Business Planning Group Design Guidelines & Standards

In MMTS Business Planning, the long range design group is responsible for the creation of long range mine plans broken down by – Phase 1, Phase 2 & Phase 3 studies. This document outlines the recommended standards & guidelines agreed to be used in long range underground design. For the purposes of this document and ultimately in use of the Business Planning Engineer Development Program, the definitions of standards and guidelines should be understood as: Design Standards - System's purpose is to standardize the best designs, subject to meeting or exceeding the Ontario Mining Regulations, of actual infrastructure such as sumps, electrical sub-stations, and refuge stations etc… Note: Regulations found in this document were valid for 2010. Refer to current regulations for possible updates. Design Guidelines - These were defined in the old Inco nomenclature as mine standards. This system's purpose is to standardize the best mine design practices (subject to meeting or exceeding the Ontario Mining Regulations) such as how to decide your ventilation requirements, the distances between sub-stations, power requirements, compressed air requirements, water requirements, pump selection, etc…

Table of Contents 1 Mining Methods .............................................................................................3

1.1 Sub Level Caving ...................................................................................4 1.2 Block Caving...........................................................................................6 1.3 Vertical Retreat Mining ...........................................................................7 1.4 Cut & Fill .................................................................................................9

1.4.1 Overhand.........................................................................................9 1.4.2 Underhand Cut & Fill .....................................................................10

1.5 Post Pillar .............................................................................................12 1.6 Room and Pillar ....................................................................................14

2 Business Planning Design Standard............................................................15 2.1 Infrastructure Sizing..............................................................................15 2.2 Development ........................................................................................16

2.2.1 Design ...........................................................................................16 2.2.2 Location of Drifts............................................................................16 2.2.3 Drift Sizing Info ..............................................................................17 2.2.4 Ramp Grades Phase 1& 2.............................................................19

3 Mineable Stope Shape ................................................................................20 4 Ground Control ............................................................................................21 5 Equipment Selection....................................................................................25 6 Ventilation ....................................................................................................26 7 Orepass Design ...........................................................................................26 8 Power ..........................................................................................................27 9 Cost estimating ............................................................................................28 10 Regulations ..............................................................................................30

10.1 Trackless Haulage Clearances.............................................................30 10.2 Overhead Clearance ............................................................................30 10.3 Refuge Station......................................................................................31 10.4 Powder Magazines ...............................................................................31 10.5 Fuel Bay ...............................................................................................31 10.6 Escapeway ...........................................................................................32 10.7 Boundary Pillars (U/G)..........................................................................32 10.8 Property Boundaries.............................................................................33

11 Rules of Thumb........................................................................................34 11.1 Passes, Bins, and Chutes.....................................................................34 11.2 Mining Methods ....................................................................................34 11.3 Backfill ..................................................................................................35 11.4 Mine Dewatering...................................................................................36 11.5 Lateral Development and Ramps .........................................................36 11.6 Mine Layout ..........................................................................................38 11.7 Cost Estimating ....................................................................................38 11.8 Misc ......................................................................................................39 11.9 Refuge Station......................................................................................39 11.10 Raise Bore Station ............................................................................40

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11.11 Existing Guidelines ...........................................................................40 12 References...............................................................................................41

List of Figures Figure 1 - Sublevel Cave Diagram........................................................................4 Figure 2 - Block Caving Diagram ..........................................................................6 Figure 3 - Vertical Retreat Mining Diagram...........................................................7 Figure 4 – Overhand Cut & Fill Diagram ...............................................................9 Figure 5 - Underhand Cut & Fill Diagram............................................................10 Figure 6 - Post Pillar Diagram.............................................................................12 Figure 7 - Room and Pillar Diagram....................................................................14 Figure 8 – Haulage Cross Section ......................................................................17 Figure 9 – ST1030 Dimensions ..........................................................................18 Figure 10 – ST1030 Turning Radius ...................................................................18 Figure 11 – Recommended Drift Size .................................................................18 Figure 12 - Factor A Strength/Stress ..................................................................22 Figure 13 – Rock Defect Orientation Number .....................................................22 Figure 14 - Design Surface Factor ......................................................................23 Figure 15 - Modified Stability Graph....................................................................24 Figure 16 - General Diagram for Transforamtions ..............................................27

List of Tables Table 1 - Infrastructure Sizing.............................................................................15 Table 2 - Ramp Grades ......................................................................................19 Table 3 - Equipment Selection............................................................................25 Table 4 - Standard Orepass Design....................................................................26

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1 Mining Methods Production can be generalized into two categories, namely selective and bulk mining. The nature of the deposit will guide the design towards the most appropriate method. As a general rule, bulk mining is best suited to larger deposits which yield a lower operating cost whereas selective mining is more commonly used to recover higher grade deposits in order to improve recovery and dilution.

Depending on the orebody characteristics, drifts can be driven either longitudinally or transversely. Each mining method describes these methodologies. The following methods are discussed in detail below:

• Sub Level Caving • Block Caving • VRM • Cut & Fill • Underhand Cut & Fill • Post Pillar • Room and Pillar

Refer to existing MTS Guidelines for additional information.

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1.1 Sub Level Caving The sublevel cave (SLC) mining method is suited to a variety of orebody shapes and ground characteristics. It has particular advantages in low grade ore. Stobie Mine is a noteworthy Vale operation using SLC mining. SLC mining is suited to steeply dipping ore, as well as flatter lying orebodies having a large vertical dimension. The integrity of the ore and overlying waste must meet certain criteria in order for the method to be viable. The ore should be strong enough to:

• Allow drifts to stand with minimum support due to the large amount of development required in ore

• Allow pre-drilled production holes to stay open until blasted • Keep a good brow during blasting

The overlying waste should be weak enough to cave on top of the ore as the ore is blasted and removed. Controlling dilution and ore losses is essential to the practical application of SLC mining. Ore is graded in the drawpoint and an economic cut-off for mucking is established. For these controls, the ore should be easily separated from the rock, visually or otherwise, or else have a vague ore boundary due to mineralization extending into the wall rock.

Figure 1 - Sublevel Cave Diagram

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Advantages

• Repetitive unit mining processes of development, drilling and blasting • Early production • Highly adaptable to mechanization • Ore is recovered in one operation - no pillars are left • Long term ground support in the crosscuts is not required • Ideally suited to low grade ore • Flexibility to mine around rock inclusions and previously mined out areas • High rate of production is achieved due to the number of producing drifts • No backfill required

Disadvantages

• The main disadvantage of SLC mining centres around the controls required to minimize dilution and maintain good recovery. The constant presence of rock requires strict quality control of the grade during the extraction process. Drift and drilling layouts must be made with dilution and recovery as guiding criteria

• A strict sequence of mining a group of parallel drill drifts must be followed. If the drill drifts on a sublevel are not retreated in unison so the free faces line up to form a common production front, more dilution and less recovery will occur

• The high amount of development, both in ore and waste, is greater than other mining methods. This is also true for waste development on each sublevel and access ramps between sublevels. Planning must allow for substantial rock hoisting and/or storage

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1.2 Block Caving Block-caving is a large scale production mining method applicable to low grade, massive orebodies with: large dimensions both vertically and horizontally; a rock mass that behaves properly, breaking into blocks of manageable size; and a ground surface which is allowed to subside.

Figure 2 - Block Caving Diagram

Advantages

• It is reasonably inexpensive underground method since drilling, blasting, ground support, and labor costs are reduced

• Ventilation is generally simplified • Production levels can approach those of large open pits • Low-grade ore bodies can be more economically mined • Grade control through drawpoints aids mine planning

Disadvantages

• It is accompanied by major surface subsidence • Development time and money can be excessive • Drift and drawpoint maintenance is expensive • Ground conditions due to weight can damage or shut down production

drifts • Draw control is a constant and rigorous endeavor

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1.3 Vertical Retreat Mining Vertical retreat mining (VRM) is a bulk method of mining used in most of Vale’s underground mines. It is a versatile method that can be used for longitudinal or transverse mining applications. VRM is best suited for ore dipping at 55° or greater and may be used in a variety of ground conditions, where the drilled holes maintain their integrity.

Figure 3 - Vertical Retreat Mining Diagram

Advantages

• Versatile system that can be used for longitudinal or transverse stope and pillar recovery

• Safety is improved because all loading operations are carried out from the same topsill location. No open hole conditions exist in the panel

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while loading is being done, with no exposure to open ground conditions

• More costly operations, such as raise boring and conventional raising for establishing slots, are eliminated, as in regular blasthole

• Broken muck left in the stope during mining provides wall support • Improved fragmentation, with larger blasts being attained relatively

quickly • Cycle time is improved, since no additional ground support systems

are required during blasting • Lower costs over conventional methods, such as Mechanized Cut and

Fill • Efficiency is better than Mechanized Cut and Fill • Dilution can be minimized by selectively leaving low grade areas

unblasted

Disadvantages • Close and continuous engineering control is required to plan and

design each crater blast. • Hole deviation and inaccurate drill set-up can cause excessive burden

in toe spacing. • Holes can become plugged during the blasting operations. • Tight collar spacing can cause problems in the top sill area such as

floor blow back and back problems requiring conditioning. • Holes can start to cone if the toe spacing becomes excessive or

charges are placed too high. • Holes may have to be re-drilled or cleaned. • Non-breakthrough holes can have excessive toe burden and attempts

must be made to locate toes (Boretrack, etc.). • High grade broken muck may oxidize. • Requires fill preparation when compared to regular open stope

preparation, and production rates are less than regular large blasthole methods.

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1.4 Cut & Fill

1.4.1 Overhand Overhand cut and fill is a method of mining in which support for the walls is provided by fill. In general, cut and fill is used under conditions where irregular ore, flat dip, weak hangingwall, weak footwall, or weak pillars would make open stope mining uneconomical because of ore losses or dilution.

The fill is placed as an integral part of the mining sequence and can be rock, sand (alluvial or mill tailings), gravel, hydraulic sand, or cemented hydraulic sand (P.P.C. slag). Generally, the orebody is excavated in small sections which are filled completely, or in part, before the next section is mined. Depending on the size of the orebody, this method can be used for either transverse or longitudinal mining.

Figure 4 – Overhand Cut & Fill Diagram

Advantages

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• Easily mechanized • Early production from start of development • Selective and flexible in mining irregular ore and sorting rock • Good labour efficiency in tons per manshift • Provides good ground support

Disadvantages

• The mining and filling cycles limit the maximum production rate • All work is done under an open back • Wall and back conditioning on each cut uses large amounts of labour and

materials • If stoping blocks have "Manway Only" access, all equipment becomes

"captive" • Ore tied up in pillars for later mining often using a more expensive

underhand mining method • Crown left below the above level for later mining, often using a more

expensive underhand mining method

1.4.2 Underhand Cut & Fill Undercut and fill mining is a system of fill method mining where each cut is mined below the last tightly poured cut of consolidated fill. The fill is maintained in position by the assistance of a suitable mat structure which is constructed before the fill is poured.

Figure 5 - Underhand Cut & Fill Diagram

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Advantages • Allows the mining of very weak ore • Allows the mining of remnant ore • Low development cost

Disadvantages

• Requires experienced miners and supervision • Slow production rate • Low efficiency in tons per manshift • High cost per ton • Difficult to mechanize

Undercut and fill mining is normally used for the recovery of good grade remnant ore where location, weak ground conditions, or risk of dilution will not allow the use of any other more economical mining method. Almost any shaped remnant can be recovered using this method by mining and filling in appropriate width slices.

Pillar remnant services are normally established in two ways:

a) A chute is raised adjacent to an existing manway. b) A raise borer chute is cored on centre line of pillar and a service

manway is carried down over the chute as mining progresses.

The detailed descriptions and calculations in this guideline are based on the above service systems used in slice widths of 11, 16 and 22 feet.

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1.5 Post Pillar The post pillar cut and fill mining method can be used to extract ore from irregular as well as regular shaped ore bodies in either a transverse or longitudinal direction depending upon the size and shape of the ore body. Sufficient size stoping blocks should be available to allow for full utilization of mechanized equipment while ground conditions must be competent enough to maintain stable ground conditions over a practical stope width as mining progresses.

Figure 6 - Post Pillar Diagram

Advantages

• Low dilution • Grade control facilitated • Only primary mining (no secondary recovery) • Less reserves tied up in pillars than in regular cut and fill, therefore higher

productivity mining maintained for a longer period of time • Mining methodology and equipment similar to that used in development

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• Internal access ramp in ore possible - Retreat 'up ramp' with associated services

• Backfill quality not as critical as some methods • Good fragmentation • Multiple faces available

Disadvantages

• Overall ore recovery probably lower since posts are not recoverable • Possible excessive open ground (based on stope/pillar ratio) to reflect

acceptable recovery ratios • Auxiliary ventilation more complicated in wider areas • Requires close technical control by surveying and geology departments

(very sensitive to operations quality control)

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1.6 Room and Pillar Room and pillar is designed for mining of flat, bedded deposits of limited thickness. Examples are sedimentary deposits, like copper shale, limestone or sandstone containing lead, coal seams, salt and potash layers, limestone and dolomite.

Figure 7 - Room and Pillar Diagram

Advantages

• The nature of the applicable ore bodies lends itself to a high degree of mechanization and labor efficiencies

• Extreme flexibility exists, and rapid changes can be made when required • Grade control can be maximized by adjusting drift dimensions and leaving

support pillars in low grade or waste • Productivity can approach or exceed block caving system if the ore

horizons are wide and high • Method can be made very safe and training due to repetition is facilitated

Disadvantages

• Ground swelling and movement as large areas are opened creates difficulties

• Roof and back maintenance can become excessive if larger open areas are required

• To provide necessary roof support, high-grade pillars must at times be left • Capital intensive due to mechanized equipment requirements

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2 Business Planning Design Standard

The Business Planning Design Standards are based on the best designs from Ontario Division mines, subject to meeting or exceeding the Ontario Mining Regulations, of actual infrastructure including electrical sub-stations, fuel, diamond drill and refuge stations, sumps, powder & fuse magazines, remucks and loadouts.

2.1 Infrastructure Sizing Type of Excavation Size (Width x Height x Length) Notes

X-Cut 18' x 15' x Dependent Length varies according to ore outline

Sill Drift 16' x 16' x Dependent Length varies according to ore limits

Ramp 16' x 16' x Dependent Length varies according to the purpose

Fuse Mags 16' x 16' x 16' Average typical length

Loadouts 16' x 20' x 55' Average typical length

Remuck 22' x 16' x 50' Average typical length

Powder Magazines 22' x 16' x 50' Height of 20 if crane is required

Stations

Fuel Station 24' x 16' x 75' Average typical length

Electrical Substation 22' x 16' x 50' Average typical length

Refuge Station 16' x 16' x 55' Depends on the number of people in area

Sumps

Sumps - Level 22' x 18' x 36' Average typical length

Sumps - Main 22' x 20' x 120 A two basin system for clean/dirty water

Diamond Drilling

Diamond Drill Station - Exploration 16' x 22' x 42' Width from MTS Guideline and height

from actual

Diamond Drill Station - Definition 20' x 16' x 32' Average typical length

Table 1 - Infrastructure Sizing *For a list of infrastructure sizing for every mine, please click here*

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2.2 Development

2.2.1 Design To be able to select the location, type and size of drift required, the following basic criteria is considered.

• Is the drift required for track haulage or is it for trackless equipment? • What is the largest piece of equipment travelling through the drift? • What is the size of the vent pipe that is to be hung from the back;

(clearance problem)? • Ground conditions relative to size. • What other purposes is this drift to be used for; ventilation, services,

transfer, etc. • Duration that the drift will be required. • Initial and ultimate function.

2.2.2 Location of Drifts When laying out drifts, the following should be kept in mind:

• Any drift that is to exist for more than two years after production from the area starts is considered capital.

• Any drift that provides short term access, cuts across the ore body, and is an intricate part of the drilling and blasting and primary haulage phase is considered operating.

• Footwall haulage drifts, not in the orebody, are generally at least 50 feet from the orebody contact.

• The span of drift intersections should be kept to a maximum diameter of 24 feet. Cable bolting is required, in addition to the primary bolting over 24 feet.

• Where possible four-way intersections should be avoided to reduce the span of the intersections opening. Hence, stagger the drift take-offs.

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2.2.3 Drift Sizing Info All haulage and access drifts used in mechanized blasthole mining development must have enough clearance as indicated in the figures below: Things to Consider:

• Support or overhead installation to provide the vehicle / operator with adequate clearance

• Wall to wall minimum clearance of 5 ft wider than the maximum width of a motor vehicle using the haulageway

Refer to regulation 854, sections 2.4.1 and 2.4.2.

Figure 8 – Haulage Cross Section

Figure 9 – Scooptram Turning Radius

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Figure 9 – ST1030 Dimensions

Figure 10 – ST1030 Turning Radius

Figure 11 – Recommended Drift Size

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2.2.4 Ramp Grades Phase 1& 2 The assumptions are:

• The ramp only flattens out at the subsequent level • Max ramp grade is 15%

Level Access

Interval Ave Ramp Grade Total Driving

Footage 50' 13.30% 377' 100' 14.10% 717' 150' 14.40% 1,052' 200' 14.50% 1,390‘

Table 2 - Ramp Grades

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3 Mineable Stope Shape

The standardized process of designing stopes for a phase I and II project is using Mineable Shape Optimizer (MSO). The MSO will assist Business Planning Long Range Planners in designing stopes very quickly for an initial high level assessment in order to give an estimation of mineable tonnage and grades. MSO can be run to analyze the effect of different parameter settings, such as cut off value, cut off grade and stope dimensions.

MSO searches for the optimal shape by taking into account the orebody geometry. It generates strings on sections using the block model and a cut off value, just as a planner would do, and creates wireframes shapes for evaluation against the block model. A similar product is the Mineable Reserve Optimizer (MRO). MRO works by identifying and evaluating mineable envelopes of material taking into account factors such as the minimum size and shapes of the stopes or mining faces. MSO results should be compared with MRO results, so before starting, the planner should send a request to a Senior Mine Geologist at each site or MMTS to run the MRO on the final Block Model.

MSO is a module within Datamine Studio 3 and the manual can be located through the following link, Stope Design Using MSO.

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4 Ground Control

In the planning stages of a mining block, a ground classification system is necessary to determine mining methods, pillar sizes, drift, station and pass locations, etc. This is the stage where the R.Q.D system finds its best application.

In the actual development stages, where the planning is complete and forewarning of problems in the short term are necessary for safety and proper support or reinforcement selection. This type of problem is solved by detailed process and general classification of R.Q.D.’s may be of secondary importance. This is because R.Q.D. does not account for certain parameters, and only partially takes into account other factors that affect rock quality. The more important of these are:

- Internal rock strength (usually compressive strength); - Joint filling (i.e. gouge, slickensided, etc.); - Joint orientation (attitude); - Continuity of Joints and - Water flow.

These geotechnical parameters are combined in the form of a stability number (N) and are plotted against an excavation factor size (hydraulic radius m). The stability number accounts for the rockmass quality, the state of stress, and the orientation of the exposed surfaces. The excavation factor accounts for the shape and size of the opening. To provide an indication of the maximum span that could be opened up along the hangingwall, the following method should be used: Stope Stability Number: N= Q’ * A * B * C Where:

- Q’ is the modified NGI rockmass rating = RQD/Ja * Jr/Jn, they are derived from the tables published by Barton.

Ja = Joint Alteration Number Jr = Joint Roughness Number Jn = Joint Set Number

- A is the stress factor defined as the ration of the uniaxial compressive strength to induced stress in the hangingwall and footwall (Refer to figure 12);

- B is the rock defect orientation number which accounts for the presence of persistent structure parallel to, or intersecting, the exposed surface (Refer to figure 13);

- C is a design surface orientation factor which accounts for the dip of the hangingwall (Refer to figure 14).

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Figure 12 - Factor A Strength/Stress

Figure 13 – Rock Defect Orientation Number

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Figure 14 - Design Surface Factor

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Stable Strike Length (Hydraulic Radius): Hydraulic Radius m = (Strike Length) x (Vertical Lift) 2 (Strike Length + Vertical Lift) The stope stability number is plotted against the hydraulic radius on the stope stability graph determined by Potvin et al (1988) to indicate the area in which unstable stope surfaces are plotted. The following is the stability graph by Potvinat et al:

Figure 15 - Modified Stability Graph

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5 Equipment Selection

A list of equipment was compiled from all mines to be used as a standard list for a phase 1 and 2 designs. The following is a list that can be used as a guide for design for Vale Sudbury Basin:

EQUIPMENT

TYPE HP CFM POWER REQUIRED

(kW) Man Carrier 134 13,400 N/A

Forklift 49 4,900 N/A Scissor Truck 147 14,700 N/A Anfo Loader 174 17,400 N/A

Grader 135 13,500 N/A Boom Truck 225 22,500 N/A

ITH Compressor 75 7,500 56 ITH 100 10,000 74.6

Boart 60 6,000 44.7 1 Boom Jumbo 87 8,700 75 2 Boom Jumbo 155 15,500 115.6 Maclean Bolter 154 15,400 112.6

Loco 55 5,500 N/A SCOOPS

3.5 113 11,300 N/A 6 310 31,000 N/A 8 325 32,500 N/A

10 350 35,000 N/A TRUCKS 30 Ton 408 40,800 N/A 40 Ton 472 47,200 N/A 45 Ton 589 59,900 N/A

Kiruna (55 Ton) 600 18,000 800 Table 3 - Equipment Selection

*For a list of all mine equipment list, please click here*

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6 Ventilation

Ventilation systems are a major asset with respect to mining and key to ensuring a healthy, safe and productive environment to ensure business goals are met. The need for careful planning and the consideration for proper design with respect to location and maintenance are critical for optimization of the system.

Certain projects require a quick evaluation (of design) to estimate capital cost, including fans required and vent raises. For these projects, Mine Site Standards can be used to design vent systems and vent raises. For a more detailed analysis, please refer to the Ventilation Tool Box. This tool box will provide guidance to planners with respect to ventilation criteria for main and auxiliary system components as well as services including locations and requirements for emergency facilities and infrastructures.

7 Orepass Design

Raises are, inclined to vertical, openings that are excavated by a number of ways, and are required for various reasons such as ore/rock and slot for production blasting. Raises can be of various sizes and shapes; however, a standard size commonly used around the mines is as follows:

Standard Orepass Design:

Size 10' x 10' Typical Dip (in degrees) 70-85 Depends on contractor Ft/day 6 On average Maximum Length 1000' *Most Economical

Table 4 - Standard Orepass Design *Need own air line, power line, transportation time gets higher and survey

control gets harder after 1000 feet.

For a quick evaluation of design and financials, standard orepass dimension can be used or else, one should refer to the MTS Guidelines in section 11.11 below.

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8 Power

Switchrooms are intended to provide controlled access to permanently installed electrical power control devices and electronic communications equipment. This equipment includes switches, starters, breakers, transformers, panel boards, relay panels…down to capacitors. For a detailed study, the electrical department should be referred to in order to properly layout power distribution systems.

However, for a phase 1 financial analysis, indications that a transformer will be required should be sufficient in estimating capital cost and evaluating the project. The need for a transformer can be assessed by adding all equipment, pumps and fans HP’s on a level to what a transformer can support (a maximum of 1000HP). If the requirements are over 1000 HP, it’s a good indication that two transformers will be required or an upgrade to a 13.8kV transformer. The following diagram is a general layout for a typical transformers station:

Figure 16 - General Diagram for Transformers

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9 Cost estimating

The Mine Design group has multiple facets within their project that need to have costs associated with them. Typically these are capital costs, such as rock development, electrical stations, sumps, refuge stations, etc.

Rock development will typically be the highest capital cost and these costs differ from mine to mine. The Mines development costs can be found in the Design Group Costs excel sheet. This link also contains the costs of purchasing equipment, infrastructure and personnel.

Using this cost estimating tool the designer can acquire the costs pertaining to their project and consequently do the financial evaluation of the project.

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10 Backfill

Follow the chart below to determine the scale of study required from the Backfill Department. Most of the design can be done rather quickly due to the project being part of an existing mine, if not, a detailed analysis may be required. Please refer to section 12.3 for quick rule of thumb.

11 Regulations

System's purpose is to standardize the best mine designs and practices, subject to meeting or exceeding the Ontario Mining Regulations. For convenience, some of the most commonly referred to regulations are listed below. Note: Regulations found in this document were valid for 2010. Refer to current regulations for possible updates. http://www.e-laws.gov.on.ca/html/regs/english/elaws_regs_900854_e.htm

11.1 Trackless Haulage Clearances Reg. 854, s.112 A haulageway used by motor vehicles, other than Motor vehicles running on rails, shall:

a) Except where pedestrian traffic is effectively prevented, be at least 1.5 metres wider than the maximum width of a motor vehicle using the haulageway; and

b) Where it is regularly used by pedestrians and it is less than two metres

wider than the maximum width of a motor vehicle using the haulageway, have safety stations as prescribed in section 109 at intervals not exceeding thirty metres.

11.2 Overhead Clearance Reg. 854, s.113

Except in an underground mine with a low clearance roof in which equipment designed to be operated herein is used, a haulageway used by a motor vehicle shall have sufficient clearance below the roof, support or overhead installations to enable the operator of a motor vehicle to sit erect at all times.

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11.3 Refuge Station Reg. 854, s.26:

Where the procedure in case of a fire in an underground mine provides for the use of a refuge station for workers, the refuge station shall,

a) Be constructed with materials having at least a one hour fire resistance rating;

b) Be of sufficient size to accommodate the workers to be assembled therein; c) Be capable of being sealed to prevent the entry of gases; d) Have a means of voice communication with the surface; and e) Be equipped with a means for the supply of,

i. compressed air, and ii. potable water

11.4 Powder Magazines

Reg.272/97, s.22

A magazine, storage container or explosive storage area that is in an underground mine shall be

a) Located at least 60 metres from i. the main access into or from a mine ii. key mechanical and electrical installations that remain in service

during a mine emergency iii. areas of refuge or other areas where workers may congregate

and iv. storage areas for fuels or other potential sources of fire

b) Located and designed to protect explosives from vehicle impact and

vehicle fires; and c) Conspicuously marked by a “DANGER EXPLOSIVES” sign.

11.5 Fuel Bay Reg.854, s.120 A service garage or fuelling station in an underground mine shall,

a) Be designed and protected to prevent inadvertent entry of an uncontrolled motor vehicle;

b) Be located so that in the event of a fire or explosion in the garage or station there will be a minimum effect on working areas of the mine or on

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underground installations including shafts, magazines, refuge stations, transformer installations and other installations;

c) Have a concrete floor without service pits in the floor; and d) Be equipped with a system to contain spills of oil and grease.

Note - Refer to ventilation guidelines for design criteria.

11.6 Escapeway Reg 854, s.37

a) Except during the initial stages of exploration and development of mine, in addition to the opening through which workers are let into or out of the mine and the ore extracted, a separate escapement exit shall be provided.

b) The escapement exit required by subsection (1) shall be, i. Located more than thirty meters from the main hoisting shaft or

ramp; ii. Of sufficient size to afford an easy passageway iii. Where necessary provided with ladders from the deepest

workings to the surface iv. Marked on all levels by signs and arrows pointing the way of exit

in a manner to expedite escape v. Made known to all underground workers who shall be instructed

as to the route to the escapement exit and vi. Inspected at least once a month by a competent person who

shall give a written report of such inspection to the supervisor in charge of the mine.

c) A structure covering the escapement exit shall be constructed of material with at least a one hour fire resistance rating

11.7 Boundary Pillars (U/G) Reg. 854, s.19

a) Subject to subsection (2), a pillar sixty thick shall be established on either side of a party boundary between adjoining underground mining properties.

b) Except for exploration headings and diamond drilling, before the pillar is mined, drawings , plans, specifications, mining methods and procedures for the mining of the pillar shall be prepared or checked by a professional engineer in accordance with good engineering practice, filed with the owners of adjoining mining properties and kept readily available at each mine site. O.Reg. 272/97, s.7(1)

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c) The drawings, plans, specifications, mining methods and procedures to be filed shall be maintained and kept up to date in accordance with subsection 29 (2) of the Act.

d) The pillar dimensions and mining methods and procedures shall, i. Provide ground support to control rockbursting, ground falls or

pillar failures; and ii. Withstand inrush of water or waterbearing materials across the

party boundary. R.R.O. 1990, Reg. 854, s.19(3,4) e) Subject to subsections (2), (3) and (4), the party boundary pillar may be

mined if the owners of the adjoining mines agree. O.Reg. 272/97, s7(2)

11.8 Property Boundaries Reg. 854, s 92.

a) Subject to subsection (2), i. Where earth, clay, sand or gravel is being removed from a

surface mine no mining operations shall be carried on within a distance from the property boundary of half the total depth of the surface mine and earth, clay sand or gravel that sloughs from within this distance shall not be removed; and

ii. Where metallic or non-metallic rock is being removed from a surface mine, no mining operations shall be carried on within a distance of six meters from the property boundary.

b) (2) Adjoining owners may, by agreement in writing, waive the provisions of

subsection (1). R.R.O. 1990, Reg 854, s. 92.

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12 Rules of Thumb

The following rules of thumb are exerts from the McIntosh Engineering “Hard Rock Miners Handbook Rules of Thumb” and the existing MTS Guidelines. Exerts taken from the McIntosh Engineering are referenced below by chapter and section between the brackets.

12.1 Passes, Bins, and Chutes • Ore passes should be spaced at intervals not exceeding 500 feet (and

waste passes not more than 750 feet) along the draw point drift, with LHD extraction. (24.05)

• The best inclination for an ore pass in a hard rock mine is 70 degrees from the horizontal. (24.06)

• The minimum inclination for a short ore pass is 50 degrees from the horizontal. For a long pass, it is 55 degrees. (24.07)

• Ore passes cannot be employed to any advantage where the ore dips shallower than 55 degrees from the horizontal. (24.08)

12.2 Mining Methods • Ore will not run on a footwall inclined at less than 50 degrees from the

horizontal (3.02) • Even a steeply dipping ore body may not be drawn clean of ore by gravity

alone. A significant portion of the broken ore will inevitably remain (“hang”) on the footwall. If the dip is less than 60 degrees, footwall draw points will reduce, but not eliminate, this loss of ore (3.03)

• The number of stopes developed should normally be such that the planned daily tonnage can be met with 60% to 80% of the stopes. The spare stopes are required in the event of an unexpected occurrence and may be required to maintain uniform grades of ore to the mill. This allowance may not be practical when shrinkage is applied to a sulfide ore body, due to oxidation (3.04)

• In any mine employing backfill, there must be 35% more stoping units than is theoretically required to meet the daily call (planned daily tonnage) (3.05)

• Blasthole (longhole) Stoping may be employed for ore widths as narrow as 3m (10 feet). However, this narrow a width is only practical when there is an exceptionally good contact separation and a very uniform dip (3.06)

• Footwall drifts for blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth, it should be not less than 30m (100 feet) (3.08)

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• A ton of ore left behind in a stope costs you twice as much as milling a ton of waste rock (from dilution). Source: Peter J. George (3.09)

12.3 Backfill

• If a mine backfills all production stopes to avoid significant delays in ore production, the daily capacity of the backfill system should be should be at least 1.25 times the average daily mining rate (expressed in terms of volume) (21.04)

• The capital cost of a paste fill plant installation is approximately twice the cost of a conventional hydraulic fill plant of the same capacity (21.03)

• Typical costs of backfill range between 10 and 20% of mine operating cost and cement represents up to 75% of that cost (21.02)

• The typical requirement for backfill is approximately 50% of the tonnage mined. It is theoretically about 60%, but all stopes are not completely filled and tertiary stopes may not be filled at all (21.05)

• It is common to measure the strength of cemented backfill as if it were concrete (i.e. 28 days), probably because this time coincides with the planned stope turn-around cycle. Here it should be noted that while concrete obtains over 80% of its long- term strength at 28 days, cemented fill might only obtain 50%. In other words, a structural fill may have almost twice the strength at 90 days as it had at 28 days (21.06)

• The quantity of drain water from a 70% solids hydraulic backfill slurry is only one-quarter that resulting from one that is 55% solids (21.07)

• Hydraulic backfill has porosity near 50%. After placement is completed, it may be walked on after a few hours and is “trafficable” within 24 hours (21.08)

• It takes two pounds of slag cement to replace one pound of normal Portland cement. In other words, HF with 3% normal cement and 6% slag cement will exhibit the strength characteristics of one with 6% normal cement alone (21.09)

• A 6% binder will give almost the same CRF strength in 14 days that a 5% binder will give in 28 days. This rule is useful to know when a faster stope turn-around time becomes necessary. Cemented Rock Fill (21.11)

• As the fly ash content of a CRF slurry is increased above 50%, the strength of the backfill drops rapidly and the curing time increases dramatically. A binder consisting of 35% fly ash and 65% cement is deemed to be the optimal mix (21.12)

• The strength of a cemented rock backfill may be increased 30% with addition of a water reducing agent (21.13)

• The size of water flush for a CRF slurry line should be 4,000 US gallons (21.14)

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• The optimum W/C ratio for a CRF slurry is 0.8:1, but in practice, the water content may have to be reduced when the rock is wet due to ice and snow content of quarried rock or ground water seepage into the fill raise (21.15)

• Experience to date at the Golden Giant mine indicates that only 46% of the tailings produced can be used for paste fill (21.18)

• The inclusion of the slimes fraction (“total tails”) means that at least some cement must always be added to paste fill. The minimum requirement to prevent liquefaction is 1½% (21.19)

12.4 Mine Dewatering • The average consumption of service water for an underground mine is

estimated at 30 US gallons per ton of ore mined per day. The peak consumption (for which the water supply piping is designed) can be estimated at 100 USGPM per ton of ore mined per day (20.01)

• Ore hoisted from an underground hard rock mine has moisture content of approximately 3% (20.02)

• The main pump station underground must have sufficient excavations beneath it to protect from the longest power failure. The suggested minimum capacity of the excavations is 24 hours and a typical design value is 36 hours (20.06)

• Allow one square foot of surface area/USGPM in the design of a settling sump (20.08)

• A tonne of water a second pumped up 100 m requires 1MW of power (20.13)

• A sump should have a live volume equal to at least 2½ times the pump operating rate to limit pump starts to six per hour (typical NEMA B motor). For example, the live volume of the sump for a 500 USGPM pump should be at least 1,250 gallons (20.15)

12.5 Lateral Development and Ramps • The overall advance rate of a lateral drive may be increased by 30% and

the unit cost decreased by 15% when two headings become available (11.02)

• The overall advance rate of a lateral drive will be increased by 2m/day when a second heading becomes available and an additional 2m/day with a third heading (11.03)

• Approximate productivity for driving trackless headings (drill, blast, scale, muck and bolt) is as follows: 0.3-0.5 m/manshift for a green crew; 0.7-0.8

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m/manshift for competent crews; and 1.0-1.25 m/manshift for real highballers (11.04)

• The minimum width for a trackless heading is 5 feet wider than the widest unit of mobile equipment (11.05)

• The back (roof) of trackless headings in hard rock should be driven with an arch of height equal to 20% of the heading width. Source: Kidd Mine Standards (11.06)

• The cost to slash a trackless heading wider while it is being advanced is 80% of the cost of the heading itself, on a volumetric basis (11.07)

• For long ramp drives, the LHD/truck combination gives lower operating costs than LHDs alone and should be considered on any haul more than 1,500 feet in length (11.08)

• LHD equipment is usually supplemented with underground trucks when the length of drive exceeds 1,000 feet (11.09)

• With ramp entry, a satellite shop is required underground for mobile drill jumbos and crawler mounted drills when the mean mining depth reaches 200m below surface (11.10)

• With ramp and shaft entry, a main shop is required underground when the mean mining depth reaches 500m below surface (11.11)

• A gradient of 2% is not enough for a horizontal trackless heading. It ought to be driven at a minimum of 2½% or 3% (11.12)

• Wet rock cuts tires more readily than dry rock. To prevent ponding and promote efficient drainage, trackless headings should be driven at a minimum gradient of 2½ - 3%, if at all possible (11.13)

• The minimum radius of drift or ramp curve around which it is convenient to drive a mobile drill jumbo is 75 feet (11.14)

• For practical purposes, a minimum curve radius of 50 feet may be employed satisfactorily for most ramp headings (11.15)

• Footwall drifts for trackless blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth it should be not less than 30m (100 feet) (11.17)

• Ore passes should be spaced at intervals not exceeding 500 feet (and waste passes not more than 750 feet) along the draw point drift, with LHD extraction (11.18)

• The maximum practical air velocity in lateral headings that are travelways is approximately 1,400 fpm (7 m/s). Even at this speed, a hard hat may be blown off when a vehicle or train passes by. At higher velocities, walking gets difficult and road dust becomes airborne. However, in pure lateral airways, the air velocity may exceed 3,000 fpm (11.19)

• In practice, the maximum air velocity found employed in lateral headings used for two-way trackless haulage seldom exceeds 1,000 fpm (5 m/s) (11.21)

• Typical gradients for track mines are 0.25% and 0.30% (11.27)

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• A minimum clearance of three feet should be designed between the outside of the rails and the wall of the drift to permit safe operation of a mucking machine when driving the heading (11.28)

12.6 Mine Layout • For pit haul roads in general, 10% is the maximum safe sustained grade.

For particular conditions found at larger operations, the grade has often been determined at 8%. It is usually safe to exceed the maximum sustained grade over a short distance (4.03)

• The maximum safe grade over a short distance is generally accepted to be 15%. It may be 12% at larger operations. Source: (4.04)

• The maximum safe operating speed on a downhill grade is decreased by 2 km/h for each 1% increase in gradient (4.05)

• Each lane of travel should be wide enough to provide clearance left and right of the widest haul truck in use equal to half the width of the vehicle. For single lane traffic (one-way), the travel portion of the haul road is twice the width of the design vehicle. For double lane (two-way), the width of roadway required is 3½ times the width of the widest vehicle (4.06)

• To avoid a collision caused by spinout, the width of an open pit haul road should equal the width plus the length of the largest truck plus 15 feet safety distance (4.07)

• Footwall drifts for blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth it should be not less than 30m (100 feet) (4.22)

• Ore passes should be spaced at intervals not exceeding 500 feet (and waste passes not more than 750 feet) along the footwall drift, when using LHD extraction (4.23)

• The maximum economical tramming distance for a 5 cubic yard capacity LHD is 500 feet, for an 8 cubic yard LHD it is 800 feet (4.24)

• The amount of pre-production stope development required to bring a mine into production is equal to that required for 125 days of mining (4.25)

12.7 Cost Estimating • Haulage costs for open pit are at least 40% of the total mining costs;

therefore, proximity of the waste dumps to the rim of the pit is of great importance (8.09)

• The economical tramming distance for a 5 cubic yard capacity LHD is 500 feet and will produce 500 tons per shift, for an 8-yard LHD, it is 800 feet and 800 tons per shift (8.08)

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12.8 Misc

• When driving development towards ore, the last 30’ of rock before ore contact is considered operating.

12.9 Refuge Station

• Each person requires 10 cu. yds of air to sustain them for a period of 8 hours. Hence, design the station to accommodate the total cu. ft. of air required after determining the practical minimum dimensions

• The refuge station is designed to accommodate the practical maximum number of individuals that could be in the area in the event of a disaster, such as a fire

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12.10 Raise Bore Station

• This guideline will give the information required to design a raise bore top station for the 64R and 71R (D.C.) raise bores equipped with a rod handler.

• For station sizing requirements refer to Diamond Drill and Raisebore • The minimum length of this straight drift is dependent on the length of the

hole to be drilled.

12.11 Existing Guidelines

MTS Guidelines Section PagesNotice - Read this first 1 Credits 2 01 - Introduction 79 02 - Geology 01- Diamond Drilling 50 02- Mapping and Plotting 16 03- Mineral Resource Inventory 41 04- Grade Control 17 05- Mineability and Dilution 12 06- Ground Control 8 07- Ore Contact Probing 6 03 - Development 92 04 - Rock Mechanics / Ground Control 134 05 - Drilling 40 06 - Loading And Blasting 31 07 - Mining Methods 202 08 - Backfill 56 09 - Mine Services 178 10 - Project Management 14 11 - Mine Closure 25

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13 References

• “Occupational Health and Safety Act and regulations for Mines and Mining

Plants”, 2007

• “SME Mining Engineering Handbook”, 2nd edition, Volume 2, senior editor

Howard L. Hartman, 1992

• “Mining Methods in Underground Mining”, 3rd edition, by Atlas Copco,

2008

• “Loading and Haulage”, 4th edition, by Atlas Copco, 2007

• “Hard Rock Miners Handbook Rules of Thumb”, 3rd edition, by McIntosh

Engineering, June 2003

• MTS Guidelines, http://sud-ccmines2/guidelines, October 10, 2006

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