Underground Mining [PDF]

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Underground Mining



Contents MODULE ONE: ............................................................................................................................................................ 3 MINE ACCESS ............................................................................................................................................................. 3 1.0 Introduction .........................................................................................................................................................3 1.1 Design of Mine shafts ..........................................................................................................................................3 1.2 Purpose of a Shaft ................................................................................................................................................4 1.3 Shaft Cross Sections ............................................................................................................................................4 1.4 Determining Shaft Size ........................................................................................................................................4 1.5 Choosing the Right Shaft .....................................................................................................................................6 1.6 Shaft Comparison ................................................................................................................................................7 1.7 Types of Shafts .................................................................................................................................................. 10 1.8 Lateral Development and Ramps ....................................................................................................................... 10 1.9 Design and Function of Lateral Headings ......................................................................................................... 11 1.10 Track versus Trackless .................................................................................................................................... 12 1.11 Shaft sinking .................................................................................................................................................... 12 1.11.1 Applications of Shafts ............................................................................................................................... 12 1.11.2 Activities required for Shaft Sinking ............................................................................................................ 13 1.11.3 Special methods for shaft sinking. ................................................................................................................ 14 Impact on surrounding infrastructures ..................................................................................................................... 15 Appropriate soil conditions ...................................................................................................................................... 15 1.12 Site layout requirements .................................................................................................................................. 15 Resources required: (cost estimation for shaft, City of Edmonton,2005) ................................................................ 16 Human resources ................................................................................................................................................. 16 Equipments and machines ................................................................................................................................... 16 1.13 Collars & Portals.............................................................................................................................................. 16 1.14 Ground Freezing .............................................................................................................................................. 17 1.15 Shaft Sinking ................................................................................................................................................... 17 MODULE TWO .......................................................................................................................................................... 19 UNDERGROUND MINE SELECTION AND PLANNING ...................................................................................... 19 2.0 Mine Planning – Some Considerations .............................................................................................................. 19 2.1 Underground or Surface .................................................................................................................................... 19 Underground methods ................................................................................................................................................. 19 Surface methods........................................................................................................................................................... 20 2.2 Preliminary Strata Mechanics ............................................................................................................................ 20 2.3 Portal/ Access Location ..................................................................................................................................... 24 2.4 Types of Portals ................................................................................................................................................. 25 2.5 Mining Method Selection ................................................................................................................................. 26 2.6 Bulk Methods .................................................................................................................................................... 27 2.7 Selective Methods .............................................................................................................................................. 28 2.8 Dilution .............................................................................................................................................................. 29 2.9 Mine Planning.................................................................................................................................................... 30 2.10 Strategy for Underground Mines ..................................................................................................................... 30 2.10.1 Ramp Haulage .......................................................................................................................................... 30 2.10.2 Belt Conveyor ........................................................................................................................................... 31 2.10.3 Shaft System ............................................................................................................................................. 31 2.10.4 Conventional Methods of Ore Transport .................................................................................................. 31 2.10.5 Main Entry ................................................................................................................................................ 32 MODULE THREE ...................................................................................................................................................... 33 SELF-SUPPORTED MINING METHODS ................................................................................................................ 33 3.0 Room and Pillar Method .................................................................................................................................... 33 3.1 Non-coal Application of Room and Pillar Mining ............................................................................................. 47 3.2 Types of Room and Pillar Mining ..................................................................................................................... 50 3.3 Shrinkage Stoping .............................................................................................................................................. 54 Underground Mining| 1



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3.4 Sublevel Stoping ................................................................................................................................................ 60 3.5 Vertical Crater Retreat Mining .......................................................................................................................... 65 MODULE FOUR ........................................................................................................................................................ 68 SUPPORTED MINING METHODS .......................................................................................................................... 68 4.0 Cut-and-Fill method........................................................................................................................................... 68 4.1 Underhand and Overhand Methods ................................................................................................................... 71 4.1.1 Overhand cut-and-fill.................................................................................................................................. 72 4.1.2 Overhand stope ........................................................................................................................................... 73 4.1.3 Overhand stoping ........................................................................................................................................ 73 4.1.4 Underhand stoping ...................................................................................................................................... 73 4.2 Back filling Method ........................................................................................................................................... 73 4.2.1 What is backfill ........................................................................................................................................... 74 4.2.2 Mining methods using backfill ................................................................................................................... 75 MODULE FIVE .......................................................................................................................................................... 76 CAVING METHODS ................................................................................................................................................. 76 5.0 Longwall Mining ............................................................................................................................................... 76 5.1 The Longwall Mining Process ........................................................................................................................... 77 5.2 Sublevel Caving ................................................................................................................................................. 86 5.3 Block Caving ..................................................................................................................................................... 93



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MODULE ONE: MINE ACCESS 1.0 Introduction- Shafts Many underground operations consist of several tunnels acting as accesses, haulages, production levels, and airways but there is only a limited number of shafts that can be developed for any given ore body and these shafts must be sunk in the right place with the correct configuration to get optimum operational benefit. Despite the extensive and widespread use of shafts as a primary means of access to the ore body for many mineral resources in South Africa, there is little current and comprehensive reference material which provides a central body of knowledge spanning the various design and cost factors involved in shaft design. The purpose of this module is to cover some aspects which have to be considered when designing and sinking mine shafts. It is hoped that this module will be used as a quantitative basis for the comparison of various mine access options both for new and existing mines. This should form an integral part of any mine design process. 1.1 Design of Mine shafts This is a module that aims to cover some aspects, which have to be considered when designing mine shafts. No attempt will be made to cover the design of hoisting systems as it will be covered in Mine Transportation and Winding. Generally speaking, it is safe to assume that shafts are vertical or inclined openings sank into the earth’s crust in order to access mineral resources which are too deep to mine economically using open cut methods or adit systems. The mineral resource must be converted into a mineral reserve before one can start designing a mine shaft. This may sound very logical but, in the past, there has been situations where the shaft was sunk (on gut feel) first and the ore body found later with some pretty obvious economic consequences. An economically mineable reserve is therefore a pre-requisite for miners to be able to start on the design process. Since shafts play a major role in the general planning of mine development, their location is usually pre-determined. The location of a shaft can be changed when adverse geotechnical site conditions are encountered. The design of mine shaft is an iterative process, which requires several variables and options to be considered in order to arrive at an economic decision. The economic decision is arrived at by comparing the net present values (NPV) and internal rate of return (IRR) from the different options considered in the optimization process. The option with the most attractive financial option is then selected. The design parameters mentioned above include, but are not necessarily limited to the following: depth of shaft, ore and waste tonnage to be handled, shift handling (work force), materials handling, mining machinery handling, ventilation requirements, capital costs, operating costs, and of course the selling price of the mineral commodity.



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1.2 Purpose of a Shaft Shafts are generally used for the following functions:  To access an ore body  To transport men and materials to and from underground workings  For hoisting ore and waste from underground  To serve as intake and return airways for the mine (ventilation)  To provide a second egress (escape route) as required by mining law  Storage of nuclear waste 1.3 Shaft Cross Sections Rectangular Shafts Most shafts that were constructed in the 1900’s were of a rectangular cross-section because of the shape of the pieces of equipment that were taken down the shaft i.e. cages, skips, and counterweights were all square or rectangular in nature and so it made a lot of sense to sink or mine rectangular shafts. Breaking a square / rectangular shutter was however problematic and this slowed down the rate of sinking. Circular Shafts Almost all the hard rock mines now have circular shafts because the cross section provides good geometry for airflow and good rock support characteristics. The circular shutter is ease to move when doing concurrent lining resulting in faster work progress during sinking operations. This is an important aspect when it comes to the cash flow of the project. Elliptical Shafts Elliptical shafts were designed as an alternative to large circular shafts by simply adding half moons along the main axis. This had the effect of reducing the circular excavation and therefore the cost of sinking the shaft. 1.4 Determining Shaft Size As mentioned earlier the first step in determining the shaft size is to estimate the total reserves in the area to be exploited by the shaft. The reserve (ore body) size will govern the rate of mining and the mining rate will determine the tonnage (ore and waste) to be hoisted, the number of persons and material to be transported in a given shift. The foregoing is then used to determine the skip and cage sizes which in turn are used to calculate the total area required to accommodate these units. The shape and size of equipment to be taken down a shaft are also included in the calculation of the final shaft dimensions. The situation described here applies to rock, men, and materials shafts. This process is equally applicable to decline / incline shafts, the difference with



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the decline being trackless equipment and conveyors instead of the skips and cages. Inclined shafts also use monorails. Determining the rate of mining can be summarized as follows:  Identify possible mining layouts  Define standard mining block (stope or panel size) per layout  Calculate steady state conditions per level  Define steady state inputs/outputs requirements per level  Determine minimum access dimensions to cater for equipment and ventilation  Calculate development requirements to get to steady state  Simulate full level production from start of block to ore body extremity  Determine the maximum number of levels that will operate simultaneously  Estimate shaft size required to cater for sum of the requirements of the maximum number of working levels  Do an economic analysis (NPV & IRR)  Decide on optimum mining layout and shaft configuration Determining the size of a ventilation shaft can be summed up as shown in Table 1 below.



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By summing up the total intake air required for the complete mining system it is possible to compute the minimum ventilation shaft dimensions required to service the chosen mining system. 1.5 Choosing the Right Shaft The size or dimensions of each shaft will differ according to the intended duty for each unit. There are three types of shafts that are considered in this module, namely: vertical, decline and inclined shafts. There are many factors that influence the location and therefore the type of shaft to be sunk. Any mining technician should know that the shape, size and dip of ore body will dictate where the shaft position should be. In addition host ground conditions and water bearing structures also influence the final location of shafts. Generally speaking there are three types of ore deposits, that is: a) Narrow tabular deposits (steep & flat dipping – gold, platinum, etc) b) Wide tabular deposits (coal, potash) c) Massive deposits (copper, nickel, iron ore) All of the above deposits can be accessed by vertical shafts except for the flat narrow tabular deposits, which may be mined more economically using Decline or Inclined shafts. The chosen site of a shaft should be such that the best and most profitable use will be made of it when in full production. Criteria for choosing a Vertical Shaft A vertical shaft should be chosen under the following conditions:  Ore body should be steep dipping  Ideal for deep ore bodies  Provides quick access to ore body  Most economic hoisting method for depths exceeding 500m  Quicker return on capital investment Other considerations are that the shaft should be sunk near or close to the centre of gravity of the ore body. The shaft depth should be sunk to where most or all of the ore will be transported down grade to reach the shaft loading stations. The shaft should ideally cover a five-year production period i.e. minimum life of five years. For aesthetic reasons the shaft should be in a position where the head frame will be out of site of the general public although this really is not a major issue in developing countries. Criteria for choosing a Decline or Inclined Shaft A decline or inclined shaft should be considered under the scenarios:  Flat dipping ore body  Shallow ore bodies  Require high throughput Underground Mining| 6



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 Require low initial capital costs  Want to avoid some of the environmental concerns (headgear) The decline or inclined shafts are associated with grade constraints and one needs to take these into account. If a decline is used for trackless haulage then the maximum grade recommended is 8,0 . However, if used for conveyor belt haulage and the decline are used by rubber tired trackless equipment on regular basis, then 15,0 is the maximum recommended. If used for conveyor belt haulage only, then the maximum grade could be 15 - 25,0 depending on material to be conveyed. If equipment has to be driven up and down to clean spillage, this will limit the gradient. It is important to note that unfortunately there are no standard designs for circular concrete shafts. Typically each new shaft is designed from scratch to accommodate the particular requirements envisioned by the mine planners. Therefore each shaft is designed and constructed on a fit for purpose basis. 1.6 Shaft Comparison The discussion on design of mine shafts will be incomplete without making the necessary comparison between the different types of shafts. One needs to assess the selection criteria and compare the merits and demerits of each type as part of the design process. Table 2 below summarizes the selection criteria including the advantages and disadvantages of each shaft type.



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Table 3 below gives some indicative development costs. It is difficult to get the exact figure as costs are influenced by support applications, ground conditions and various other factors specific to each site. The numbers used here are the August 2007 costs using an exchange rate of R7.00 to the US dollar. These figures do not include shaft equipping.



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1.7 Modern Shafts Shafts sunk today in hard rock mines are mostly limited to the standard three-compartment timber shaft and the circular concrete shaft. The standard three-compartment timber shaft has two hoisting compartments that measure six feet by six feet inside the timbers. The third compartment, used for a manway and utilities, is sometimes slightly shortened from the six-foot width. These timber shafts are still contemplated for exploration entries in general and production shafts for small hard rock mines. For remote sites, the shaft timber may be replaced with steel sets to save on weight and the cost of transportation. It is now widely believed that any savings thus realized are later lost in the shaft sinking costs and schedule, mainly due to the increased difficulty in installing blocking (which cannot be nailed), catch pits, and launders (water rings). In addition, omission of hanging rods makes the sets more difficult to hang and align. Although a number of ingenious methods have been developed for timber shafts to successfully traverse bad ground conditions (jacket sets, pony sets, squeeze blocking, etc.), timber shafts are no longer considered when bad ground or highly stressed ground is anticipated. In some countries, suitable timber has become scarce and expensive. For this and other reasons, shaft sinking by this method is now mainly confined to deepening existing timber shafts. The circular concrete shaft, sunk vertical, is invariably employed for large shafts and most often employed for any deep shafts. The circular shafts may be as little as 12 feet in diameter for shallow applications, but deep shafts are better tooled for shaft sinking if they are of larger diameter. At hard rock mines, only a few shafts of appreciable depth have been sunk larger than 26 feet in diameter outside of South Africa. 1.8 Lateral Development and Ramps For underground hard rock miners, the term “lateral development” means the horizontal headings in a mine, such as the drifts and cross cuts at a mine level. Lateral development includes the inclined headings (ramps and declines) between levels. Because it constitutes by far the major portion of mine development, lateral development is of significant consequence. Pre-production mine development concerns are well recognized by the mining community (and discussed in other chapters of this module). Ongoing development during operations is not given similar attention. Part of the reason is that the major portion of the development costs in an operating mine are capitalized and not accounted for in mining costs statements. As a result, the significance of ongoing lateral development is partially disguised. Despite significant efforts directed at developing new equipment, techniques, and procedures, the productivity and advance rates of lateral development have shown no significant gain during the past twenty-five years. Part of the reason is that most research and development efforts have been directed at modifying mechanized equipment, such as the tunnel boring machine (TBM), continuous miner (gathering arm mucking unit), and the road header. Such machinery is capable



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of achieving acceptable advance rates in hard rock; however, other problems remain unresolved. These include dealing with high silica dust counts, poor visibility, low cutter wear life, squeezing ground, highly stressed ground, difficult equipment access, poor equipment mobility, difficulty in mechanizing ground support, and inflexibility with respect to gradient and curvature. As a result, the traditional drill and blast method remains the least expensive and most practical means of advancing lateral headings. For this reason, the next part of this module is primarily devoted lateral headings driven by drill and blast. No universally accepted standard definitions exist for terms that refer to inclined lateral headings. In this chapter, a “ramp” is a heading containing horizontal curves used as a transport corridor for rubber-tired mobile equipment. A “decline” is a straight heading suitable for installation of a belt conveyor that may also permit travel of mobile equipment. 1.9 Design and Function of Lateral Headings The design starts with determining the cross-section of the drift, cross cut, ramp, or decline. Lateral headings are contoured to the minimum dimensions required to safely permit passage of the largest vehicle while providing space for roadway dressing (or rail and ballast), ditches, utility lines and ventilation duct. Safe clearance must be provided for pedestrian traffic, especially if no safety bays are to be cut. The minimum clearances and the spacing of safety bays are usually specified in the applicable statutory mine regulations. If the heading is to become a main airway, its cross section may have to be enlarged for this purpose. In the recent past, the size of typical lateral headings has continued to increase because larger and larger haulage vehicles are employed. The philosophy has been to reduce costs by economy of scale. Larger headings are not advanced as rapidly as smaller headings, which has the effect of slowing the pre-production development schedule for a new mine and ongoing development at an existing operation. Employing remote operation and guidance systems that enable one operator to run two or three units of equipment simultaneously allow greater productivity improvement than further increasing equipment size. In this regard, a unique system employed at the Savage Zinc mine in Tennessee may have good application elsewhere (the mine employs a trackless “road train” with a 75-ton payload that travels underground at speeds of 25 miles per hour in a relatively small heading at a gradient of 6%). High-speed haulage is also employed at several underground operations in South America using a “throw-away” haul truck. For this purpose, the mines have dispensed with the typical slow-moving articulated underground truck and replaced it with one designed for highway travel. The truck is a stock model modified with a supercharger and oversize inter-cooler. After a useful service life of only two to three years on mine grades of 10%, a truck is sold to the after market for light duty and replaced with a new one. Some of these same mines employ surface-type front-end loaders underground instead of LHD units to muck from draw points and load the haul trucks with apparent great success.



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The Australians have employed relatively high-speed truck haulage for many years using offroad haulage trucks modified for underground service at typical gradients of 9:1 or 12%. (Refer to Chapter 4 for more information on truck haulage in Australia.) Track headings are normally fully arched with a vent duct hung at the crown when driven. For trackless headings, it is common practice to hang the vent duct and utility lines on the ditch side to save space and help protect them from wayward vehicles. When a very large vent duct is required, two ducts are used instead. These are hung from the back on either side of the heading, leaving the central portion open to permit passage of the heaped load on a haul truck. The back of these headings may be gently arched or driven flat backed, depending on the ground conditions (mine company standards may dictate an arched back). 1.10 Track versus Trackless A “track” mine refers to one that has rail installed in its lateral headings to provide travel for trains drawn by battery-operated, trolley, or diesel locomotives. A “trackless” or “mechanized mine” refers to the use of rubber tired mobile equipment to advance the lateral development and haul the ore. The basic component of an operation is the LHD unit. Of course, some mines employ a combination of track and trackless headings and many employ conveyors in drifts and declines for ore handling. The trend for some time has been away from track development. Even the smallest of mines are now considered best served by trackless methods, mainly due to flexibility. At the larger mines, rail haulage has been largely displaced by conveyor transport fed from an underground crusher placed near the orebody so that it can be gravity fed by trackless equipment enjoying a short haul distance. Trackless headings have other significant considerations besides flexibility. Both the productivity (i.e. feet per man-shift) and rate of advance (i.e. feet per month) are normally significantly higher for trackless headings than for rail headings. Following are the principal disadvantages of trackless headings.  The need more ground support because trackless headings are larger in cross section  The equipment that drives trackless headings requires more ventilation  The roadway is more difficult to maintain Employing electric powered LHD units and trolley-line truck haulage reduces ventilation requirements; however, flexibility is impaired. 1.11 Shaft sinking 1.11.1 Applications of Shafts Shafts are usually used for the following purposes: “  Mining mineral deposits  Temporary storage and treatment of sewage  Bridge and other deep foundations Underground Mining| 12



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



Hydraulic lift pits  Wells  In conjunction with a tunnelling system or network, for the purpose of lifts, escalators, stair and ladder ways, ventilation, conveyance of liquid, carrying pipes and cable in river crossing, drainage and pumping, particularly from sub aqueous tunnels.” They also can be temporary or permanent, (R. Tatia 2005) has classified the techniques which are used for sinking shafts: 1.11.2 Activities required for Shaft Sinking We can divide the operations for sinking a shaft into three parts: 1. Reaching up to the rock head 2. Sinking through the rock 3. Sinking through the abnormal difficult ground, if any, using special methods A sinking cycle includes the following operations:  Drilling  Blasting  Mucking and hoisting  Support or shaft lining  Auxiliary operations o Dewatering o Ventilation o Lightning or illumination o Shaft centering Drilling We use sinkers to drill holes of 32–38 mm diameter, The length of the holes vary between 1.5 to 5 meters. There are three types of cuts  Wedge cut  Step cut  Pyramid cut 1 and 2 are common drillings that are used and in rectangular shafts. Wedge cut is used most of the time. Pyramid cut is often used in the circular ones. Step cut is adopted if water is high and the shaft is of a large cross section. Blasting In practice, at the bottom of shaft is usually full of water during sinking. therefore, high density, water-resistant explosives are used. Lashing and mucking Lashing is made for the loading of muck into a conveyance for its disposal. This activity is a time consuming activity due to Presence of water, limited space.



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Support or shaft lining There are two types of lining,  Temporary  Permanent The type of water and strength of the rock and soil layer where sinking operation is done determine which option to select. Therefore, in some cases, temporary support is not adopted, while in others it becomes essential to protect the crew and equipment from any side fall. The permanent lining can be made of bricks, concrete blocks, monolithic concrete, shotcrete and cast iron tubing. Auxiliary operations  Dewatering: When the shaft is reached to the water table or beyond it, water inflows inside it, to remove this water usually face or sinking pumps are used.(tatiana) Removing water can also be done by driving deep wells or well point systems around the shaft, that results in lowering the water table around the shaft.(zhou)  Ventilation: Fresh air, supplied by a forcing fan installed at the surface, which can be provided by rigid ventilation ducts for below 6 m depth or flexible ones for more than 6 m depth.  Illumination: A pneumatically operated light, is used to provide illumination at the working face during construction work.  Shaft centering: Using the reference points, which are fixed before, to fix the shaft center. The shaft center is checked from time to time by the use of centering device installed at the surface. 1.11.3 Special methods for shaft sinking. In the process of shaft sinking, it becomes necessary to adopt a special method if the ground through which the shaft is sunk is loose or unstable such as in sand, mud, gravel, or alluvium, or when an excessive amount of water is encountered, which cannot be dealt with by sinking pumps. In some situations, both sets of these conditions my be encountered. Listed below are special methods that can be used to deal with the situations outlined above: 1. Piling system 2. Cementation 3. Freezing method 4. Grouting 5. Shotcrete 1. Piling system (Soldier pile): These piles are driven and after installing the steel beams can be concreted. Piling method and the spaces between piles depend on the soil conditions 3. Freezing method ”Sometimes when we can’t control the groundwater by pumping, we may use freezing or grouting. This procedure consists of sinking pipes around the area to be excavated and



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circulating a cold brine solution through the pipes, thereby freezing a wall of soil, this process needs 2 months to complete,” this method will be discussed in more detail later in this module. 4. Grouting: In this method we drill rows of grout holes around the shaft perimeter, then inject grout into them, but freezing is more reliable comparing to this, 5. Shotcrete: Shotcrete is sprayed concrete can be applied immediately to freshly excavated rock Impact on surrounding infrastructures Shaft sinking can have the following impacts on the infrastructure and environment around it:  blocking the streets and causing traffic in the area around it  making noise and dust which can bother the people around the construction area  bad effects on soil because of making vibrations in the ground while construction  cutting some trees and clearing the area for construction site Appropriate soil conditions The appropriate soil condition For each method is mentioned during the construction method, and if the soil is not strong, we should use piles and temporary linings to take care of that The space between columns depends on the soil conditions and amount of ground water existing, piles can be close to each other or have the appropriate distance. However, strong and consolidated soil is the most appropriate soil for driving shafts. 1.12 Site layout requirements Tommelein (1989) defines Construction site layout and its benefits as below: “ identifying the facilities that are temporary needed to support construction operation on a project but that do not form apart of the furnished structure: determining the size and shape of these facilities; positioning them within the boundaries of the available on-site or remote areas” “the so called temporary facilities usually remain on site for a period ranging from a few days to several months or even years, a time period that ranges from duration of a construction activity to the duration of a major phase of the entire construction period” (Fangyi Zhou, 2006) “ a well-organized site facilities inventory control, cuts travel times, reduces noise and dust, prevents obstructions and interference, increases safety and security, and improves site access” (Fangyi Zhou, 2006) According to Fangyi Zhou, considerations affecting the site layout are shown below: Efficiently using site space to accommodate resources throughout a construction project is fundamenta to success of project. So optimizing the construction site layout using models such as physical and computational is the interest of many researchers. (Fangyi Zhou,2006)



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Site Layout has a great effect on project costs, therefore, models are used to simulate the different site layouts and choose the best one. (Fangyi Zhou,2006) Resources required: (cost estimation for shaft, City of Edmonton,2005) Human resources Equipment operators, labourers and workers, foremen, supervisors Equipments and machines  Drill rig, compressor, excavators, explosives  Cranes, hoists, trucks  Welding truck  Lumbers (laggings), liners, ribs, tie rods, support beams, tie wire  General purpose concrete, concrete forms, concrete pump, rebars  Water pumps  Illumination and electrical equipment  Communication systems  Personal protective equipment  Ventilation System  Instrumentation to determine the concentration of flammable gases



1.13 Collars & Portals On the surface of an underground mine, a collar is required for a shaft or raise entry, while a portal refers to the entrance for an adit, decline, or ramp. Collar Besides providing a mine entrance, a shaft collar for a production shaft performs the following functions.  Keeps the shaft watertight.  Provides a top anchor for the shaft sets and the plumb lines required for shaft surveying.  Provides space for the shaft sinker to install equipment before the main excavation process begins.  May support a portion of the headframe. Collars are also required for ventilation shafts, service shafts, and for all raises that reach surface. Constructing collars in a rock outcrop or in shallow overburden is relatively straightforward; however, if the soil overburden is deep and especially if it is water bearing, collar construction can become a major project. The same is true of a portal, but in this case, if the overburden is deep and water bearing, the construction may be more difficult or even impractical. Shaft and raise collars are normally lined with concrete.



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Portals Portals may be left open to the elements in tropical zones; however, the entrance is normally enclosed with a weather-tight structure in temperate or arctic climates. This structure was once built with timber or reinforced concrete, but now miners usually employ corrugated metal archways similar to those used for large highway culverts. For a ramp entry in overburden, this structure can become very long. In the case where the portal carries a conveyor, the archway is designed with enough strength to accommodate hangers that suspend the conveyor support frames. Portals for ramps and declines usually incorporate a reverse slope at the start to prevent surface water from running into the mine. 1.14 Ground Freezing Ground freezing is considered the most reliable means to support a collar excavation in deep overburden. The method may be used for ramp entries; however, ground freezing is unusual for deep entries due to the large number of pipes required and because of the difficulty in arranging the piping to obtain a freeze in the overburden above and beneath the proposed excavation. Ground Freezing Procedure Shaft collars have been sunk employing ground freezing for over a century and so today the procedure is well understood and straightforward. Normal practice is to engage a contractor that specializes in ground freezing. A number of vertical freeze holes will be drilled around the perimeter of the proposed excavation to form the “freeze circle.” The spacing between the freeze holes varies from 2½ feet (0.8m) for shallow excavations with small freeze pipes and to 6½ feet (2m) for very deep excavations with larger freeze pipes. Each freeze column extends into the bedrock. Two pipes are installed in each freeze hole, one inside the other. The larger pipe is sealed at the bottom so that chilled brine directed in a continuous flow down the inner pipe will return to surface in the annular space between the pipes. Having taken up heat in the ground, it is then recooled in a refrigeration plant. Traditionally, the brine velocity in the annular space was designed high enough to obtain turbulent flow and assure good heat transfer. More recently, it has been determined that the transfer will not be affected if the velocity is lowered into the “mixed- flow” region; however, it should not be so slow that the flow is laminar. The brine selected is almost always calcium chloride (road salt), which in theory is capable of lowering the freeze point of the brine to a minimum of -510C (-600F) at a concentration of 29.6 % calcium chloride (SG =1.290). 1.15 Shaft Sinking-conclusion “Now when a miner finds a vena profunda, he digs a shaft collar, sets up a hoist, and builds a headframe…. Then a shaft is sunk, 10 feet long by three and one-half wide.” Georgius Agricola, 1556



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Of all the headings driven in hard rock mines, shafts are the most costly and time consuming. Moreover, the shaft sinking procedure is intricate and arduous. While a few shafts are advanced by big-hole drilling methods, the great majority employs the traditional “drill and blast” cycle to which this chapter is devoted. Shafts for smaller mines are traditionally sunk rectangular and rely on timber for support. Larger mines typically employ circular shafts lined with concrete poured in place as the sinking advances. Today, independent mining contractors sink most shafts. While there have been significant technical advances, no world records have been broken for rate of shaft-sinking advance in hard rock since 1962. Part of the problem is that mining contractors have no discretionary funds to invest in research and development, while mining companies and government agencies have other priorities for what little resources are available. Except at great depth, shafts sunk in hard rock mines do not normally require special considerations to maintain wall stability.



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MODULE TWO UNDERGROUND MINE SELECTION AND PLANNING 2.0 Mine Planning – Some Considerations This module discusses the pre-planning stage of mine development. We assume here that we have a mineral deposit of sufficient quality to justify mining. At this point, one needs to decide if a method of mining should be surface or under-ground. Usually, the depth and size of the deposit makes this decision obvious. However, that is not always the case. The factors affecting the two options are given below†: 2.1 Underground or Surface Underground methods -



-



Good for relatively deep deposits. Usually, the depth is more than 100 feet. In a lot of situations, the depth factor makes the decision easy. If the deposit is too deep inside, surface methods are ruled out. Generally cost per ton is higher than surface methods. Therefore, good for high quality grades only. Less disruptive environmentally. In the past, however, reckless underground mining left behind large tracts of subsided land. Mining is rarely affected by the climate. However, artificial ventilation and lighting is required. In some very deep mines, mine production is affected by heat (due to the depth). Definitely more hazardous than surface mines. Mining in a coal seam is affected by present/old workings in other seams. Return on capital is generally not quick. Generally less dilution when mining. Especially good for complex ore bodies where selective mining can be carried out.



†



The factors listed here cover both non-coal and coal applications. Therefore, some factors may not be applicable in coal and vice versa.



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Surface methods -



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-



-



Good for shallow deposits. The maximum depth that can be mined by this method is dictated by the technology used. With rapid advances in technology, surface mines have gone significantly deeper than before. For example, Bingham Canyon copper mine in Utah is about half a mile deep. Generally lower cost per ton than underground methods. Therefore, even relatively poor grades can be mined too. Very disruptive to the environment. According to a study in 1978, 75% of the land affected by surface mining is due to mining of coal, gravel and crushed stone. Reclamation can be expensive. Sometimes, companies prefer to go underground (despite bad economics) simply to alleviate environmental concerns. De Beers only considered underground mining in their Snap Lake diamond project in Mackenzie Valley, Canada. Mining is affected by weather. Inclement weather can lead to mining stoppages. Less hazardous. Does not require artificial lighting during day hours. Multiple seams can be mined without being subject to ground control problems. In Yellandu, India, the Singareni Colleries Company Limited mined seams by surface methods that were previously mined by underground methods‡. High capital required for modern mines. Cannot be used for selective mining. Generally higher productivity than underground mines.



It should be mentioned here that not all factors mentioned above are hard and fast. For example, if a company has several open pit mines, and decides to open a new one, the capital required will be less if they divert some equipment from the existing mines. Usually, companies may schedule the closure of a mine with the opening of another just to reduce capital costs. Having decided to mine by a surface or u/g method, the next step is to plan and develop the mine. The surface mining methods will be presented later in the subject Surface Mining. In this module, we will discuss planning of an u/g coal mine. 2.2 Preliminary Strata Mechanics A very good practice in mine design is to look at the strength data available from the boreholes. This can help in avoiding problems in the future. Even when strength data is not available, one can look at the lithology and make a qualitative assessment. This assessment, though entirely based on experience, is very useful. An immediate use of strata characterization is in the layout of the mains and sub-mains in coal mines. Mains have a long life and therefore, their size and location must be planned carefully. They are the main arteries of material and human transport. ‡



Since the u/g workings used very little technology, only a small fraction of coal was mined and most got left behind due to ground control problems.



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The same is true for non-coal mines as well. The main entries/drifts to the mines are always in strong strata. Their location is also determined with regards to future mining activities, i.e. they are sited so that future mining activities do not deleteriously harm them. Background on Roof and Floor Characterization: Roof (or “back” in non-coal mines) and floor stability is very important for ensuring smooth and safe production. While roof falls can be fatal and disruptive to production, floor disturbances generally slow production. A good floor is essential for conveyors and most production machines. In severe cases, floor heave can essentially close an entry. An inspection of the borehole logs gives a very good idea of the roof and floor that can be expected around the seam. In coal mines, one must estimate the immediate roof from the borehole logs. Generally, the immediate roof is the strata between the top of the seam and the plane of contact of weak shales and rocks with strong strata (Figure 1). If the weak rocks immediately above the seam are thick in comparison with the coal seam, then the immediate roof depends on the expansion ratio of the rocks, as when they fall, they occupy a volume indicated by the expansion ratio. Depending on the expansion ratio, the immediate roof varies from two to five times the excavation or seam height. Immediate roof is important to know since that is the roof that needs to be supported for safe mining.



sandstone



Immediate roof



shale siltstone coal



Figure 1. The immediate roof is the distance between the roof of the seam and the floor of the first hard layer.



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The roof and floor can be characterized using criteria such as compressive strength, shear strength, tensile strength, moisture content and presence of joints/fractures. Typically, roof falls are caused by tensile stresses, at the center of the opening, and shear stresses, at the corners of the openings. Weak immediate roof, or presence of clay and moisture, or fractures/slickensides indicates greater support requirement. Soft floors, on the other hand, may be punctured by pillars. Other things to watch out for are shale/sandstone contacts. These contact regions have high stresses due to the significant difference in their Young’s modulus. We discuss some quantitative characterization methods next. Rock Quality Designation (RQD): This is a very handy classification tool. It is defined as the ratio of the cumulative length of core sticks (NX size) each greater than 100 mm long to the total length of the drill hole. This method assumes that the length of the core pieces depend on the structure and strength of rock. Sum of length of core sticks greater than 100 mm long



RQD=



X 100 Total length of drill hole



It is evident from above that computing the RQD is very easy. One must, however, be careful. For example, when the holes are parallel to bedding, high RQD is indicated, while for the same rock if the hole is perpendicular to the bedding, low RQD is indicated. Rock Mass Rating (RMR): This is one of the most popular characterization criteria. It uses six parameters in its classification: the uniaxial compressive strength, RQD, spacing of discontinuities, condition of discontinuities, groundwater conditions and the orientation of discontinuities. Depending on the conditions, the strata in question is assigned a rating for each of the first five parameters. These ratings are summed to arrive at the basic RMR. This rating is then adjusted for orientation of discontinuities. The adjustment takes the form of a penalty for harmful discontinuity orientation. The adjusted RMR is the rock’s RMR. The classification also provides a table for translating the RMR into average stand up time for tunnels, cohesion of the rock mass and the friction angle. Q system: This system was developed out of a study of 212 tunnels in Norway. The Q rating was given by the formula: RQD x Jr x Jw



Q= Jn x Ja x SRF



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where, Jr = joint roughness number, Jw = joint water reduction number, Jn = joint set number, Ja = joint alteration number, and SRF = stress reduction factor. The RQD and Jn are indicators of the overall rock structure, Jr and Ja are indicators of shear strength of the joints, Jw is a measure of water pressure, and finally, SRF is a function of i) load reduction due to shear zones and clayey rocks, ii) stress existing in rock and ii) squeezing and swelling loads in soft plastic rock. Jw and SRF are indicators of confining stress. One aspect of the Q system that stands out is that the rock strength is not directly taken into account.



An Example Characterization: Figure 2 below shows a coal property with borehole locations and their RQD’s. The method of polygons is used to obtain the areas of influence of each borehole. It is assumed here that the polygons shown in the figure accurately represent the weight of each borehole. We color code the polygons according to the scheme: Good or RQD  90 Green Moderate or 90>RQD 80 Yellow Bad or RQD < 80 Red



75



65



89 87 93



95 Mains 91



75



84



This map is very good for deciding on the location of the mains. As we would want to lay the mains inFigure areas2.ofRoof competent roof, the only forproperty. Figure 2 is to lay the mains East-West as characterization using RQDoption for a coal shown. Most of the mains will have very good roof in that case. During mining, they should expect roof control problems in the north, northwest, and south side of the property. If possible, one could also size the pillars in the different zones differently. .



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2.3 Portal/ Access Location It is very important for a mine to have a reliable access to the deposit. The type and location of the portal significantly affects mine operation. The only way for men and machines to reach the deposit is through them. Portals are the first to be developed and the last to be abandoned, and therefore, they have the longest life among other mine structures. Some of the factors that should be taken into account during portal construction are: -



-



Portal should be located in strong strata. Else, the portal could be subject to strata disturbance jeopardizing mine production. Portal area should be free from flooding. Should be as close to the center of the property as possible. The portal should be at the lowest elevation of the seam/orebody so that the grades favor hauling and mine drainage. Transportation costs are lower if coal/ore is hauled down gradient and drainage is easier when the final destination is lower than the origin. Transportation, both inside and outside the mine, is affected by portals. Conveyor belts, loading chutes and other elements of coal hauling network inside the mine are planned on the basis of portal location and size. If for some reason a portal cannot be used, arranging for alternate routes can be very expensive. The location of the entrance to the mine affects the trucking distance between the mine output and the next destination for the coal/ore (coal/ore processing plant or customer). This is especially true in hilly areas such as the Appalachia, where a portal placed on the wrong side of a hill adds several miles of additional transportation. Generally also, remote location of a portal requires building a road. A deeper/longer portal is often preferred to a shorter portal when the shorter portal is farther away from the customer or the washing plant or the highway (Figure 3). Bad location: requires going around the hill



to the highway or the customer



Longer portal Good location even though the portal is deeper (one time large expense).



portal coal seam



Figure 3. Effect of the location of the portal on transportation.



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Underground Mining



It is often found that the best conditions do not occur together. For example, a portal location that is the closest to the customer may have the worst strata conditions. Therefore, one must carefully weigh all the options before deciding on the location. 2.4 Types of Portals There are three types of portals: shafts, inclines or slopes, and adits or drifts (Figure 4).



adit



incline



shaft



Figure 4. The three types of portals.



Adits or Drifts: An adit or a drift is the access of choice when the coal seam outcrops or where the ore/coal is accessible through a horizontal drift. No special equipment besides what is needed for the mine is generally necessary to drive them. They are the cheapest to construct and have a low operating cost. However, their application is limited to those with favorable geology. Inclines or slopes vs. shafts: Slopes have certain advantages over shafts: - They can be driven quicker than shafts, and cheaper on a per foot basis (up to a certain depth). Figure 5 shows the comparative costs (unit and total) for shaft and slope. - They support continuous hauling, by use of conveyors, unlike shafts. They can also handle large volumes, making the operating costs cheaper. - Provides access to the outside of the mine in the case of emergencies. While this is true of shafts too, climbing hundreds of feet vertically is difficult. - Transporting heavy and wide/long equipment is relatively easy and may not require dismantling.



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The disadvantages of slopes are: - Even at the maximum angles for coal transport by conveyor belt, slopes are about three times in length, for the same depth, compared to shafts. - Capital cost is higher than shafts. - If the strata is poor, the increased length results in higher maintenance costs. - Increased length causes greater pressure drops affecting the ventilation.



2000



5000 Unit cost, slope



1600 Unit cost, $/vertical meter



4000



Unit cost, shaft



1200



Total cost, ’000s $



3000 Total cost, slope



800



Total cost, shaft 2000



400



1000 122



244



366



486



610



732



853



914



Total depth, meters



Figure 5. Comparative costs for shafts and slopes.



Driving an incline or a shaft is a technical endeavor. Therefore, the economics also depends on the type of equipment used. Mismatch between the strata and the equipment can prove costly.



2.5 Mining Method Selection To select a mining method, certain data describing the orebody is required.  Geological cross sections and a longitudinal section  Level maps  Block model (grade model)  Geomechanical characteristics of the host and surrounding rock. One approach is to find one or more comparable ore bodies that are being or were mined successfully and use that mining method(s) to determine the most likely methods to investigate further. Another approach is to determine applicable mining methods and develop a short list for detailed consideration through a process of rationalization. Underground Mining| 26



Underground Mining



Following are typical considerations to be weighed in selecting a mining method (listed roughly in order of importance).  Maximize safety (integrity of the mine workings as a whole or in part).  Minimize cost (bulk mining methods have lower operating costs than selective extraction).  Minimize the schedule required to achieve full production (optimize stope sequencing).  Optimize recovery (80% or greater recovery of geological reserves).  Minimize dilution (20% or less dilution of waste rock that may or may not contain economic minerals).  Minimize stope turn around (cycle) time (drill, load, blast, muck, backfill, set).  Maximize mechanization (trackless versus track and slusher mining).  Maximize automation (employment of remote controlled LHDs).  Minimize pre-production development (top down versus bottom up mining).  Minimize stope development (selective versus bulk mining methods).  Maximize gravity assist (underhand versus overhand).  Maximize natural support (partial extraction versus complete extraction).  Minimize retention time of broken ore (open stoping versus shrinkage).  Maximize flexibility and adaptability based on size, shape, and distribution of target mining areas.  Maximize flexibility and adaptability based on distribution and variability of ore grades.  Maximize flexibility and adaptability to sustain the mining rate for the mine life.  Maximize flexibility and adaptability based on access requirements.  Maximize flexibility and adaptability based on opening stability, ground support requirements, hydrology (ground water and surface runoff), and surface subsidence. Following is a list of mining methods most often employed underground listed roughly in the order of increasing cost (direct mining cost, including backfill where applicable). The order is generally true, but can be deceiving because some methods, such as Blasthole can have a wide range of costs. 2.6 Bulk Methods 1. Block Caving/Panel Caving – columns of rock are undercut wide enough to cave under the weight of the column. Caving is initiated by undercutting the ore zone. Block Caving involves a significant capital investment in pre-production development and may be especially risky. It should only be implemented in consultation with a block-caving expert. 2. Blasthole/Sublevel/VCR – the ore is drilled in rings or by long hole and the ore is drawn off (“mucked”) as it is blasted. A common variation is to pull only the swell and



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leave most of the broken ore temporarily remaining in the stope to support the walls (deferred pull). 3. Sublevel Caving – the ore is drilled in rings and drawn off (pulled) after blasting in successive lower lifts. Unless the ore dips steeper than 70 degrees, a great deal of ore may be left behind as production losses. One difficulty with Sublevel Caving concerns grade control. A gradual dilution occurs toward the end of the draw and it can be difficult to determine when it is best to stop pulling. Recovery may be improved if the draw point layout is staggered from one level to the next. 4. Room and Pillar/Post Pillar – a grid of rooms is developed on a near horizontal plane, leaving pillars of ore to support the back (roof). The pillars left in the Post Pillar method are undersized (posts) and designed to fail in a controlled manner. Typically, a zone of low-grade mineralization or host rock (“barren”) must be mined with pay grade ore to maintain access and control stress distribution. On the other hand, Post Pillar (and even Room and Pillar) may be considered to be selective when the pillars can be arranged in zones of lower grade material, as opposed to a regular geometric pattern. 5. Modified Avoca – the ore is drilled by long hole and drawn off in retreating vertical slices, followed closely by placement of rock fill dropped “over the bench” or “over the fill” (via access to the back of the stope from the footwall drift). 2.7 Selective Methods 1. Shrinkage (narrow vein) – the ore is sliced off in successive horizontal lifts (overhand). Only the swell is drawn off leaving broken ore to support the walls and provide a working platform for the next lift. Narrow vein shrinkage stoping is classed as selective because it permits mining to variations in the horizontal contour of the vein and even removes pockets of ore extending into the wall rock. It is not selective with respect to the fact that once initiated, a shrinkage stope has to “take it all.” Normally, a barren portion of the vein cannot be left behind. 2. Resuing (narrow vein) – a method that reduces dilution when the vein is narrower than the heading. Historically, a drift round was taken in two passes. First, the waste rock was drilled blasted and mucked out, then the narrow vein was slashed to recover the ore, undiluted. In most cases, waste rock quantities can be significantly reduced by innovative measures. Single pass resuing uses appropriate delay blasting to throw the ore and waste rock into separate piles. 3. Cut and Fill (Overhand/Underhand) – access is provided by first ramping down from a cross-cut access and then taking down back in successive slices. After mucking, the stope is filled but enough space is left to mine the next slice. Mining equipment is captive unless an access ramp is employed. If underhand (undercut), slices are taken from the top down under cemented backfill or a concrete matte.



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4. Drift and Fill (Overhand/Underhand) – a modification of Cut and Fill in which drift sized cuts are taken adjacent to one another and, upon completion, packed with cemented backfill. The process is repeated next to the backfill once it has consolidated. Many additional mining methods exist; the foregoing are the most commonly employed. The selection of a mining method and its application to a new orebody may be simple in some cases, but it is more often a challenge requiring not only logical and practical reasoning, but creative minds working in three dimensions. 2.8 Dilution Modern bulk mining methods reduce direct operating costs and facilitate management of the mine operations, but a common drawback is often increased dilution. For ore bodies of vast expanse, dilution is not a problem; however, most mines deal with ore zones of finite width and many experience dilution as high as 20%, or even 25% when bulk-mining methods are employed. Dilution is the great nemesis for miners because the cost of dilution is not only the obvious direct cost (dilution tons displace ore tons in the ore handling and process circuits), but also includes significant indirect costs. For example, each ton of sterile rock or backfill that circulates through the mill carries mineral values with it to the tailings. The minimization of dilution should be given weight in the selection and subsequent application of a mining method. The causes of excess dilution include using the wrong mining method and related factors. The causes can be illustrated in the following fish-bone chart (Figure 3-1).



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Underground Mining



2.9 Mine Planning Once a mine is producing, a series of mine production schedules based on current reserves and proposed mining methods should be prepared and updated on a regular basis. These schedules form an integral part of the ongoing reconciliation that must be performed to monitor the success of the selected mining method. Most active mining operations in the world have the following production schedules readily available for review and audit by senior staff or consultants. Three month (updated every three months) Six month (updated every six months) One year (updated every year) Two year (updated every year) Five year (updated every two to three years) Ten year (updated every five years) 2.10 Strategy for Underground Mines 2.10.1 Ramp Haulage



For small ore bodies, ramp haulage is the default selection because it normally provides the most flexible and economical choice. (In a cordillera, the terrain may provide relief adequate for a



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level entry or “adit.”) A ramp (or adit) drive can typically be oriented to provide an underground diamond-drilling base and provide shorter crosscuts to the ore zone. The crosscuts are provided rapidly and economically because they provide a second heading for the main drive. It is possible to sink and develop from a shaft at the same time; however, this is a difficult and expensive procedure. Another advantage to the ramp or adit entry is direct access by mobile equipment when trackless mining is to be employed. For a typical shaft, the equipment must be dismantled and reassembled underground. The set-up time required to initiate ramp driving is usually shorter than for a shaft. One to three months may be required to provide access and collar a ramp portal, while the collar, hoist, and headframe required for a shaft may take six months of site work. For medium sized ore bodies, ramp haulage may still be the best choice where the orebody is relatively flat lying. In this case, the ramp may have to be enlarged to accommodate larger trucks. In some cases, it may be practical to provide twin ramp entries to handle two-way traffic. 2.10.2 Belt Conveyor



For large, flat-lying ore bodies, a belt conveyor is typically the most economical method of hoisting ore. The legs of the conveyor are put into a ramp that has been driven straight (i.e. a “decline”) for each leg of the proposed conveyor way. If the soil overburden is very deep, or deep and water bearing, a ramp or decline may not be a practical method due to the extraordinary cost of excavating and constructing a portal. If the ground (bedrock) beneath the overburden is not competent or is heavily water bearing, a ramp or decline access may be impractical due to the driving time and cost. 2.10.3 Shaft System



For large steeply dipping ore bodies, a shaft system is usually best. In this scenario, it may be advisable to have a ramp entry as well to accelerate the pre-production schedule and later to provide service access to the mine. 2.10.4 Conventional Methods of Ore Transport



At the conceptual stage, it is normally better to consider only conventional methods for the transport of ore and resort to the unusual methods only under unusual circumstances. A good example of “unusual” is the aerial tramway installed across the Black Angel Mine in Greenland to access an orebody located high on a cliff face. Table 4-1 lists methods employed at underground mines for ore transport.



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2.10.5 Main Entry



The foregoing strategy determines a main entry to the underground on the basis of ore transport. In many cases, this entry also serves for personnel and materials transfer, particularly at small operations. Consideration should be given to a separate entry for man and materials handling when it can be afforded. For example, some mines use the production shaft for ore/waste hoisting, main exhaust, and alternate escape while a second shaft provides cage service in the main fresh air entry. If a shaft system is employed at an operation of substantial capacity, it is not uncommon to find a ramp access from surface as a third entry. This is a logical progression when an internal ramp system is required by the mining method to be employed.



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MODULE THREE SELF-SUPPORTED MINING METHODS 3.0 Room and Pillar Method The next stage in mine development is to actually open the mine. Typically, coal is mined by one of two methods, room-and-pillar (or bord and pillar, in British terminology) mining and longwall mining. The method of mining is chosen based on several factors (to be discussed in the next few modules). We begin with room-and-pillar mining. The first part of this handout presents the nomenclature. Room-and-pillar mining involves opening a series of tunnels in parallel (called entries) with tunnels going across them (called crosscuts) at regular intervals leaving solids blocks of coal in between (Figure 1). The crosscuts can be perpendicular to the entries or be at an angle. The tunnels themselves (be it the entries or the crosscuts) are called rooms (locally, i.e. in a give spot) while the solid blocks of coal are called pillars. Pillars are left to support the rooms. During mining, these entries are extended all the way up to the end of the deposit after which, the pillars are removed in retreat. Removal of the pillars (called pillaring) from an area concludes mining in the area (the roof collapses very quickly after a pillar is removed, thereby reducing the stress).



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Underground Mining



Entries are generally named from left to right, facing inby. Inby and outby are terms frequently used in mining to describe location relative to the observer. Inby is towards the direction of advance, while outby is in the opposite direction (generally towards the exit). In Figure 1, A is inby of B, while B is outby of A. Mine Plan Format: If you are seeing a room-and-pillar mine plan for the first time, you will notice that in the mine plan given to you, rooms are depicted as lines, unlike in Figure 1. To help you make the transition to mine plans, Figure 1 has been redrawn in a typical mine plan format in Figure 2 (not in the same scale, however).



Before opening a mine, few parameters need to be decided on. These are presented next.



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Underground Mining



Number of entries: The example mine plan that was given to you uses a 7 entry system (ask me if you do not have the mine plan). This is evident from the 7 entries for mains, sub-mains and panels. Mains, as the name suggests, are an important part of the mine. They form the skeleton or backbone for the mine, allowing access to various parts of the mine. Every ton of mined coal passes through the mains. Due to their long life and pivotal role in coal hauling, mains usually have pillars larger in size than in other areas of the mine. Sometimes, the same is true for submains also. Sub-mains are the system of entries that branch off mains, giving access to various portions of the mine. Sub-mains may be absent in small mines. Panels are entries that branch off the sub-mains. Mining is generally done off the panels. However, mine geometry may sometimes necessitate driving panels off mains. Also, the distinction between mains and sub-mains, and sub-mains and panels may not always be clear. The decision to drive a certain number of entries in a panel should consider the following: - The purpose of the panel. For longwall development, typically 3 or 4 entries suffice. It is difficult driving anything lower than three entries since belts, air-return and track/human transportation each (typically) require a separate entry. Additionally, congestion can be an issue for fewer entries; for example, even with three entries, there is very little room for equipment parking (for repairs or routine daily maintenance). Figure 3 shows a 4 entry longwall development section. Note how congested it is. - When the number of entries are small, there are only a few available working faces. This usually results in delays. This is because mining is cyclical (including continuous mining). Cutting is followed by bolting and servicing (clearing the floor, advancing ventilation and rock dusting), and therefore, any delay in any of the activities affects the others. When there are many working faces available, an activity can be shifted to another location without any interruption. For example, if the bolter is down in entry#1, the miner could work in entry #2 rather than wait for the bolter to be repaired. - More entries mean more space. Therefore, machines can maneuver easily and quickly, saving time. In case of shuttle car sections, more entries may mean different routes for loaded and empty cars, thereby reducing travel times.



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Underground Mining



- The more the entries, less is the air resistance. Therefore, volume of air flow is higher than with fewer entries. However, with too many entries, section ventilation may be difficult and prone to leakages (too many curtains). Underground Mining| 36



Underground Mining



- One disadvantage of many entries is that it slows down development and therefore, pillaring. Hence, cash flow may be small for a long time. - Too many entries can also be cumbersome and uneconomical. Section ventilation and management becomes difficult in a very spread-out section. The number of entries in a panel is usually 3 to 7; with 3-4 entry systems being primarily for longwall development. In the mine plan given to you, it was decided to use bridge sections and, therefore, the mine was planned with 5 entries (since any less number of entries causes congestion with bridges). Many times, when the number of entries is large, two sets of equipments (i.e. two sets of miner-bolter-car combinations) are used. These sections are termed super-sections. Angle between entries and crosscuts: The angle between crosscuts and entries depends on the machinery. Bridges and ramcars, for example, require oblique angles, while shuttle cars require perpendicular angles. The angle between the panel entries and main/sub-main entries in the mine plan given to you is oblique to reduce spillage in conveyor transfer points. Conveyor spillage is high when the direction of flow takes a sharp turn (Figure 4). Therefore, one must plan oblique angles wherever the conveyor flow direction is expected to change.



Entry/crosscut height and width: The mining height is dependent on geology (seam heights) and panel requirements. Typically, as long as it can be economically justified, a height comfortable for humans, is preferred. However, many times heights are also justified based on current equipment and personnel. For example, I have known mines to drive certain entries higher to accommodate longwall shields. Sometimes, entry height is determined by roof control; for example, a thin layer of slate is better mined down instead of being supported. The width is primarily dependent on equipment needs. Most entries are 16-20 feet wide. Pillar sizes: These are determined by ground control and systems engineering. The ground control aspects are discussed at length in the next handout. As far as the systems aspect go, as pillars become longer, travel times from one crosscut to another becomes longer. It impacts the haulage cycle times (shuttle or ram car) since when one car is in an entry, others have to wait until the entry is cleared. For larger pillars, such idle waiting times can be long. While driving longwall panels, there is always a conflict between driving longer pillars (reduces total amount of cutting) and reducing idle waiting times by reducing pillar lengths. Mining Progression: Figures 5 through 8 show the progression in mining activity during the life of a mine. This discussion assumes a pure room and pillar mine (i.e. no longwall panels). The first stage (Figure 5), development, involves driving the mains all the way to the end of the Underground Mining| 37



Underground Mining



property. The second stage may involve further development, especially in large deposits. In Figure 6, panels are being developed on the flank, while small sections are being developed off of the mains. In the next stage (Figure 7), the sections off of the main are extracted or pillared, while panels continue to being developed. At this point, the production levels in the mine are very high due to the presence of pillaring sections. Mines like to have few pillaring sections as they are very productive. Most mines remain at this stage for a long time, i.e. most mines have both pillaring sections and development sections for a majority of their lifetime. Mining progresses this way (Figure 8) till the deposit is completely mined.



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Development around a panel i. Extraction in some areas (dashed lines indicate pillaring or extraction) ii. ii. Development in others



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Underground Mining



Caveats: Many times, however, due to financial reasons, a mine may start developing panels as soon a possible, without waiting to drive to the end of the property. This is especially true of longwall panels, that take about a year to develop. In Figure 9, advance of the mains is delayed, while the longwall panel is developed. The main reason for doing this would be to start the revenue stream quicker.



A possible problem with the above is that in the future when panels are developed farther from the mouth of the mine, travel to these panels require that one pass by the older sealed off panels (Figure 10). While sealed off panels are generally safe, they can prove hazardous as seen in the Sago mine accident in 2006.



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Underground Mining



Production Cycle: There are two types of production methods in room-and-pillar mining, continuous mining and cyclical mining. Cyclical mining consists of drilling, blasting, mucking or loading and supporting. As these actions occur in sequence or cycles, this type of production is called cyclical. As drilling and blasting is becoming rare in coal mines, and due to time constraints, only continuous mining will be discussed in this course. Continuous mining assumes non-cyclical production. This is attained by the use of a continuous miner that cuts the coal while simultaneously loading it into a conveyor or a shuttle car, thereby, eliminating the need for drilling, blasting and mucking cycles. However, in reality the term continuous mining is a misnomer. Given the continuous miner, and assuming a good haulage system, one can theoretically mine continuously. However, mining is not continuous. This is because mining laws require ground control and ventilation measures to be undertaken after each cut. For example, no opening can be unsupported for more than 40 feet. This means that any time the depth of the cut reaches 40 feet, one must stop cutting and allow roof supports to be installed. Therefore, cutting is only one aspect of a cycle of activities. The cycle usual consists of: - cutting coal - install roof supports, usually roof bolts - extend ventilation using ventilation devices such as brattices or tubes - service the face (scoop it) including rock dusting - pump water, if necessary - survey, to ensure advance in the proper direction - periodically one must re-locate the power center so that the machines can advance - also periodically, one must advance the static conveyor. The last two need not interfere with the production cycle when a mine only has two shifts of production per day and one shift of maintenance. These can then be accomplished during the maintenance shift. Any backlogs in roof support is generally also accomplished in the Underground Mining| 42



Underground Mining



maintenance shift. Surveying is not necessary after each cut though foremen measure the advance from each cut using a tape. Note that miner-bolters are equipment that allow bolting, while cutting coal, thereby not requiring that mining be stopped for roof bolting. Additionally, these machines also have a builtin rock duster. Therefore, when using miner-bolters, coal mining can be very close to continuous mining. Equipment: The typical equipment requirements for room-and-pillar mining are (per section): - continuous miner - roof bolter - haulage machinery such as shuttle cars and bridges - utility equipment – a scoop and a LHD - conveyors The various types of continuous miners, roof bolters, haulage machinery such as shuttle cars and bridges, and conveyors are described in the textbook. The capacities of the machines, however, are a little outdated.



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Note: In the above example, we did not take into account the tram time between cuts or the time to service the face. We are assuming that the miner and the bolter have the same tram time and that there is an additional face that the scoop services, while the bolter is bolting the fresh cut. In a four entry system, typically only one free face is available and, therefore, bolting and servicing have to be completed, while the miner is cutting the face. The ventilation also determines cycle time. A four entry longwall development section is presented next, where it becomes clear that miner-shuttle car-bolter times are not the only factors that determine cycle time. Figure 11 presents a 4 entry longwall development section. As in most modern mines, the section has two miners (one for each side, left and right) and two bolters (one for each side, left and right) in addition to four shuttle/ram cars. Note the section ventilation. In the left side, the intake air travels from #2 entry to #1 entry before returning, while in the right side, it travels from #3 entry to #4 entry before returning. Since miners cannot work downstream of a continuous miner (due to high dust levels), it precludes #4 entry being bolted or serviced, while #3 entry is being mined. Similarly, when #2 is being mined, #1 entry has to be idled (as shown in Figure 11, where the bolter is simply parked). When the miner is in #1 entry, #2 can be bolted and serviced (except, when #2 is being rock dusted, the miner in #1 would have to be briefly stopped). Same applies to #3 and 4 entries.



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Therefore, as far as equipment utilization goes, only the miner is being utilized on one side when the inside entries (#2 or 3) are being mined. The bolter is effectively idled since space limitations require it to be parked in the outside entries (#1 or #4 entry), which gets dusted out. The scoop can hopefully be kept active doing other work. Lets look at a longer cycle to see how the actual run might be. Considering only the right side (the analysis applies to right side since it’s a mirror image), when the miner is done cutting #3, it moves to #4. While it is cutting #4, #3 gets bolted and serviced. If everything works perfectly, #3 will be ready when the miner gets done with #4. But so what? Can the miner go to #3? No, because #4 now needs to be bolted and serviced, which means, you cannot dust out #4. Therefore, the miner will have to wait, so that #4 can be bolter and serviced. Hopefully, during this idle time, some maintenance is done on the miner (oil and bits changed, general cleaning done besides other routine maintenance).



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In any case, these issues are not taken into account in computations done previously. Therefore, one must take into account the entire cycle when estimating section productivity. In my experience, I have seen management very often over-estimate section advance rates since idle times such as from moving the return up, or those due to belt and power moves, or inbuilt cycle delay time (like the one presented earlier) are not incorporated into time studies.



3.1 Non-coal Application of Room and Pillar Mining (Stope and Pillar Mining) Application - Large, flat or nearly flat deposits (less than 30 degree dip) • has been used for small bodies too • in very thick deposits, pillar degradation is possible, requiring larger pillars or pillar reinforcement • problems with very thin seams too. - competent roof and floor • however, this could be applied to various kinds of orebodies (host rock strength between 344.7 – 27.5 Mpa and at depths from 15-915 m). Essentially, that range covers any rock that will withstand development w/o massive supports - examples of room-and-pillar mines are in lead (Buick mine, Missouri), zinc, salt, trona (Wyoming), sandstone mining, limestone Three variants - horizontal stoping of a flat deposit - stoping of an inclined deposit (between 20-30 degrees) - horizontal stoping or step mining of an inclined deposit. Development - minimum development in flat deposits - roadways for access and transport of ore - development and mining can be combined Production ♦ thick deposits - mined in slices. - production starts at the top and roof is bolted. Sometimes, however, some ore may be left at the top due to • the ore may not have been known to be there. In such a case, “overhand stoping” (see below) can be used to mine the upper slice.



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• the ore may be uneconomical at the present time - after a suitable advance, the bottom is excavated by benching. This method is called “underhand stoping”. • The vertical face on the top layer is called “breast” • Since benches are developed in a downward direction, this method is also called “underhand stoping”. • Another variant is “overhand stoping,” where horizontal slices are taken in an upward direction. Sometimes, such as in irregular deposits, it becomes necessary to leave some ore on the floor to reach the full height of the stope. • In underhand stoping, the point of contact between the top floor and the vertical face is called the “bluff line,” whereas in overhand stoping, the contact between the lower roof and the vertical face is called the “brow line”. • In underhand stoping, drilling benches with vertical holes has the advantages of i) larger tonnages per blast (this in turn allows better scheduling of loaders, trucks, jumbos etc.), ii) better pillars due to pre-splitting, iii) lower capital cost of drilling equipment and iv) the possibility of presplitting being done long before the actual blast. When blasting benches with horizontal holes, the advantages are i) better control of mined grade (due to larger number of benches as each bench is now smaller), ii) better breakage of sedimentary strata resulting in lower powder factor, iii) ability to re-use face drilling equipment for bench drilling when the face is close to the bench thereby reducing inventory and personnel, iv) more even floors and v) no need to clean up flyrock or be concerned with back break at the top of the bench. - ramps necessary for transport vehicles to change levels ♦ inclined deposits - several openings made at different levels of the deposit - production starts at the bottom level and moves up, along the dip, to the next level - uneven levels make mechanization difficult. - slushers, in combination with rail cars, used for the transport of ore ♦ inclined deposits mined in steps - mining starts as horizontal openings that branch off the access drifts - similar other openings are made, one level below and parallel to the previous one - ramps connect these parallel openings - trackless haulage - equipment same as for flat deposits ♦ Face areas can be large (from 30m2 in metal to 180m2 in limestone). Compare this to face sizes in coal, where face width is usually 6.5 m and height depends on the seam height and is typically 1-4m resulting in a face area of 6.5-26m2. Face height in non-coal applications can be as high as 10m, even though in most cases thickness greater than 6m are taken in slices.



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- Presplit holes are a good idea if the pillar heights are expected to be greater than the width. Pillar recovery: Rare, because - pillars are small. Therefore, removing them is hazardous - irregular pillars and therefore, systematic removal not possible - first mining aims at maximum recovery eliminating the need for second mining. Extraction rates up to 85% are possible. Ore handling - diesel front end loaders and LHD’s generally. For inclined w/o steps: slushers and track mounted haulage - when large spaces available conventional dump trucks (up to 100 ton capacity) can be used - belts used on soft ore bodies - all equipment should be mobile due to the large horizontal extent of the ore body - jumbo drills, air leg mounted drills Advantages - very flexible (adaptable to changes) • can be selective • flexible around varying thickness - can be mechanized • high productivity 27-70 tons/shift - can be applied to several level simultaneously w/o affecting other levels structurally - multiple faces possible - sometimes mineral production occurs in development also - mobility of equipment is a big advantage (maintenance is easy) - easy to install good ventilation Disadvantages: - roof needs to be supported over a longer time • especially problematic in bad strata • for high backs, expensive equipment may be required to inspect and maintain - capital intensive due to mechanization



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3.2 Types of Room and Pillar Mining



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3.3 Shrinkage Stoping This method is very similar to sublevel stoping. Here, the caved ore is not loaded in its entirety. Instead, it is used as a base for further mining activity. Shrinkage stoping is a flexible mining method for narrow orebodies that need no backfill during stoping. Successive horizontal slices of ore, usually about 3 metres (10 feet) high, are taken along the length of a stope, in a manner similar to cut-and-fill. The ore is removed from the stope through drawpoints at the bottom horizon spaced about every 7.5 meters (25 feet) along strike. Just enough ore is left in the place to provide a floor from which to work when taking the next cut. This requires considerable planning and coordination. When the ore is blasted, it fills a space about 1.5 times the size of the space it filled as a solid mass. This is called swell and is an important factor in determining how much ore to draw from the bottom of the stope in order to maintain adequate working room. The broken ore is drawn down from chutes below, thus “shrinking” the volume of broken ore in the stope. The process is continued upward until the stope either reaches the next level or is stopped at some predetermined elevation. Horizontal crown pillars are left behind at the top of the stope. Shrinkage stoping depends on gravity to keep the broken ore moving to the draw points, so it works only in steeply-dipping orebodies. There is no provision for support, so the wall rocks must be strong and competent. The orebody must also be wide enough to allow a working width all the way up the stope, generally no fewer than two meters. Application - steep orebody (dip greater than angle of repose) • the best dip is in the 70-90 degree range. Ore support of the hanging wall is less for less steep dips. At a dip equal to the angle of repose, support for the hanging wall is close to zero. - firm and competent orebody - minimum thickness of about 1 m and maximum anywhere from 3-30 m - stable hanging wall and footwall as broken ore doesn’t give much support - regular boundaries - must not be affected by storage in the stope. Must not emit toxic gases • for example, some sulfide ores decompose Development - elaborate development for loading the mined ore • develop a haulage drift parallel to the bottom of the stope • drive cross-cuts from the haulage drift into the bottom of the stope • undercut the stope, about 10 m above the crosscut level • finger raises and cones to be developed from the undercut to the crosscuts



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• This can be further simplified as in sublevel stoping. Instead of finger raises and cones, one could alter the undercut to allow loading from the crosscuts • Widely spaced drawpoints imply more ore left behind in the stope. This will be fine if the sill is mined later. - develop a raise, for access and ventilation, from the haulage drift to the main level at the top of the stope - stope height usually no more than 75 – 100 m Production - overhead drilling and blasting - due to uneven floor, mechanization not possible • air-leg drills and stopers Ore Handling - direct loading from chutes above into cars or use draw point system with LHDs Comments - capital is held up in the stope for a long period - labor intensive as mechanization impossible and, therefore, low production - dangerous working environment - however, viable for a small mine where investments are low



Example: Mouska Gold Mine (Canada) (Underground Mining Methods.., Hustrulid and Bullock, 2001) Orebody - lode type deposit, with mineralization in narrow quartz veins (about 30cms wide). The host rock is diorite (UCS: 175Mpa, Poissons ratio: 0.15, RQD: 80). The sp.gr. of ore is 2.8 - cut-off 9.8 g/t of gold - Shrinkage stopes are 1.6m wide Mine - 100,000 tpy @ 5 days/week. Underground employment: 63 - 72% of the production comes from shrinkage stoping, 20% longhole stoping and rest development - Daily production: 400 tons of ore and 200 tons of waste - Access by a four compartment 485m shaft



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- Levels at 60m center and equipped with rail haulage. Mine advances from top to bottom. Stope - An ore drift (or sill) is driven into the ore prior to development of the haulage drift (see figure). This drift serves to explore the profile of the vein. It also forms the bottom level of the stope. Establishment of the vein profile helps in not only reducing dilution, but in stope planning as well (location of drawpoint, haulage drifts etc) o The width of this drift is critical, especially in the first 4 lifts. Dilution is caused by too wide a drift. A typical drift (and hence stope) is 1.5-1.6m wide. o The sill is no more than 30m ahead of the haulage drift. o Slushers are used to bring the ore to the neared drawpoint from where they are loaded by rubber mounted mucking machines into trains. - Drawpoints are 10m long, 2.7mX2.9m in cross-section, and 10m apart. The roof at the intersection of the drawpoint and haulage drift is made higher to accommodate loading into trains. - Remote controlled muckers recover ore between drawpoints. The drawpoints include a small bay for the remote mucker operator. Also, slasher holes are drilled (to be blasted if needed) if remote mucking is necessary. The extra space allows for easy movement of muckers. - Haulage drifts are in waste rock and 10m away from the stopes - Ave width of stopes is 1.6m. Each cut is 2.4m high. - 10 stopes required to produced 400 tons per day, of which, 6 are mining stopes, 2 are development stopes and 2 are in the final phase. - On a given lift (2.4m high), drilling is horizontal (breasting). o Easy to track the vein o Unexploded holes are better taken care of o Better controlled blasts o Miners are not exposed to the ground they are actually drilling - Blasts are about 25 ton Stope Issues - A slusher is used to ensure a level working surface. In some cases, wooden platforms are built for the same purpose. - The tonnage of ore drawn from each draw point is carefully controlled to ensure a level working floor - Safety protocols are followed when drawing ore from the stope - Hangups are common in the narrow stopes even when the ore is finely broken. Mucking is stopped if the surface level of the broken ore has not moved even though about 50 tons of ore has been removed. The hangup is removed by a high pressure water spray.



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The water hose is brought to the desired location on a remote controlled carrier that rides on a wooden rail installed in the back of the stope. - Ore recovery from the stope is about 95%. The empty stope is not backfilled. Ore Inventory - 35% of the ore (swell) removed after a blast - Blasting stops when the blasted ore muck surface is within 5m of the top level - Ore can remain in the stope for almost 8 months. The quantity in the stope can be as much as 10,000 tons. At 15g/ton grade, the value of the trapped ore can be as much as $2.75 million @$550 /oz Ground Control - Problems caused by: o Diorite, as it is brittle under pressure o Intersection of major joints forms destabilizing blocks o Significant changes in dip and strike as that results in changes in stope profile as the stope chases the ore. - Hanging walls, footwalls and backs are bolted in a 1.2mX1.2m pattern using 1.2m long bolts Service Raise - Raises are excavated for supplying the stopes from the upper level with services such as compressed air, water, ventilation and access) - It is driven in ore and connects the drawpoint of an upper level stope. 0



0



- The raise is 2.1mx1.8m and dips at 60 -65 - It consists of a ladder way for men. Ladder segments are 7m in length and offset from each other for safety. - A wooden slide is also installed to allow miners to operate a winch and bucket for transporting supplies such as drill rods, bolts, explosives etc Manway - A stope requires at least two access ways (for ventilation and safety). The service raise (in the middle of the stope) constitutes one access way (access from the top). The other is the manway (access from the bottom) that is typically located at one end of the stope. - The manway is connected to the drawpoint at the end of the stope by a side opening. - It has wood on one side and follows the rock profile on three sides. Sometimes, wood becomes necessary on all sides. - It is raised one lift at a time along with the mining cycle.



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Equipment and Haulage - Primary drilling equipment consists of handheld drills. Rubber tired mucking machines load ore into rail cars - 5.5 ton battery powered locomotives haul seven 5-tonne ore cars to ore passes - Ore is crushed and then taken to the surface on 6t skips. A new shaft is being excavated to access deeper levels. Economics - Capital expenditure (long term development, construction, exploration drilling, and equipment) has been about US$1.7 million per year. - Operating cost at the mine is about US$160 per ounce. Gold price in 2006: $550-650/oz Comments - Labor intensive. Workers tend to suffer from back problems - Equipment used in this type of mining has not changed in the past 40 years



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3.4 Sublevel Stoping Sublevel stoping recovers the ore from open stopes separated by access drifts each connected to a ramp. The orebody is divided into sections about 100 m high and further divided laterally into alternating stopes and pillars. A main haulage drive is created in the footwall at the bottom, with cut-outs for draw-points connected to the stopes above. The bottom is V-shaped to funnel the blasted material into the draw-points. Short blastholes are drilled from the access drifts in a ring configuration. The ore in the stope is blasted, collected in the draw-points, and hauled away. The stopes are normally backfilled with consolidated mill tailings after being mined out. This allows for recovery of the pillars of unmined ore between the stopes, enabling a very high recovery of the orebody. Application - steep orebodies • dip should be more than the angle of repose - strong surrounding walls - competent orebody - regular shape of orebody Development: Significant. Constitutes about 30% of the cost - haulage drift at main level below bottom of the stope - raises to provide access to sublevels and for their development - drilling drifts through ore on the sublevels • when down-hole-drills (DHD) are used, longer and larger holes can be drilled. Therefore, fewer levels are required • also for DHD, the holes do not fan out and, therefore, stopes may have to widened (Fig 2.6) to ensure that the entire width of the orebody is mined - undercut at the bottom of the stope • the undercut is to provide a free face for blasting and room for blasted muck - loading draw point system at the bottom of the stope for ore recovery • drawpoints must be optimally spaced • they must have smooth floors and long life (shortcrete the walls or use bolts) • avoid secondary blasting - one can combine the undercutting and draw points by locating a drilling drift at the undercut/draw point level (Figure 2.5, p54). Drilling in a vertical fan will result in draw points. This drift should obviously be ahead of the drifts above it. - slot raise at the end of the stope to provide a breaking point for stope initiation • can be raised by raise boring (using a raise borer) or drop raising (blasting in lifts, from bottom up) • widened after initial raising, usually by blasting



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Production - dependent on drilling, usually DHD, electric hydraulic or rotary percussive (and sometimes percussion long hole drills) - drilling diameter from 50mm to 200 mm (2-8 inches), lengths up to 90 m - best pattern is parallel vertical - maximum size of blast is designed with consideration to potential damage - nonel (non-electric) caps are becoming more popular. Various kinds of explosives can be used (dynamite, slurry, ANFO etc) - LHDs used to transport ore from draw point to rail/crusher - Very productive (15-40 st/worker-shift) as drilling can be done well in advance of blasting. Therefore, drilling, blasting and loading are almost independent of each other. - Everything inside a stope is mined. If the geology is not known intimately, mistakes could be made by mining waste/poor quality ore. A typical mistake would be in misjudging the ore/waste boundary (i.e. the ore body outline).



Example (Mine C). - Iron ore, ~137 m thick, steeply inclined - Stope dimensions are 137m x 48m x 19m - 10,600 short tpd, at 3 shifts/day, 6days/week - underground employment of 400 with a productivity of 26.5 st per worker-shift



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Development and Mining of Stope - The haulage drifts are located outside the ore at 46 m vertical intervals - 60 ft wide and 45 ft long stopes are transversely opened across the ore body. The pillar width between stopes can vary from 60 to 70 ft. - A 12 ftx12ft undercut is driven at the bottom of the stope (along the centerline). Holes are drilled upwards (in a fan) to create a funnel for the stope. Loading crosscuts are connected to this undercut by means of drawpoints. - Four drifts, parallel to each other and to the undercut, are driven in the top level. These drifts are used for production drilling in the stope. - To start production, a slot raise in driven in the footwall of the stope. It is widened to the full width of the stope. - Fan holes from the bottom undercut and the downholes from the top of the stope are drilled in the same vertical plane. Drilling and Blasting - bottom fan holes are 76 mm in diameter and 18.2 m high vertically • for the holes, toe burden of 2.5m and ring burden of 1.2m - two holes (28.3 m each) drilled per drift on the upper level, using 114 mm (4.5 in.) percussion drills - productivity is 16.5 st/m of hole drilled - ANFO or cartridged slurry initiated by detonating cord Ore handling Underground Mining| 62



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- 6.1 m3 LHD’s used (with smaller buckets (4.9 m3), however, as the ore is very heavy). LHD productivity is 200-250 st per machine per shift, with a 305 m haul (1000 ft). - Hoisting is by means of a 19-st skip, operating in a 732 m shaft (2400 ft). A separates shaft for waste rock (about 45000 tons/annum) - Ore is sent to the underground crusher (gyratory) in the bottom level. After crushing, it is sent on a conveyor belt to storage bins (underground), from where it is loaded onto the skip. Case Study: Pyhasalmi mine, Finland (Underground Mining Methods.., Hustrulid and Bullock, 2001) Orebody - Copper (1%), zinc (1.8%), sulfur (38%) ore. 3400 st/d with 270 employees. - Surface to 1400 m depth, inclined at 700. Initially opened as a surface mine in 1962. - Length: 650 m, width: 100m. Specific gravity of the ore is in the range 4.0-4.5 - High horizontal stresses (70 Mpa or 10144 psi)). The ore is competent (92-202 Mpa). Host rock is around 180 Mpa (26000 psi) with RQD of 90. Upper levels consist of weak rocks (60 Mpa). o Almost all drifts require shotcreting and grout bolts. Cable bolts (5-10m long) are used in ore drifts. Access - Surface to 470m via main shaft. 470m to 702m via second shaft. A 1:7 decline also connects the 1400m level to the surface. Mining Method - Bench stoping (above 1050m level) and sublevel stoping. - Sublevel stopes are developed along the strike (longitudinal stopes) when the thickness is less than 20m. Else, the stopes are transverse. In the deeper parts, the stopes are parallel to the principal stress direction.



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- Sublevel stopes are 50 m high. The pillars between the stopes are bench stoped in 25m high stopes. The sublevel stopes are backfilled to mine the pillars. The sublevel stopes are narrow to reduce backfill costs. A primary stope is typically 40,000 tons. - A 4.5mX4.5m drift is drilled in the top level. Two parallel drifts are driven in the sublevel and in the loading level. The top level drift is sometimes widened to 11-12m so that an oval stope profile is created. This increases stability and mining recovery. - The holes in the sublevel stope are 25m downholes. They are 76 mm (3 inch) in diameter. Production rings are 3m apart, with toe spacing of less than 2.5m. Powder factor is 150g/tonne (0.30 lb/st). Two Tamrock drill are used for production drilling. One would suffice, but due to high horizontal stresses, the holes get blocked easily and have to be re-drilled.



- Primary stopes are backfilled with consolidated material (classified mill tailings, furnace slag and lime) that is piped through diamond drill holes. The secondary stopes are backfilled with waste rocks (from development activities) by means of LHD’s. - Four LHD’s are used for production (0.9 mtpy). 3 of these are remote controlled while one is video controlled (used during poor visibility). 15% of ore is loaded by remote controlled loader. By leaving an ore triangle, use of remote loaders were reduced. The ore triangle directs the ore from the upper levels to a limited area within the loading level. This triangle is mined later. - The biggest mining cost is hauling since the ore is carried from mining areas to the crusher in 660m. The average haulage distance is 3.5 km. Ore Transport - In three stages. Trucks carry the ore to the underground crusher. It is hoisted to the surface in two stages in skips. A new shaft has been constructed to connect the 1400m level to the surface.



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3.5 Vertical Crater Retreat Mining Crater Blast: This theory propounds that when a spherical charge is placed at an optimum depth below ground, it will break the maximum volume in the shape of an inverted crater. The optimum depth, N is obtained from the relationship: 1/3



N = EW where E = strain energy factor (a constant) and W is the weight of the explosive. In practice, experiments are conducted where the rock type and explosive weight are kept constant. A spherical charge is a column of explosives whose length is less than 6 times the hole diameter.



Result of an experiment: A 4.5 kg of explosive charge was blasted in the same rock, once as a charge with length-to-diameter ratio of 15:1 (cylindrical charge) and again with a length-todiameter ratio of 2.7:1 (spherical charge). While the crater radii were close in both cases, the 3



3



crater volume was significantly different, being 1.1 m for the former and 4.4 m for the latter. Use of spherical charges restricts the maximum amount of explosive that can be loaded into a hole per column segment. In one mine, it came to about 34 kg/hole/segment.



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Application: Started as a pillaring technique - steep dip and large vertical extent - regular ore boundaries Development - diamond drilling to obtain ore boundaries - the top level and the bottom level are cut • the distance between the top level and the bottom level depends on ore consistency, drilling accuracy, accessibility and rock strength. No more than 34 kg of explosives in each of these segment stemming The top level must be large enough for drills to operate • A cut is sometimes made just above the bottom level to minimize blasting damage - crosscuts made from haulage level to bottom level, for ore removal - development equipment can be jackleg drills or jumbos and LHD’s - large diameter holes drilled from the top level, all the way to the bottom level • hole size selection needs deliberation. Factors that are considered are: ore and rock strength, stope geometry, and production requirement • down-hole-drills are used because of their accuracy and ability to drill large diameter holes efficiently • blasting pattern (or the hole burden and spacing) can vary depending on the stope (usually, about 3x3 m) • hole breakout location checked at the bottom level and if unsatisfactory, more holes are drilled Production - bottom of production holes loaded in sections and blasted to obtain a lift of predetermined height • each hole measured before loading • it is determined if there is a sufficient volume of stope available for expansion of blasted muck (the part of the stope that has the sufficient volume is blasted first. This part is called the slot, and is equal to one-third of the stope volume) • holes are blocked before explosives can be loaded • drilling factor about 11.5 tons per meter in some mines, and the powder factor is about 0.5 kg/ton • explosives must be restricted to a maximum amount per delay



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- broken ore can be left in the stope for support of the hanging wall (as in shrinkage stoping) - mucking by electric or diesel LHD’s. Ore is usually dumped into a ore pass. - ground control of crown pillar when the upper lift is close to the main level Comments - cheap, due to reduced development, ground control, man power requirements and high mining rate



Case Studies Falconbridge Nickel Mines - used for rib pillar recovery (before VCR was employed, cut and fill, and sub-level stoping were employed for pillar recovery) - each pillar is 107-122 m long, 61 m high and 6.7 m wide. - main stopes were 15 m wide and mined by cut-and-fill, the fill being 1:32 cement-fill ratio mixture - hole diameter is 165 mm (6.5 in.) - delays are used in blasting Carpathia Orebody, Australia. - tin orebody, 40 m below surface, holes drilled from the surface 80-90 m long Birchtree Mine, Canada - 38 m long and 33.5 m high stope of width varying between 3 and 9m. 152 mm (6 in.) diameter holes, and 3 m lifts per round. - Satisfactory fragmentation (some large pieces due to over break from hanging wall)



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MODULE FOUR SUPPORTED MINING METHODS 4.0 Cut-and-Fill method Cut-and-fill mining removes ore in horizontal slices, starting from a bottom undercut and advancing upward. Ore is drilled, blasted and removed from the stope. When a stope is mined out, the void is backfilled with sand tailings and cement or waste rock. The fill serves both to support the stope walls and provide a working platform for equipment when the next slice is mined. There are two types of cut and fill mining – overhand and underhand. In overhand cut and fill, the ore lies underneath the working area and the roof is backfill. In underhand cut and fill, it is the opposite, the ore overlies the working area and the machines work on backfill. Cut-and-fill mining is applied to steeply dipping orebodies in stable rock masses. It is a selective mining method and is preferred for orebodies with irregular shape and scattered mineralization. Because the method involves moving fill material as well as a significant amount of drilling and blasting, it is relatively expensive and therefore is done only in high grade mineralization where there is a need to be selective and avoid mining of waste or low grade ore. Application - steeply dipping orebody - reasonably firm and competent orebody • use of hydraulic fill makes it possible to apply this method to a vast variety of ore/rock type - no restrictions on ore boundaries Development - haulage drift along the orebody at the bottom of the stope - stope should be undercut - manways and raises to the undercut - ventilation, transport and service raises for the stope - ramps for access to the stope Production - Two major drilling options: overhand drilling and horizontal drilling - Overhand drilling • Vertical or inclined holes drilled upwards • Large amount of drilling possible as face is available. Therefore, blasting rounds can be large



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• High headroom after blasting. Back may be ragged and difficult to control after blasting - horizontal drilling • simple breasting • after filling, only a narrow gap between fill and previous back • drifting jumbos can be used for drilling • drilling/blasting round smaller than overhand drilling method as face area is limited. However, due to the mobility of modern drilling and blasting equipment, this does not affect efficiency • more even face as holes are horizontal • headroom is optimum rather than large (as in overhand drilling method) • horizontal drilling offers selectivity so that poor grade ore can be left behind untouched Ore Handling - ore is brought to the ore pass - LHD’s or other rubber tire mounted equipment are used as the floor is smooth Comments - wide range of applications due to selectivity, recovery and applicability in weak rock - filling interrupts production, even though this is significantly reduced with hydraulic fills Cut and Fill at Mt. Isa, Australia† Orebody - Ag-Pb-Zn deposit - Cut-and-fill used in all the narrow and long orebodies - Width restricted to 3-11m Development - stope divided into two equal divisions wherever possible, and sometimes even a third section (for a very large stope) - access to the stope is by means of footwall raises (inclined at the same dip as the orebody) - main service raises (3x2.4 m) are at the center of the stope • the service raise contains air, water, fill and fuel lines as well as telephone, welding and firing cables - vent raises are located at the end of the stope - raise wall is supported by rock bolting or meshes



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Stope Preparation - initial lifts started at the sill, 12 m above haulage level - horizontal stripping is initially done to obtain full width (11 m) - fill at the bottom slices contains 10% portland cement to form a firm base and to enable future floor pillar recovery with minimal dilution - slice height is about 3.7 m Drilling - uphole drilling used in some stopes • twin boom longhole percussion drills with hole dia. 48-51 mm and hole length of 4.5 m to give a slice of 3.7 m - horizontal holes (flat backing) used in stopes where strata does not permit vertical holes, and control of hanging wall is essential • hand held drills used from above blasted material • drill hole length is 3 m. Four rows are used to obtain a vertical lift of 3 m Blasting - occurs from return air raises to the center of the stope - for up holes, ANFO with electric detonators. 50 holes (6 rings) are blasted before mucking becomes essential - powder factor: 0.2 kg/ton Ore Handling 3



- 3.8 m LHD’s dump into ore passes that are 90-120 m apart 3



- pneumatic chutes control ore flow into 7m cars. The cars are pulled by 20t locomotives - the ore is crushed (primary and secondary) and hoisted to the surface Filling - filling follows mucking - all ore passes are closed off with 1-1.5m concrete brick bulkhead - filled until a 3 m headroom is left - water must be drained from the as soon as possible so production can start and fill pipes extended - drain towers are placed above drain raises to filter the water that is drained Fill material - deslimed mill tailings, 69% solids by weight - interlevel flow is though 100 and 150 mm pipes (4in. and 6in.) - when more support is required (such as in the sill), 105 portland cement is used



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Ground Support - rock bolting of back using 16mm tension bolts of lengths 1.5 - 2.3 m - hanging wall rock bolts are 2.3-3m long (metal straps and wire mesh used when necessary) - crown pillar will be cable bolted by 18 m long cables Ventilation - return raises located at ends of stope, while the service shaft is the intake 3



- airflow is 12-14 m /sec for each stope - circuit fans exhaust air from return raises whenever necessary Man power - labor intensive (three mining crews) - production: 33 st/miner shift and 41 st/miner shift in uphole and flat back methods - mucking: 364 st/miner shift 4.1 Underhand and Overhand Methods The process of breaking and extracting the ore from the vein was known as 'stoping'. It could be conducted in two ways. The ore could be cut from the floor of the gallery, thus proceeding in a downward direction, with the 'deads' or handsorted rubbish being thrown up on to 'stemples', which were wooden or stone arched platforms erected across the workings above the miners' heads. Alternatively, the ore could be cut from above the roof of the gallery, thus driving in an upward direction, with the rubbish falling on the floor for sorting and packing into exhausted spaces or on platforms below the miners' feet. The former downward method was known as 'underhand stoping' and the latter upward method as ‘overhand stoping’. With both methods the miners behaved rather like moles, cutting the lodes from above or below them and filling in behind with the handsorted waste material which it was not worth hauling to the surface for dressing. On occasions, when the lodes were wide and the ground unstable, this packing was also used to support the walls and roof of the workings. Where the waste produced by stoping or dead work was insufficient to provide such packing, it was not unusual for mines to quarry rock at the surface and take it underground to supplement the normal supply. This was done at the great Van lead mine in Montgomeryshire during the 1870s, for example, where several men were kept constantly employed at the quarry adjoining the mine in obtaining slate which was trammed and tipped down two specially designated rubbish shafts for distribution underground. Underhand stoping was probably the oldest method and was the usual practice in most mining areas until the late eighteenth century.



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Overhand stoping, offered the advantage of requiring less labour for the disposal of deads; less timber for the storage of deads, i.e., it required only one long stope instead of several successive stages; and greater safety for the miners, particularly where the walls of the vein were unstable and stemples could not be securely fixed. The underhand system appears to have survived only where the lodes were weak and friable and where there were rich silver ores, since it made it easier for the miners to pick over the cut ore and extract the rich pieces of galena or silver. The early predominance of the underhand system probably resulted from the practice in small, under-capitalised mines of simply pursuing ore lodes downwards f rom the surface. It was only with the expansion in the scale of mining in the late eighteenth and nineteenth centuries and larger initial investment in sinking shafts and opening gallery systems as a preliminary to full-scale working, that ore lodes could begin to be approached from below.



4.1.1 Overhand cut-and-fill



In this method, two level drives are first connected, the lower and upper one by a raise, from the bottom of which mining is begun. The work proceeds upwards, filling the mined-out room, but in the filling, chutes are left through which the broken ore falls. In inclined seams the chutes, also inclined, have to be timbered. The lower-level drive is protected either by timbering or vaulting, or by a fairly strong pillar of vein fillings. Stoping in the different cuts always proceeds Underground Mining| 72



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upwards, but as a whole it proceeds between the two level drives in a horizontal direction. Overhand cut-and-fill, esp. in mining irregular orebodies of greater size, is also called back stoping. 4.1.2 Overhand stope



a. Stope in which the ore above the point of entry to the stope is attacked, so that severed ore tends to gravitate toward discharge chutes and the stope is self-draining. b. An overhand stope is made by working upward from a level into the ore above. 4.1.3 Overhand stoping



a. In this method, which is widely used in highly inclined deposits, the ore is blasted from a series of ascending stepped benches. Both horizontal and vertical holes may be employed. Horizontal breast holes are usually more efficient and safer than vertical upper holes, although the latter are still used in narrow stopes in steeply inclined orebodies. b. The working of a block of ore from a lower level to a level above. In a restricted way overhand stoping can be applied to open or waste-filled stopes that are excavated in a series of horizontal slices either sequentially or simultaneously from the bottom of a block to its top. Stull timbering or the use of pillars characterize the method. Filling is used in many instances. Modifications are known as backfilling method; back stoping; block system; block system of stoping and filling; breast stoping; combined side and longwall stoping; crosscut method of working; cross stoping; Delprat method; drywall method; filling system; filling-up method; flatback stoping; longwall stoping; open cut system; open stope and filling; open-stope method; open-stope, timbering with pigsties, and filling; overhand stoping on waste; resuing; rock filling; room-and-pillar with waste filling; sawtooth back stoping; side stoping; slicing-and-filling system; stoping and filling; stoping in horizontal layers; transverse with filling 4.1.4 Underhand stoping



The working of a block of ore from an upper to a lower level; mining downward. The method is particularly suitable for narrow, highly inclined deposits. Syn:horizontal-cut underhand; underbreaking; underground milling. CF:overhand stoping



4.2 Back filling Method Backfill is an increasingly important component of underground mining operations around the world. This module will give a brief overview of the current application of mining with backfill technology. The disposal of mine tailings underground not only reduces the environmental impact but provides the base of an engineering material which can be used to improve both the ground conditions and the economics of mining. Carefully engineered and efficiently run backfill systems can significantly enhance a mining operation. By contrast, badly engineered and poorly run backfill can be a serious impediment to the mine and, worst of all, compromise safety. Underground Mining| 73



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This module will briefly describe how backfill contributes to a mining operation, the types available and some mining methods that use it. It will cover the method of selecting the appropriate backfill system and some of the challenges and opportunities that arise from the selection process. The module will discuss each of the backfill types in turn, commencing with hydraulic fills. The concept of mine tailings disposal will be expanded to include waste rock in order to cover the important rock backfill systems, which are used in the biggest underground mines. The module will then cover pastefill, which has been newly reintroduced into mining. The following terms will be used throughout: 1. Hydraulic backfill Deslimed mill tailings slurries, with densities raised to over 70%Cw (solids by weight). The coarser fractions are placed underground as hydraulic backfill and the slimes rejected to the surface dam. 2. Paste backfill Total mill tailings filtered or thickened to around 80%Cw to which cement and water is then added to achieve the required rheological and strength characteristics. Any rejects to the dam are at the full tailings sizing range. 3. Rock backfill Waste rock from surface or underground and crushed to a typical top size of around 40mm. This can be placed as is or with cemented hydraulic backfill slurry or cement water slurry.



4.2.1 What is backfill?



Backfill refers to any waste material that is placed into voids mined underground for the purposes of either disposal or to perform some engineering function. Waste materials used include waste development rock, deslimed and whole mill tailings, quarried and crushed aggregate, and alluvial or aeolian sand. Other exotic backfill types used overseas include ice and salt. The waste materials are often placed with very lean cement or other pozzolanic binders to improve the strength properties. Gravity based delivery methods are utilised for slurry based systems with the dense tailings slurry being delivered by pipeline to the disposal point in the stope. These pipelines can range from low pressure (less than 1MPa) turbulent flow systems for deslimed slurries to high pressure (greater than 5MPa) laminar flow systems for pastefill. Sometimes backfill only acts as a void filler and needs only sufficient strength to prevent any form of remobilisation. Where backfill is used as an engineering material it requires sufficient strength be exposed by ore pillar mining in tall vertical faces or undercuts. Lean cement addition is used to generate unconfined compressive strengths ranging from 0.5 to 4 MPa. The other essential requirement is that backfill must be of low cost. Typical costs of backfill range from $2 to $20 per cubic metre, depending on the duty required. These costs can be a significant contribution to the operating costs of the mine. Where cemented Underground Mining| 74



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backfills are used, these costs tend to be between 10 and 20% of the total operating cost of the mine and cement represents up to 75% of that cost. 4.2.2 Mining methods using backfill



Early mining methods either left open voids after the ore had been removed or permitted the caving of the surrounding waste rock. Caving methods often resulted in surface subsidence. Temporary and permanent timber supports enabled larger sized workings and there is a rich history of mining in the 18th and 19th centuries using timber alone. One of the earliest records of backfilling in Australia as a discrete technique was the placement of aggregate from lead jig wastes at Mount Isa in 1933 tipped directly from the mill by conveyor to square set timber stopes. This was done both for disposal purposes and for stabilising the working areas by providing an improved platform. Cut and fill mining methods evolved to use hydraulic fill as a bulk filling material. The tailings slurries were deslimed and densified to achieve a reasonable permeability and to minimise the water that would drain out. The fill would provide an effective surface to walk on after a few hours and for equipment to be moved on it within a day or so. The fill would provide a mucking horizon and confinement to the exposed rock below. In some mines, cement was routinely added to improve cohesion but at significant additional cost. Rock and other mine wastes were routinely disposed in cut and fill stopes. At the Black Swan nickel mine in WA, a marker bed of finely graded and compacted sand is placed on the top of the placed rockfill to act as a digging horizon and to prevent the loss of high grade nickel ore into the fill (McGurk & Lock, 19983). At Henty Gold Mine in Tasmania, paste backfill is being placed in the cut and fill stopes (Henderson et al, 19984). The benching method evolved from experience with sub level open stoping and applying the long hole drilling methods to the cut and fill mining areas. At Mount Isa, the cut and fill stopes (MICAF) were initially mechanised to use trackless development equipment (MECAF) and then long hole drilling rigs were introduced to mine several lifts in one go, over short strike lengths. Hydraulic backfill, dry aggregate and mine development waste are used when available (Villaescusa & Kuganathan, 19985). The controlling factors on backfill placement are the need to maintain stable bench sizes and the need for working platforms for the heavy mobile equipment. Long hole open stoping methods with post backfill have become a major low cost mining method around the world. By designing integrated mining and backfill systems, very high ore extraction ratios are possible. Using combinations of cemented and uncemented backfill optimises costs. At Mount Isa, crushed rock was added at high ratios to cemented hydraulic fill to create very lean cement masses. Early experience quickly showed that exposures 40 metres wide and virtually unlimited in height could be created with little risk of fill failure (Leahy & Cowling, 19786). A large number of stopes were filled with cemented hydraulic fill alone where geometry precluded the addition of rock. All stopes not requiring future re-exposure were filled with uncemented hydraulic fill. These stopes Underground Mining| 75



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were an important part of the backfill operations since these stopes often formed low priority “sand dumps” where deslimed tailings could be placed whenever other sources were not available. Considerable quantities of deslimed tailings have been placed underground at Mount Isa through the life of the mining operation. Bloss (19967) records that 64 million tonnes of backfill has been placed into the 1100 Orebody since 1973, over half of which has been deslimed tailings. Coal mining in Australia has not traditionally used backfill, relying instead on caving of the overlying sediments with surface subsidence occurring in shallower operations. By contrast, Polish coal operations have widely used hydraulic backfill to both minimise surface subsidence and to enable thick seam extraction methods (Palarski, 19938). An Australian desktop study (ACARP, 19979) has identified the use of washery tailings to improve the extraction potential of punch mining in highwall coal operations. More recently, Wambo Mining Corporation placed cemented backfill into a series of headings in the path of the Longwall No. 9 at Homestead Colliery near Singleton. The sand, flyash and cement mix was designed to 4MPa strength. This permitted the uninterrupted mining through the heading area without the need for several weeks of shutdown. No ground stability problems were encountered and Wambo concluded that the cost of the backfill project was revenue positive. This technique has a wide range of applications to underground coal mining problems. Lane10 will be discussing details of disposal of tailings into open pits later in this summit. Other uses of backfill in open pits include the early extraction of ore in the crown between the pit and the underground workings and replacing it with a stiff and impermeable backfill pillar.



MODULE FIVE CAVING METHODS 5.0 Longwall Mining



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5.1 The Longwall Mining Process Fig. 1.1, below, shows a cutaway diagram of a typical longwall mine. The main features of the mine are indicated in the key below the diagram. The longwall face is indicated by the number 8 in the diagram.



In longwall mining, a panel of coal, typically around 150 to 300 metres wide, 1000 to 3500 metres long and 2 to 5 metres thick, is totally removed by longwall shearing machinery, which travels back and forth across the coalface. A typical section through a coal face is shown in Fig. 1.2 and a photograph of typical longwall face equipment is shown in Fig. 1.3. The shearer cuts a slice of coal from the coalface on each pass and a face conveyor, running along the full length of the coalface, carries this away to discharge onto a belt conveyor, which carries the coal out of the mine.



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The area immediately in front of the coalface is supported by a series of hydraulic roof supports, which temporarily hold up the roof strata and provide a working space for the shearing machinery and face conveyor. After each slice of coal is removed, the hydraulic roof supports, the face conveyor and the shearing machinery are moved forward. Fig. 1.3 shows the arrangement of machinery on a typical longwall face, with the hydraulic roof supports on the left hand side and the coal face on the right hand side of the picture. The drum in the background is the rotating cutting head of the coal shearer and the chain conveyor can be seen in the foreground.



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Terminology: - Panel width (262m average in 1999, typical range: 200-360m center to center) - Panel length (2470m average in 1999, typical range: 900 – 5280 m) - Head entry (usually 3 entry system) (total width ~ 30-110 m) - Tail entry (usually 3 entry system) - Recovery room - Setup room - Barrier pillar - By the set up rooms, to protect the bleeder entries, solid block, 60-150m, (can be the bleeder system too) - By the recovery rooms, to protect the mains (90-150m, solid block) Bleeders



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Applicability: Depth 60-820m. Works best at greater depths as the roof will collapse easily. Seam thickness 1.1 – 4.0 m (multiple lifts possible for greater thickness) Deposit should be large to justify high costs 0



Seam dip preferably less than 15 Roof must cave. Massive sandstone or limestone can cause problems.



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Panel Width: Effect of increasing the width - reduced number of panels - reduced development costs - increase in recovery rate - increase in production. However, after a certain maximum width (~300m), gains negligible. Operational difficulties such as difficulty in maintaining a straight face line and hence, support line and conveyor line is common. If the face is not straight, roof falls can occur ahead of the shearer.



This is because roof falls/caving typically occur in a straight line. If one side of the face is ahead of the other (or as happens most of the time, ahead of the middle), the caving process starts, hurting the face conditions where it lags behind. - reduced number of face moves. However, each move is longer due to increased equipment to be moved.  Longwall face moves are very intricate and time consuming. Each face contains tens of heavy supports, armored face conveyor (AFC), shearer, stage loader, power center and the regular conveyors. All these have to be disassembled and moved to the new face. It is challenging not only due to the logistics and scheduling problems, but also due to the fact that handling large and heavy equipment is not easy to handle and move in confined spaces. - higher capital investment - higher power and structural requirements at the face - for unstable roofs and unidirectional cutting, longer exposure times for roof between the toeline of supports and the face Number of Panel entries: - minimum of three entries should be planned for the head gate and tail gate, even though, a lot of mines operate with two entries only (per gate). When the three entries are being developed for each gate, it allows for easy movement of man, material and air (one entry for man movement and intake air, one entry – neutral - for material, and



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one for return air). Once the gates are developed, man, material and intake air use the main gate or head gate, while the return is via the tail entry. Size of chain pillars: - One can use yield pillar design (for deep mines) or other pillar design strategies we discussed before. Use ALPS, the NIOSH program. In most mines, this may be done based on past practices. Supports: - face support consists of self advancing supports such as 2-leg or 4-leg chock-shields. Each units’ support capacity ranges from 300 to 1200 tons. Units are about 1.3-1.6m wide. - head entry support is a combination of hydraulic props, roof bolts, or 3-set beams Stress: - front abutment stress is felt about 160 m ahead of the face. However, increase in stress is negligible. The stress starts increasing at a distance of 50-60m from the face. The peak front abutment (1.5-5 times the cover load) is felt 1-5 m from the face. The face itself is de-stressed. - The highest stress is at the ends of the face as this is where the side abutment meets the front abutment stress. In the figure below, “goaf” refers to gob.



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- 13-82 fpm Production: - Coal cut by the shearer gets loaded onto the AFC, which takes it to the crusher. The crusher outputs the coal into a stage loader which dumps it into the main conveyor. - nomograph gives cycle production - unidirectional or bi-directional cutting 



In unidirectional cutting, the shear cuts when it travels from the tail gate to the head gate. In a double ended shearer, the leading drum cuts the coal while the trailing drum is either free or cuts the floor. On the way back to the tail entry, only cleaning of loose coal is done.  In bi-directional cutting, the coal is cut in both directions. - Manpower (about 10-13: 1 supervisor, 6-10 miners, 2 mechanics, 1 electrician)



Equipment details (from a mining machinery company) as in Year 2000 (240-300m wide face): - Shearer, 1400 hp, $2.1 M - AFC, 2400 hp (1200 head gate, 1200 tail gate), $2.9-3.0 M - Shields, $100-130K each, 146 for 240 m face, totaling $17-20 M - Crusher, 350-400 hp, $190K - Stageloader, 350-400 hp, $500K - Face voltage is typically 2300/4160 Total cost of equipment for a face, 240-300 m in width, is approximately $25 M Total power requirement approximately 4500 hp (not including that required for supports). More Details from a Coal Company The shearer costs about $1.75 million (this is a Joy 4L-S). The cutting motors are about 600 hp each and the tram motors are, I think, about 70 hp. A 1000-foot AFC + crusher + stage loader costs about $5 million. (Joy quotes these as a unit.) The AFC is capable of handling 3000 tph and has a total of about 2700 hp (three drives, 900 hp each). The crusher will have another 200-400 hp (depends on the space available), as will the stage loader. 1250-ton shields, 1.75-meters wide, will cost about $110,000 each. (A 1000-foot face will take 176 shields.) Seven of the shields must be larger than the others (they must push the stage loader assembly and the drives at the head gate and tail gate), and will cost about $135,000 each. Don't forget to buy a few spares. Other panel equipment and their power consumption Hydraulic pumps for the shields: Probably about $600,000 (??). These are four units of 200-300 hp (??--sorry, I forgot to ask about these), only two of which operate most of the time. Underground Mining| 83



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Lighting -- about $100,000; Monorail unit --- $???; The "mule train" components (emulsion car, water pressure pump, electrical transformer and electrical control units) -- about $500,000 to $1million. Total power consumption for the face The power center for the face is usually sized for 5000 KVA. The shearer will cut at a steady 3000tph, so you can get some idea of the power draw (pretty close to the 1200 hp capacity). The AFC will also draw (I believe) about 2000 hp on a steady basis when the pan is full. For a longwall mine with three continuous miner units for development, I would use a 20,000 KVA main transformer at the mine. Quote for the entire face is about $33 million. This includes the spare equipment, but not the section belt, the electrical controls or any of the retrieval equipment to move the face. Ordinarily, a spare of everything except for the mule train, the hydraulic pumps and shields is purchased. Then, when the face is moved, only these items need to be moved. The rest of the items are sent out for rebuild after they are taken off the face, and the rebuilt items are then placed in the subsequent face. Thus, there is an AFC, stageloader, shearer etc. for panels 1,3,5,7, etc. and a second set for panels 2,4,6,8, etc. This keeps move time down to 4-5 days versus the 20-30 days if everything must be disassembled, rebuilt and reassembled. (And, the difference in capital cost is only the time-value of money for the items involved.) Here is a summary of the Joy equipment:



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5.2 Sublevel Caving†



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Application - weak walls and strong ore preferred though weak ore can be mined too - steep dip • a vertical dip is best, while dip>60 is fine too. Loss of ore minor in non-vertical steep dips • considerable loss of reserves in flat dips - preferably, the ore and the rock should be easily separable - surface should be amenable to caving (not an inhabited or the watershed area etc) Development - significant. Almost 20% of the ore is mine during development - sublevels are established at 7.6-12.2 m vertical intervals (25-40 ft) and about 10.7 m horizontal intervals



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• the vertical interval is dependent on the drilling accuracy and the dip of the orebody • the horizontal and vertical spacings affect the eccentricity of the cave - the size and shape of the production drift affects the draw • drift should be as wide as possible • should give good support to the back and the brows • if the back is arched, the draw is mostly at the center and none on the sides • if the back has to be arched for ground support, the drifts should be closer • ore remnants left behind (due to being out of reach of LHD’s) increase with height. Therefore, drift height should be as small as possible (usually about 3-3.2 m). - slot raises are driven at the hanging wall end of the production drift all the way up to the next level - haulage levels driven in waste • in wide orebodies, transverse layouts may be used (ore widths should be > 12-15 m) • here, the production drifts are perpendicular to the strike • recovery better than longitudinal layouts • haulage drift in waste (footwall), about 9 m from ore contact. This distance is maintained so that blasting does not occur too close to it. Diamond drilling is done to obtain ore boundaries so that the haulage drift is neither too close nor too far away. o



o



- ramps may be driven at 15 -18 , to production levels to provide access Production - long, fan holes drilled 70-80 degrees forward (about 8 holes totaling 100 m in a ring) o



- side holes, if drilled, should be steeper than 70 • side holes reduce length of holes driven up from the level below • if flat holes drilled to the sides, however, the blasted material from the side holes cannot be extracted from the working level. Therefore, the neighboring fan hole blasted material does not get enough room to expand - brows should be supported if necessary • if brows collapse, ore floods the drift and covers some rings • also, the loader has problems loading as the ore does not stack high • if brows are uneven, the ore funnels down the high spots • if sloughing or high brows are noticed from development, it may be decided to blast more than 2 rings just to advance through the area - if drilling is not accurate, a bridge/arch may be left in the stope • re-slotting may become necessary - powder factor high as blasting is always against blasted muck (almost twice that of blasting against an open face) - LHD’s used for mucking - Good ventilation necessary as all working faces are dead ends Underground Mining| 88



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- Productivity of 36 ton/miner-shift Comments - High dilution from caved waste. Ore losses occur as well since not all mined ore can be recovered. - Ore/waste flow cannot be predicted accurately prior to mining. - Probably the most economical when mining in weak strata - Development openings are not kept open for the entire life of mine. Once a level is extracted, the development openings are consumed. - Safe since all mining activities are in small, protected openings. Besides, a variety of equipment do not interact with each other unlike other methods. For example, the drilling equipment (on development/drilling levels) do not interact with mucking equipment (lower levels). - High degree of mechanization possible. - Method is flexible. For example, production and equipment can be varied due to mobile equipment.



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5.3 Block Caving†



Required Ore Characteristics



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- large massive orebodies (veins should have steep dip) of regular shape with sides dipping steeply - should have proper fracture pattern • For good fracturing, there must be at least 2 vertical joints, perpendicular to each other, and one horizontal joint. At least 50% of the ore should break in sizes less than 1.5 m, as most finger raises are of that diameter. • An idea on potential rock fragmentation is also obtained by various methods such as RQD or MRMR or Laubscher Caving Stability Graph. MRMR or mining RMR is similar to RMR but includes mining induced and blasting induced stresses in its rating. The Laubscher graph plots the MRMR against the hydraulic radius to identify stable and caving regions. Hydraulic radius is ratio of the surface area of the unsupported area and its perimeter. - To relate to active mines, Palabora mine (copper) in South Africa has MRMR between 57-70, which is on the higher side for block caving (typically block caving is not advised for MRMR over 50). The Henderson molybdenum mine outside of Denver, CO, has an average RQD of 49, while RMR ranges from 27 to 60 respectively. • Note that regional stress fields also have a say in the fragmentation pattern as well as the stability of the blocks. - should be able to withstand undercutting - no restrictions on grade, though usually used on low grades Required Cap Characteristics: Cap is the waste rock above the ore - the cap should be caveable • to prevent sudden massive failure • to transfer overburden weight to ore so ore is crushed. If the overburden weight is not transferred to ore, then ore pieces are large • to prevent weighting on excavations near production area - the cap should not break into fine pieces as that dilutes the ore. Ideally, it should break into small pieces and be resistant to attrition - the surface/overburden should be amenable to subsidence (low/no rainfall preferred, no inhabitation, not under a river, not under a large water table) Development - extensive. requires • undercutting  first, several parallel drifts driven at undercut level  long holes drilled and blasted from these drifts. Drawn from draw points • in some cases, cones, grizzly level, and finger raises • production level and/or haulage level



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• for trackless mining, some of the development is reduced (see Fig. 2.14 in the textbook) - block sizes depend on ore characteristics • when ore is weak or highly fractured, small blocks are preferred as a large block may not be able to take the undercutting • when ore is of medium strength, panel caving is best • for strong ore, mass caving is used as a large undercut is necessary to get caving started - the height of the block should be as high as possible as i) development per unit height gets reduced and ii) ratio of capping to ore reduced • the height also depends on ore geometry and strength - the following should be considered in deciding the draw point spacing: • ore breakage sizes. The area affected by a draw point or its zone of disturbance is small when ore pieces are small. Therefore, draw points should be close when small ore pieces are anticipated. The opposite is true for large pieces • the zone of disturbance for adjacent draw points should overlap so that no ore is left behind • to ensure quick loading of trains, the spacing should be adjusted so that multiple cars can be loaded at the same time Production - a balance should be maintained between draw and caving • this is to ensure that uniform contact is maintained between broken ore and ore/cap above • it also reduces dilution - if a stable arch forms, making caving difficult, widen to re-start caving • in rectangular openings, widening the length may not help if the arch formed over the breadth - draw rapidly following an undercut to ensure no pillar is left as support - control draw to control dilution Caving - initial caving area about the same as undercut area in weak rock. For strong rocks, it is smaller than the undercut area - eventually, the caving area extends beyond the undercut area, following a 45 degree line from the undercut area - most secondary blasting is done during the first 30% of caving - sometimes the boundary is weakened to limit spreading of caving



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Equipment - designed for high production • multi-boom drill jumbos • high tonnage mucking • large scapers with low scraping distance, or • LHD’s - good haulage system • long trains with big cars can be used for haulage • in-pit crushing and conveying if long hauls



Advantages - cheap, as little drilling and blasting - drilling and blasting may be higher if the fragmentation is bad. In Palabora, 70% of the ore may require secondary breakage in the first year. Depending on layout, mines can also use rock-breaker type equipment for breaking large fragments rather than blasting. In Palabora, a special remote controlled high reach (21m) drill rig is used to d&b high hang-ups. - centralized production leads to easy supervision and safe working area - easy ventilation - If diesel LHD’s are used, ventilation requirements get stringent - high production rates - Some mines cites 317 tons/hr from their LHD’s - good for low grades undercut caving limits Disadvantages - narrow range of applicability - high initial (development) cost - difficult to maintain drifts in production area - sudden increase in demand difficult to accommodate - stoppage of drawing may close ore block due to weighting. Stoppage typically happens when product prices go down (example: Henderson mine and price of molybdenum) Underground Mining| 96



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- ore recovery could be low in adverse situations. - bad draw practices may lead to high dilution. - difficult to switch method of mining once started



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