 11 12 13
Equipment $ 6,729,956 $ 6,716,486 $ 6,707,313 $ 6,704,690 $ 6,704,531
Processing $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638
Total $ 7,813,594 $ 7,800,123 $ 7,792,950 $ 8,000,884 $ 7,724,005
14 15 16 17 18
Equipment $ 6,697,299 $ 3,837,168 $ 5,814,977 $ 6,917,247 $ 6,640,367
Processing - $ 1,083,638 $ 1,083,638 $ 1,083,638 $ 1,083,638
Total $ 1,918,584 $ 4,920,806 $ 6,898,614 $ 8,000,884 $ 7,724,005
19 20
Equipment $ 1,918,584 $ 3,837,168
Processing - $ 1,083,638
Total $ 1,918,584 $ 4,920,806
125
16.23 Depreciation and Tax Calculations
Table 16-39 - Depreciation (at 20%) and tax (at 30%) calculations
