Conceptual Mine Design, Grum YT

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

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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.

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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

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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

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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

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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

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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

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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

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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

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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

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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

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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

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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:

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● 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.

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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

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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
126
157
188
219
250
281
312
343
374
405
436
467
498
529
560
591
622
653
684
715
746
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 Allowance
Subtract 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)

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