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sulfuric acid is generally used as an acidic lixiviant which is more powerful and places more metals into solution generally leads to faster leaching rates and can be necessary in specific types of deposits the choice of lixiviant depends on the metallurgical characteristics of the deposit specifically acid consuming gangue the economics associated with recovery rates the groundwater regime and the sociopolitical acceptability of using an acidic lixiviant in addition it is necessary to consider other groundwater uses and class of use within the region as well as the potential for self amelioration of the deposit once mining has ceased processing method uranium contained in the pregnant lixiviant extracted from the ore body is pumped to the processing plant where it first undergoes ion exchange to load the uranium onto the resin the now barren lixiviant is recycled to the well field the loaded resin is stripped of the captured uranium through an elution process the eluate is subjected to precipitation and filtering processes which generate yellow cake slurry that is dried in a low temperature vacuum dryer dehydrating the product while controlling emissions by capturing particulates after drying is completed the yellow cake is packaged and stored for shipment in some cases where salinity precludes the use of ion exchange solvent extraction methods may be utilized the recovery of uranium within the processing plant is about 98 the design of the plant is typically specific only to uranium but other by products can be produced uranium deposits typically contain other metals such as molybdenum vanadium and copper and it may be possible to economically extract these metals from lixiviant solutions the overall processing plant can be designed to include modules to either remove or recover these metals as is economically feasible the uranium production process generates waste products under u s regulations these solid waste products are classified as 11 e 2 material and must be disposed of in a nuclear regulatory commission licensed disposal facility the primary methods for liquid waste disposal are by deep disposal wells or by evaporating the liquids and disposing of the solids under the directives of regulatory requirements in australia liquid wastes are disposed of in the mining zones and solid wastes are disposed of in lined disposal facilities under the directives of the state or territory governments groundwater restoration the ideal goal of groundwater restoration is to return the production zone the area of an aquifer where uranium extraction occurs to premining conditions in the event this is not feasible the nuclear regulatory commission regulations allow for the achievement of restoration to the aquifer s quality of use class prior to mining operations these standards have been deemed to be protective of public health and the environment the type of reagent used determines types of contaminants

to the aquifer isr using acidic reagents causes a drastic decrease in ph and dissolution of other minerals in addition to the target uranium ores concentrating most of the elements present in the surrounding rock to levels above the maximum admissible for drinking water supply microbes in the sulfide environment of the leach zone can produce a reducing environment which may cause metals to precipitate in their sulfide forms meanwhile bicarbonate reagents selectively dissolve the uranium ore and contamination is noticeably lower however radium selenium and arsenic are still issues that must be addressed during remediation two approaches can be used for groundwater restoration direct cleaning and self cleaning as described in table 11 5 8 in kazakhstan the mining aquifers are allowed to attenuate naturally whereas in australia some aquifers have been considered unusable for human or animal consumption and hence require no restoration more recently however some of the proposed isr projects are in areas that may have better quality and more valuable groundwater resources in these areas it will be necessary to perform some measures of restoration to at least initiate increases in ph and return to class of use current production according to the world nuclear association primary global uranium production for 2008 was approximately 52 000 t 114 million lb of u3o8 world nuclear association 2010 primary producers include canada at 10 600 t 23 4 million lb kazakhstan at 10 000 t 22 1 million lb australia at 9 900 t 21 9 million lb namibia at 5 200 t 11 4 million lb russia at 4 200 t 9 2 million lb niger at 3 600 t 7 9 million lb and the united states at 630 t 1 4 million lb most of the production from the united states and kazakhstan comes from isr production which is recognized as the most costeffective and environmentally friendly method of uranium production uranium is primarily used in nuclear reactors throughout the world total demand forecast which for 2009 was 77 000 t 170 million lb u3o8 is met by both primary production from mining operations and secondary supply which includes drawdown from inventories and highly enriched uranium sources nonetheless most researchers forecast that demand will continue to outstrip supply unless more uranium mines are placed into production or expanded frasch sulfur mining the process by which elemental sulfur is produced from a geological formation is singularly credited to herman frasch his concept of the mechanism of melting sulfur and then airlifting the molten liquid to the surface required a persistent effort in technical development and an expenditure of time and capital for a span of more than 10 years before realizing significant success the initial production of sulfur by the airlift frasch process began in 1896 with the operation of three wells that produced 2 134 t 2 352 st of sulfur haynes 1942 from a worldwide standpoint however

the frasch process is applicable to situations where fuel and water are available and the geological circumstances are favorable for example the middle eastern petroleum production areas poland and russia basic requirements for frasch sulfur mine three basic resources are needed to develop a profitable frasch sulfur mine 1 large deposit of ore of good grade 5 sulfur is suggested as a lower limit 2 adequate reliable and inexpensive source of water preferably containing little dissolved solids 3 low cost source of fuel other factors such as market price royalty payments tax liabilities and transportation costs also influence the economic feasibility of the project utilities that mine sulfur have been quoted requirements for average water 19 2 m3 t 5 000 gal long ton of sulfur mined fuel 20 000 to 22 300 mm3 of gas m3 2 700 to 3 000 million ft3 of natural gas million gal of mine water and power 118 to 138 w m3 600 to 700 hp million gpd of mine water oil and gas journal 1967 reliable and economic transportation systems are also essential for distribution of the bulk sulfur in this regard transportation and storage of molten sulfur has been growing particularly where low cost water transport can supply shore based distribution terminals sophisticated production personnel are needed to operate and maintain the production and delivery systems ore deposit the frasch process was originally applied to removing sulfur from deposits associated with salt domes formed by the intrusive flow of deep bedded salt that penetrates the upper layers of sediments typically the sedimentary layers are mud shales limestone gypsum and anhydrite the salt plug normally is covered by an insoluble mineral cap below which the sulfur is found in the vuggy porosity of the limestone the sulfur location in the cap rock of the domes is highly variable and not all salt domes contain sulfur bearing rock sulfate reducing bacteria in conjunction with hydrocarbons and water are clearly capable of producing elemental sulfur thus sulfur can be formed from the gypsum and anhydrite and be precipitated in the porous limestone with the progression of water circulation the west texas deposits of sulfur are associated with layers of gypsum anhydrite limestone and dolomite these deposits are not associated with salt plugs and the porosity and permeability of sulfur bearing beds follow the natural layered porosity of the strata and fracture system zimmerman and thomas 1969 hentz et al 1987 field development efficient extraction of sulfur from the deposit depends on several interrelated factors the hot water has to be introduced in such a way and at such a rate as to melt the sulfur so it will form a pool below the hot water which is then airlifted to the surface the piping system consists of concentric pipes which extend into the mineable formation figure 11 5 5 the operating strings are a series of these concentric pipes the

outer casing is set into the cap rock above the sulfur bearing strata the next string conducts the hot water and is set through the sulfur bearing strata and is perforated at the bottom and also at some distance up the pipe another pipe is set concentrically within this pipe and an internal seal is placed between the upper and lower perforations so that the flow of hot water is above the lower level at a higher level another pipe is run concentrically the piping diagram in figure 11 5 5 shows how hot water is introduced first melting the sulfur which is heavier sinks to the bottom runs into the lower pipe and fills the second string the molten sulfur is then airlifted with the compressed air and emerges as shown in the figure the molten sulfur is sent to storage which can be handled in a number of ways since it may contain solid materials or even small amounts of acid hydrocarbons further processing may be required steam heated leaf filters are commonly used for sulfur filtration precoating usually with diatomaceous earth is almost universal the precoat should do the actual filtration to minimize blinding and will also neutralize some acid and reduce corrosion however small amounts of hydrated lime or ammonium bicarbonate may be added to the filter feed to neutralize acid such a system will produce sulfur that is free of ash and acid if hydrocarbons must also be removed activated clay may be added to bleach the dark sulfur other filter aids are also available to remove hydrocarbons filtration rates and precoat requirements will depend on the quality of the sulfur coming from the wells the sequence of the development of the well is shown in figure 11 5 6 the initial injection melts a pool of sulfur and seals the bottom as shown hot water is forced farther up the string and a larger cylinder of rock is heated melting more sulfur which is airlifted to the surface this yields a concentrated flush production this scenario is simple in concept but in practice where variability of permeability sulfur content heat flow problems and the influence of other wells exist the actual operation is complex it is necessary to reinject water produced in bleed wells to control flow save heat and dispose of wastewater the mining process is highly variable because of the nature of the deposits and the uncontrollable nature of the reservoir system makes the process something of an art that is gained by years of experience attempts have been made to apply reservoir engineering by substituting paradichlorbenzene as a meltable material for sulfur with some interesting insights rayne et al 1956 as mining proceeds subsidence and collapse of overburden become a problem that can sometimes be helped by reducing the volume of hot water necessary but may also shear tubing and cause many difficulties wells are often separated into groups to help control these problems and avoid total loss of field control corrosion due to

molten sulfur contact is quite negligible when the sulfur pool is drawn down to where the hot water and air come up together the water flashes as steam and a blow takes place hot water rising up the sulfur production string produces a corrosive environment and with a blow it becomes appreciable the salinity of the water influences the rate markedly if seawater is used and a blow occurs the corrosion rate is disastrous hackerman and shock 1947 shock and hackerman 1949 one of the practical solutions to controlling corrosion has been to use specially produced cement lined pipes as well as limiting the salinity of the water treatment of the water to minimize corrosion and scale problems is essential bleed wells are required to balance the water pumped into the formation at the production wells location of the bleed wells will depend on such site specific factors as the depth permeability and hydrology of the sulfur bearing structure because of environmental concerns bleed water should be recycled to the maximum extent possible when required disposal of bleed water needs to be carefully planned because of the high concentration of dissolved hydrogen sulfide and other sulfur bearing compounds aeration in ponds chemical treatment and use of deep disposal wells are all methods of handling these waters facilities needed at frasch sulfur mine when a sulfur deposit has been located and defined by exploration drilling a water supply located and a fuel supply established the design and construction of the needed facilities can proceed typical layouts showing both the well field and surface facilities are provided in the literature oil and gas journal 1970 most mines will require some or all of the following pumps pipelines and reservoir for the water supply power plant for heating and treating the water compressing the air and generating the electricity and steam equipment for drilling and servicing wells pipeline system for carrying the mine water and air to the wells and the sulfur to collecting stations control stations for the producing wells sulfur collecting loading and storage facilities purification equipment needed to meet customer specifications shipping facilities for solid and or liquid sulfur bleed water disposal system equipment needed for the maintenance of these facilities employee transporting dining and housing facilities if the mine is remotely located offices and buildings for material storage laboratory engineering personnel administration and supervisory staff coal seams were formed over millions of years by the biochemical decay and metamorphic transformation of the original plant material this process known as coalification produces large quantities of by product gases the volume of by product gases increases with the rank of coal and is the highest for anthracite at about 765 m3 t hargraves 1973 most of these gases escape to the atmosphere during the coali

fication process but a small fraction is retained in the coal the amount of gas retained in coal depends on a number of factors such as the rank of coal the depth of burial the immediate roof and floor to the coal geologic anomalies tectonic pressures and temperatures prevailing at the end of the coalification process in general the higher the rank of coal and the greater the depth of the coal seam the higher the gas content gas contents of coal seams generally range from 0 1 to 25 m3 t methane is the major component of gas in coal accounting for 80 95 of the total gas content the balance is made up of ethane propane butane carbon dioxide co2 hydrogen oxygen and argon because the coal bed gas is mainly methane it is generally called coal bed methane cbm methane recovered from active and abandoned coal mines is called coal mine methane coal bed methane production cmp from longwall gobs started in europe in the 1940s but methane drainage from virgin coal prior to mining began in the united states in the 1970s the primary need for cmp was to make the coal mines safe from explosion hazards the methane gas control techniques are discussed in chapter 15 4 in detail as the cmp process developed and became widespread other benefits became apparent in moderately gassy and very gassy mines immediate improvement in coal productivity was realized in some very gassy mines in the united states the productivity improved from a low 15 t worker shift to 40 t worker shift it significantly reduced the cost of mining by minimizing the gas delays caused by excessive gas emissions in the 1980s the volume of cbm produced prior to mining and postmining continued to increase eventually leading to gas processing and commercial marketing extensive drilling of nonmineable coal deposits such as cmp activities in the san juan basin of the united states also produced a large volume of cbm at present nearly 10 51 gm3 billion cubic meters of u s natural gas production is from cmp this number is likely to increase to 20 in the near future because cmp is less cost intensive and financially less risky than deep drilling for natural gas production the cbm after processing to remove impurities is exactly like natural gas with a calorific value of 35 8 mj m3 million joules per cubic meter another financial incentive is driving cmp forward methane gas is a powerful greenhouse gas ghg with a radiative forcing that is 21 times higher than carbon dioxide radiative forcing is a phrase used to measure the capacity of a gas to trap infrared energy in many industrialized countries e g the united states and australia a financial incentive is provided to capture cbm from active mines the current u s incentive is 6 50 t of carbon dioxide thus 1 t of cbm can qualify for a carbon credit of 136 50 if it is captured and marketed cmp today is a standalone profitable enterprise besides saving lives and reducing the cos

t of coal production it provides a clean burning fuel and reduces ghg emissions with adequate financial incentives it can become a major supplier of natural gas in many countries such as the united states china india and australia coal bed methane reserves as illustrated in figure 11 6 1 coal deposits occur in 19 major basins around the world and global reserves of coal and cbm are large coal has been a global energy source for a long time but for the past 200 years it has played a vital role in the growth and stability of world economy at present 70 countries around the world annually mine about 6 500 mt of coal china the united states and india the three largest coal producers derive most of their electricity from coal the total proven world reserve of mineable coal is 1 tt trillion metric tons but indicated reserve to a depth of 3 000 m ranges from 17 to 30 tt landis and weaver 1993 tes but are now prevalent in the rest of the world at present most cmp is realized from relatively shallow 1 000 m deep mineable coal deposits however coal deposits reach a depth of 3 000 m in many basins deep coal deposits are rich sources of cbm but production techniques are not very successful yet because of very low permeability new techniques are currently being developed thakur 2002 the vast global deposit of deep coal can be used in three ways 1 cmp from deeper horizons 1 000 3 000 m 2 sequestration of carbon dioxide a ghg production of cbm from deeper formation is assisted by co2 flooding coal seams have a much higher affinity for co2 hence they absorb co2 and at the same time release methane 3 underground coal gasification ucg coal seams not used for co2 sequestration can be burned in situ with a limited supply of oxygen to produce low calorificvalue combustible gases in situ combustion of coal also greatly enhances methane production by heating the coal formation thus there is a synergy in methane production from deep coal seams co2 sequestration and ucg the subject will be discussed in the order that a typical cmp project is undertaken reservoir properties wire line logging production technology shallow coal medium depth coal deep coal well servicing and maintenance produced water disposal gas gathering and processing co2 sequestration ucg coal bed reservoir properties the most important reservoir properties that not only influence the gas production rate but also determine the production techniques are gas content of coal seams and their gas isotherms permeability reservoir pressure diffusivity of coal water content and quality of water and ground stresses and elastic properties of surrounding strata e g compressive strength and young s modulus of elasticity coal porosity is generally low 1 4 and does not have a significant influence on gas production or production techniques measurement of gas contents the volume of gas contained in a ton of co

al is termed gas content of the coal and is expressed in cubic meters per metric ton m3 t it is generally accepted that the gas is stored in a monolayer on the micropore surfaces of coal the volume of gas contained in coal is dependent on the rank temperature and pressure or depth of the coal seam the micropore surface in coal is large a ton of coal has a surface area of approximately 200 mm2 one cubic meter of coal can store two to three times the amount of gas contained in a typical sandstone reservoir for natural gas of the same volume for commercial gas production it is best to core drill the entire field on a grid pattern and do a direct measurement of gas contents of all coal seams that comprise the gas reservoir a typical spacing for core holes is one in every 200 ha gas content measurement methods are classified as a conventional and b pressurized desorption techniques in the conventional technique coal cores or drill cuttings are retrieved from the core hole and immediately put in a sealed container to measure the desorbed gas the method suffers from the uncertainty in the estimation of gas lost during sample retrieval and handling to eliminate this problem the pressured core desorption technique has been used in this method gas loss is minimized by sealing the coal sample down the core hole both methods provide a positive proof of gas presence and data on the gas desorption rate that can be used to calculate diffusivity to be discussed later desorbed gases are chemically analyzed to determine the composition and calorific value of the seam gas the industry standard for direct measurement of gas content is the u s bureau of mines method shown in figure 11 6 2 kissell et al 1973 coal cores or drill cuttings are deposited in a sealed vessel and desorbed gases are measured periodically until the desorption rate is insignificant total volume of gas thus produced is known as the desorbed gas the cumulative volume of desorbed gas plotted on the y axis against the square root of time yields a straight line and its intercept on the y axis is a measure of the gas lost from the core before it was deposited in the sealed container subsequently a portion of the core is ground in a hermetically sealed mill to release the residual gas the sum of the three component volumes desorbed gas lost gas and the residual gas is the total gas contained in the sample this total volume of gas is divided by the weight of the sample to obtain the gas content of the coal seam total gas reserve gas in place for the coalfield must be established before any commercial undertaking is started gas reserve g ahv where density of coal t m3 a area of deposit m h height of coal seam m v gas content m3 t if there is more than one coal seam in the reserve as is the normal case the total gas reserve is the sum of individual coal seam reserves permeability the permeability of a coal seam is the most important res

ervoir property next to gas content permeability of a porous medium is defined as 1 darcy when a fluid with a viscosity of 1 cp centipoise flows with an apparent velocity of 1 cm s under a pressure gradient of 1 atm cm atmosphere per centimeter the permeability is usually expressed in millidarcies 1 md 1 1 000 of a darcy coal seams can have a permeability range of 1 microdarcy to 100 md depending on the depth of the seam shallow coal seams have higher permeability but deep coal seams show very low permeability reservoir pressure the reservoir pressure can be calculated from the depth of the coal seam but it is best measured directly a vertical well is drilled and cased just above the coal seam the coal seam is cleaned up with high pressure water jets a commercially available recording pressure gauge is installed in the well and the borehole is shut off at the surface the gas pressure begins to build up and in 4 to 6 days the final pressure is reached the pressure gauge is retrieved and the final pressure is read from the chart the coal seam reservoir pressure is typically less than the hydrostatic head about 0 7 times thus a 300 m deep coal seam in the northern appalachian basin of the united states has a reservoir pressure of approximately 2 mpa whereas 600 m deep coal seams in the central appalachian basin have a reservoir pressure of 4 5 mpa the reservoir pressure will rarely exceed the hydrostatic head but if it does it is a very fortunate situation for commercial gas production the fairway area of the san juan basin in the united states is an example of an overpressurized basin many highly productive cbm wells have been drilled and produced in this area diffusivity of coal seams the diffusivity of a coal seam is a parameter that measures how fast the coal seam will desorb the gases when the confining pressure is reduced to the atmospheric pressure it is commonly measured by a parameter called sorption time it is the time a coal seam needs to desorb 63 of the original gas content and is measured in days sorption time for some coal seams in the san juan basin is less than 1 day whereas it can be as high as 900 days for coal seams in the northern appalachian basin rogers et al 2007 coal seams with high sorption time will take a long time to yield a high percentage of in situ gas and should be drilled at a closer spacing and produced for a long time on the other hand coal seams with shorter sorption times should be drilled at a wider spacing greater than one well per 20 ha but produced for a shorter time usually fewer than 10 years water content and quality of water the coal seams are usually saturated with water actual fluid flow in coal seams is a two phase flow invariably the water phase is produced first as the water content depletes the gas production begins to increase reaching a peak when water is almost drained out a typical cbm well can produce from a few cubic meters per day t

o hundreds of cubic meters per day of water the amount of water produced and its quality have a serious impact on the economy of a cbm project and will be discussed later in the chapter ground stresses and rock properties every point in a coal seam lies in a stress field that comprises the vertical stress v the major horizontal stress h and the minor horizontal stress h the comparative magnitudes of these stresses are the determining factors for a successful hydrofracing of a vertical well for successful hydrofracing the sequence h v h must be true this results in the most favorable vertical fracture because the fracture is always orthogonal to the least stress it will also travel in the direction of h that is the azimuth of the fracture is parallel to h major directional permeability of a coal seam is also parallel to h refer to chapter 15 4 rock mechanics properties such as compressive strength elastic modulus and poisson ratios are also needed for a successful design of gas production techniques wire line logging the vertical boreholes drilled to establish the gas reserve and determine reservoir properties are also used for wire line logging the process involves lowering special sensors down the hole to measure various characteristics of the coal seam and sending the data to the surface using electrical wire line hence the name the most commonly used logging techniques to evaluate coal seams for cmp are gamma ray logging passive and active resistivity logging sonic logging neutron logging density logging and induced gamma ray spectrometry logging all these techniques had their origin in the oil and gas industry they are only slightly modified if needed for coal seam logging details of their application are described in literature scholes and johnson 1993 figure 11 6 3 shows basic log responses for multiple coal seams in a vertical well for coal seams the gamma ray response is low the resistivity response is high the density response is high and the neutron density response is also high most commonly only gamma ray logging is used to precisely locate coal seams in a borehole some tools measure the natural gamma ray emissions from the formation whereas others use an active gamma source and measure the scattered gamma rays the only other logging technique commonly used in cmp projects is sonic log particularly the cement bond log technique several acoustic sources mounted on a tool are lowered into the borehole the sound wave is reflected by the steel casing and the cement in between the casing and the formation the velocity of sound and or the amplitude of sound waves indicate whether the cement is there or only a void is present these two logs are highly recommended to locate the coal seam and to make sure the steel casings are well set all other logging techniques are optional new logging tools that measure the gas contents gas isotherms and directional per

meability in situ are being developed but they are not yet commercially available as of 2010 production technology the cbm resource as shown in table 11 6 1 is huge production strategy varies from basin to basin and even in the same basin based on local reservoir properties since all coal seam reservoir properties are depth dependent the cbm resources can be classified into three broad categories to identify the best production technique for different depth ranges shallow medium depth and deep the main characteristics of these reservoirs are summarized in table 11 6 2 shallow reservoirs shallow reservoirs are characterized by high permeability low gas content low reservoir pressures and low diffusivity vertical stresses 3 mpa are smaller than the horizontal stresses hydrofracing of a vertical well in these reservoirs results in horizontal fractures that are inefficient for gas production the following are preferred production techniques mine shafts with multiple laterals horizontal boreholes at the bottom of the shaft a plan view of this drilling pattern is shown in figure 11 6 4 a 3 m diameter shaft is drilled in the target coal seam and the coal seam is reamed out to a larger size to accommodate drill rigs in active coal mines these shafts can be 6 8 m in diameter and are later used as ventilation shafts the entire mine property may need six to eight shafts to recover coal they can be sunk 10 to 20 years ahead of mining for cbm production depending on the price of gas the marketed gas can defray all costs associated with shaft sinking specific gas production from these shallow coal seams ranges from 10 25 m3 d m of the borehole the specific gas production for a coal seam is a characteristic of the coal seam and is a measure of the rate of gas production per unit length of a horizontal borehole drilled for gas production assuming six horizontal boreholes drilled to 1 000 m an initial production of 60 000 150 000 m3 d can be realized another production technique uses surface based rigs to drill horizontal laterals see chapter 15 4 for details the property is drilled with production wells on a grid pattern typically four to six horizontal laterals are drilled from access wells 1 km away to intersect the production wells to realize commercial cmp figure 11 6 5 a special case of gas production from a shallow reservoir is the powder river basin of wyoming united states coal seams in this basin are very thick 30 60 m and shallow the maximum depth is 200 m the gas is of biogenic origin unlike deeper seams where gas is of thermogenic origin and gas content is quite low 3 m3 t permeability and water content are quite high the production technique used here is a vertical well with hydrojetting but no hydrofracing a typical well completed with a cost of 50 000 can yield gas production rates of 7 10 km3 d making it a profitable venture more than 10 000 wells have been drilled and complet

ed this way in the powder river basin so far the technique can work successfully in similar deposits in montana united states or other places in the world medium depth reservoirs medium depth reservoirs are characterized by low to medium permeability high gas contents relatively higher reservoir pressures higher diffusivity and the coal tends to be of higher rank vertical wells with well designed hydrofracing are ideally suited to such reservoirs for gas production specific gas productions for vertical fractures created in these seams range from 4 to 8 m3 d m the hydraulic fracturing process basically involves pumping a fluid such as water nitrogen foam or a high viscosity gel with fine sand 0 1 2 mm diameter at a pressure high enough to hydrofrac the coal seam under ideal circumstances h v h a vertical fracture is created in the coal from the floor to the roof figure 11 6 6 depending on the volume of the frac fluid pumped the total frac length can be 300 600 m the fracture will typically have two wings of 150 300 m that travel in the opposite directions but on the same azimuth that is parallel to the direction of h for commercial gas production all coal seams with a thickness 1 m are hydrofraced in the vertical interval 500 1 000 m multiple coal seam fracing techniques are well established but beyond the scope of this chapter a typical gas production from each horizon of 1 5 2 m thick coal seam is 3 000 m3 d if three or four horizons of this thickness are hydrofraced in a well gas productions of 9 000 12 000 m3 d can be realized observations of nearly 200 hydrofraced wells and mapping of the subsequent mining of the fracture created can be summarized as follows 1 the fracture volume length width height or lwh is always proportional to the fluid volume q pumped in lwh cq where c is the leak off coefficient it is typically 1 or 2 orders of magnitude higher than that for sandstone or limestone reservoirs 2 the fracture plane is always orthogonal to the least stress h or v whichever is smaller 3 the fracture azimuth is always parallel to the major horizontal stress h 4 the fracture always travels up into the roof but not in the floor following the law of physics that the fluid will always flow down the pressure gradient 5 when a coal seam bounded by shale or sandstone is fraced the formation with the smallest modulus of elasticity e will hydrofrac first in fact strong shale and sandstone provide a confining barrier containing the fracture in coal 6 the fracture width is proportional to fluid viscosity e raised to the power of 0 25 figure 11 6 7 shows two methods of completion namely coal bed a open hole and coal bed b through perforations if the coal seam has a high compressive strength the lowest coal seam can be fraced open hole the upper coal seams are fraced through perforations in the steel casing if the coal seam has low compressive strength it is pr

eferable to hydrofrac it with perforated casings the choice of frac fluid varies from basin to basin in the depth range of 500 to 750 m gelled water or nitrogen foam 70 quality has given good results in the central and southern appalachian basin of the united states on the deeper end 750 1 000 m depth and for thick coal seams 10 m cross link borate gels with very high viscosity have produced better results an example is the san juan basin of new mexico and colorado sand concentrations in the frac fluid vary from 0 1 t m3 to 0 5 t m3 depending on the viscosity of the fluid fluids with higher viscosity can carry higher concentrations of sand a special case of cmp from deep and thick coal seams is the cavitation technique figure 11 6 8 shows a completion technique called cavitation through a perforated casing if the coal seam has high compressive strength cavitation can be done in an open hole in the cavitation technique the borehole is pressurized with water and the pressure is released the process is repeated several times in a quick sequence with the hope that coal will cavitate in the depressurizing phase and enhance permeability leading to good gas production cavitation completions have worked well yielding 30 000 60 000 m3 d of gas production in 10 20 m thick coal seams deep coal reservoirs the bulk of the worldwide cbm reserves are in deep coal seams these reservoirs are characterized by high gas content and high reservoir pressures but extremely low permeability a few microdarcies coal is a viscoelastic material an extreme reduction in permeability with increasing depth and hence pressure is a quite natural phenomenon because of depth no horizontal drilling from a mine or from the surface has been done as such specific production of horizontal boreholes is not known but it is likely to be low 1 m3 d m mine shafts are obviously too expensive for these seams even vertical wells with hydrofracing are very inefficient actual hydrofracing efforts in virginia united states at a depth of 1 400 m and in eastern france at a depth of 1 800 m ended in failure the only feasible technique at present is to drill long horizontal laterals from the surface figure 11 6 9 shows a plan view of 12 horizontal laterals drilled in four groups from the same surface location the outlying horizontal laterals in each subgroup are hydraulically fractured at intervals of 300 m hydrofracing should be designed such that the fractures can extend at least 300 m on each side of the vertical well in the beginning all hydrofraced horizontal laterals will have a high gas production rate but after the gas production rate has significantly declined the middle lateral in each subgroup should be injected with carbon dioxide coal seams have a much higher affinity for co2 it will be preferentially adsorbed by the coal matrix releasing methane at the same time thakur 2002 well servicing and maintenance after drilling and completi

on all cbm wells will produce water and gas in that order all wells need some servicing and maintenance figure 11 6 10 shows a typical gas and water production profile depending on the water content of the coal seam a cmp well can produce moderate to large quantities 1 100 m3 d of water as the reservoir pressure is reduced and more water is produced the water production declines and gas production rises reaching a peak after that the gas production rate also declines horizontal and vertical wells need different kinds of servicing and maintenance horizontal laterals drilled from the surface these wells do not need much maintenance the produced water collects at the bottom of the production well if several horizontal laterals discharge water in one production well the volume is large enough to warrant the use of a progressive cavity pump as water production declines the pump can be put on a timer the gas pressure in the reservoir is enough to clean up the borehole this method of well completion still suffers from two drawbacks an inability to stay in the coal seam and an inability to completely dewater the well the first problem is due to drilling instrument limitations the borehole exits the coal seam and often enters the roof or floor limiting gas production and the ability to completely dewater hence the specific production cubic meters per day per meter is generally only 50 or less compared to in mine drilling which is more precise see chapter 15 4 it is also difficult to reenter a horizontal lateral for maintenance especially if multiple laterals are drilled from the same vertical well vertical hydrofraced well the servicing procedure and equipment for these wells is borrowed from the oil and gas industry after hydrofracing a well is generally swabbed clean to remove any debris and a dewatering pump is installed depending on the depth gas and water production rates and reservoir pressure one of the following systems is used sucker rod pump progressive cavity moyno pump submersible electric pump gas lift a sucker rod pump is most often used when water flow is less than 20 m3 d for heavier flows a progressive cavity pump is used in the dewatering phase of gas production the pump is usually kept slightly above the coal seam and a 30 60 m water column is maintained above the coal seam to prevent a blowout i e a sudden discharge of gas and proppant sand submersible electric pumps are used when the water production is excessive and progressive cavity pumps are unable to handle the flow gas lift is used when no electric power is available to operate a dewatering pump produced water disposal all produced water must be metered to determine the flow rate if the water is clean a turbine meter is used to precisely measure the water flow if the produced water contains some solids a positive displacement water meter is used the latter is more durable and commonly used where the water flow

is intermittent and low a 20 l bucket is used to collect the pump discharge and the time of discharge is measured by a stopwatch from which the flow rate is calculated gathering water and its proper disposal is often a significant part of the cmp the following are the possible options for water disposal surface disposal deep well injection evaporation in a pond or a boiler land application off site disposal in abandoned mines reuse for hydraulic stimulations the choice of the technique depends on the volume of water produced per day and its quality permitting from national and state authorities is often needed and it is a prolonged process most commonly only the first two choices are used and they are the only ones discussed here figure 11 6 11 shows a layout of various options surface disposal the simplest method of surface disposal comprises ph a measure of water acidity adjustment of water to meet the permit requirements aeration of water with compressed air and addition of some polymers to facilitate settling the clean water is used for land applications or simply discharged into a local stream the total dissolved solids tds and chloride concentrations are monitored to match the quality of treated water with that of water in the stream if the produced water has a higher concentration of contaminants it is pretreated and the reverse osmosis process is used to remove the tds average efficiency of tds removal is in excess of 95 only in the presence of bicarbonates does the efficiency go down to 50 80 the main limitation of this process is membrane fouling caused by suspended solids or oils in the produced water deep well injection if there are no streams in the neighborhood of a cbm field or the necessary permits to discharge into the local streams cannot be obtained the treated water can be injected into a deep well deep well injection also requires a permit but it is easier to get typically these injection wells are old gas wells with a depth of 3 000 m or so they are often hydrofraced to create good permeability produced water must be properly treated before injection to keep the injection pressure below the permitted limits thick sandstone or limestone reservoirs with high porosity and permeability are ideal for this purpose preinjection filtration is almost always essential scale and corrosion inhibitors and bactericides are often added to the injected water to increase long term use of the injection well finding or even drilling an injection well in the vicinity is crucial because the cost of gathering and hauling water can be quite substantial gas gathering and processing cmp must be gathered to a central location for processing and commercial marketing gas gathering and measurement gas flows from individual wells must be measured the primary measurement instruments are either an orifice meter or a turbine meter the orifice meter measures the differential pressure upstream and downst

ream of a measured orifice the pressure drop is a function of the flow rate gas pressure and temperature the pressure differential is continuously recorded on a chart that is calibrated to give the gas production this is the most common method of measuring individual well production the turbine meter measures the flow rate and total volume of gas produced gas flows over rotary blades and magnets in the meter housing pick up signals that are proportional to gas velocity the area between the blades has fixed volumes so that the number of turns equates to the volume of gas produced the turbine meter is more accurate than the orifice meter but it is more expensive and needs frequent maintenance gas from each producing well is gathered using polyethylene pipes and compressors usually a cluster of wells is connected to a small compressor these small compressors deliver the compressed gas to a central compressor facility where it is processed for marketing in the design of the gas gathering system the main object is to maintain a low wellhead pressure usually 3 5 kpa pipe diameters and compressor sizes are designed to minimize line losses in general one may start with 50 mm diameter pipe at a single wellhead but progressively the pipe diameter may increase to 250 300 mm at the main gas gathering and processing plant gas composition and processing for commercial marketing the cbm composition must meet the pipeline specification all major pipeline companies require a minimum calorific value and put a limit on oxygen moisture and noncombustible concentrations cbm composition varies from basin to basin but generally it contains 85 95 methane impurities must be removed to raise the calorific value to about 37 mj m3 other requirements are that the noncombustible content of gas should be 4 moisture should be 0 11g m3 and oxygen is restricted to 20 ppm in many applications cbm is mixed with coal mine methane that contains higher oxygen and nitrogen in which case excess nitrogen is removed by a cryogenic process figure 11 6 12 shows a typical gas processing flow diagram first the raw gas is passed through a cyclone to knock solid particles out next the gas is passed over an oxidation catalyst where methane is oxidized that consumes all oxygen and produces water and carbon dioxide co2 next the co2 is removed by passing the gas through an amine unit finally the moisture in gas is removed by dehydration units containing glycol if there is excess nitrogen in the gas stream it is generally removed by a cryogenic process the clean pipeline quality gas is compressed to pipeline pressure typically 7 8 mpa for delivery into a commercial gas pipeline small rotary compressors are used for primary gathering because they are more efficient in compressing somewhat dirty but large volumes of gas the main station compressors are generally reciprocating piston type compressors that are more efficient for duty the prime

mover for compressors can be electric diesel or even a cbm engine to keep the cost low molecular gate separation system the molecular gas separation system is a recent breakthrough for processing gases where gas molecules differ in size by only a fraction of an angstrom a new family of titanium silicate molecular sieves has a pore size that can be controlled down to 0 1 0 01 nm in upgrading cbm and coal mine methane the desired product methane has a molecular diameter of 3 8 whereas the impurities such as nitrogen 3 6 carbon dioxide 3 3 and oxygen 3 5 are all smaller in size and easily fit within the pore of an adsorbent where they can be adsorbed and removed in one step from the product stream mitariten 2001 figure 11 6 13 illustrates the basic flow sheet for a molecular gate separation process the feed gas is compressed to 1 5 mpa and fed into the molecular gate system several repetitions cycles are made to achieve high purity the product gas comprising methane ethane and propane is recovered at practically the original pressure which avoids recompression the tail gas contains most impurities but often has enough methane that it can be used for steam production the process has a methane recovery efficiency of 95 at 2 4 mpa but it becomes more efficient at higher pressures it can remove co2 n2 and o2 in one step it is available in sizes of 30 000 900 000 m3 d of methane the cost of gas processing declines significantly with increase in the flow rate carbon dioxide sequestration it was mentioned earlier that coal has a great affinity for co2 and it can displace methane stored on the micropores of the coal matrix this property of coal has been used to produce methane from deep coal seams and sequester co2 in it simultaneously laboratory experiments to displace methane from coal using nitrogen helium and carbon dioxide show that the latter is the most effective flooding agent the storage capacity of coal for co2 is at least twice the capacity it has for methane collins et al 2002 figure 11 6 9 shows an ideal layout for co2 sequestration when gas production begins to decline co2 injection can start the optimum parameters of co2 sequestration have not yet been established but many research projects are in progress to determine the following optimum storage capacity for co2 in various coal seams rate of injection and travel velocity of co2 in coal optimum pressure of injection generally the reservoir fracturing pressure should not be exceeded ideal mix of gases co2 by itself is slow moving and nitrogen is often added to accelerate the process economics of co2 sequestration co2 sequestration research is encouraged by many countries because it is considered a ghg in many countries there are financial incentives for sequestering co2 the current carbon finance instrument in the united states is fixed at 6 50 t of co2 two sources of co2 can be related to coal a power plan

ts that burn coal and b cbm itself that may contain 10 50 co2 in some coal seams co2 can be stripped using gas processing techniques discussed earlier and its sequestration may become a profitable venture underground coal gasification cmp from deep coal seams can also be enhanced by heating the coal matrix the drilling layout shown in figure 11 6 14 is a schematic for ucg heating of the central horizontal boreholes can be done with steam or radio frequency devices but the most synergistic approach is to ignite the coal with a limited supply of air oxygen combustible gaseous blends the ucg process is in fact already well established but it has not been done in conjunction with methane recovery in this process the outer horizontal laterals are used for gas production when most methane is driven out the coal matrix is combusted in situ to produce a lower calorific value gas the process is also called in situ gasification of coal the product gas composition depends on coal composition and process parameters such as operating pressure outlet temperature and gas flow these parameters are not only monitored but constantly adjusted to optimize the process the ucg process is quite flexible it can be used in coal seams with a thickness of 1 60 m and calorific values of 10 30 mj kg including lignites and other low quality coal the coal deposits shown in figure 11 6 1 have a reserve of 17 30 tt only 1 tt is amenable to actual mining thus the remaining reserve can be economically exploited by the ucg process ucg experiments carried out in the 1970s by the u s department of energy in west virginia and wyoming united states produced only low calorific value gases 10 20 mj m3 but new processes are being developed to improve the calorific value products of different ucg projects can be mixed with gases with higher calorific values to meet a consumer s specific need cmp and ucg processes supplement conventional mining by recovering energy from coal seams that are abandoned or cannot be economically mined the gaseous product of the ucg process can be used in a variety of ways such as boiler fuel to produce steam for electric power generation feed to chemical plants to produce fischer tropsch liquids or fertilizers and a clean gaseous fuel after the raw gas is processed using techniques discussed earlier in the chapter the u s environmental protection agency is making a great effort to export cmp technology to all major coalproducing countries this is mainly being done to make mining safer and more productive but it also creates a clean source of energy that supplements the natural gas supply additional incentives such as the carbon credit can provide a great impetus to develop this valuable resource even further the combination of cmp with co2 sequestration can be used to enhance cmp from deeper coal seams and eventually using the ucg process to convert coal into natural gas can greatly increase the world sup

ply of natural gas further research is needed to make the ucg process more efficient in the future the in situ coal may also be converted into liquid fuel to increase the petroleum supplygeneral basis of mine planning many details go into the planning of a mine the information gathered must come from several sources first is the geological structural and mineralogical information combined with the resource reserve data this information leads to the preliminary selection of potential mining method and sizing of the mine production from this the development planning is done the equipment selection is made and the mine work force projections are completed all leading to the economic analysis associated with mine planning planning as just described however will not necessarily guarantee the best possible mine operation unless the best possible mine planning has been done correctly any sacrifice in mine planning introduces the risk that the end results may not yield the optimum mine operation planning is an iterative process that requires looking at many options and determining which in the long run provide the optimum results this chapter addresses many of the factors to be considered in the initial phase of all mine planning these factors have the determining influence on the mining method the size of the operation the size of the mine openings the mine productivity the mine cost and eventually the economic parameters used to determine whether the mineral reserve should even be developed physical and geotechnical information needed for preliminary mine planning many pieces of engineering and geologic data must be gathered before mine planning can take place these are covered in the sections that follow this section draws heavily from chapter 2 of underground mining methods bullock 2001 technical information assuming that the resource to be mined has been delineated with prospect drilling study will be based primarily on information supplied through exploration the results of the exploration are recorded in a formal report for use in project evaluation the exploration report should contain the following information with appropriate maps and cross sections property location and access description of surface features description of regional local and mineral deposit geology review of exploration activities tabulation of geologic resource material explanation of resource calculation method including information on geostatistics applied description of company s land and water position ownership and royalty conditions mining history of property rock quality designation rqd values and any rock mass classification work that has been done results of any special studies or examinations the exploration group has performed metallurgical tests geotechnical work etc report on any special problems or confrontations with local populace of the area any other pertinent data such as attitude o

f local populace toward mining special environmental problems availability of water and hydrologic conditions in general and infrastructure requirements this critical information should be established to assist the mine planning if the exploration project is one that has been drilled out by the company exploration team this information should have been gathered during the exploration phase and passed to the mine evaluation team or the mine development group more information on each of these subjects may need to be obtained but if the inquiry can be started during the exploration phase of the project time will be saved during the feasibility evaluation and development phases of the project if this is an ongoing mine operation then most of the technical information will be available from other mine planning projects geologic and mineralogic information knowledge of similar rock types or structures in established mining districts is always helpful in developing the first mine in a new district there is far more risk of making costly errors than in the other mines that may follow the geologic and mineralogic information needed includes the size length width and thickness of the areas to be mined within the overall operations area to be considered including multiple areas zones or seams the dip or plunge of each mineralized zone area or seam noting the maximum depth to which the mineralization is known the continuity or discontinuity noted within each of the mineralized zones any swelling or narrowing of each mineralized zone the sharpness between the grades of mineralized zones within the material considered economically mineable the sharpness between the ore and waste cutoff including whether this cutoff can be determined by observation or must be determined by assay or other means whether this cutoff also serves as a natural physical parting resulting in little or no dilution or whether the break between ore and waste must be induced entirely by the mining method and whether the mineralized zone beyond above or below the existing cutoff represents submarginal economic value that may become economical at a later date the distribution of various valuable minerals making up each of the potentially mineable areas the distribution of the various deleterious minerals that may be harmful in processing the valuable mineral whether the identified valuable minerals are interlocked with other fine grained mineral or waste material the presence of alteration zones in both the mineralized and the waste zones the tendency for the ore to oxidize once broken and the quantity and quality of the ore reserves and resource with detailed cross sections showing mineral distribution and zones of faulting or any other geologic structure related to the mineralization structural information required physical and chemical structural information includes the following the depth of cover a detailed

description of the cover including the type of cover approximate strength or range of strengths structural features in relation to the proposed mine development and presence of and information about water gas or oil that may have been encountered the quality and structure of the host rock back floor hanging wall footwall including the type of rock approximate strength or range of strengths noted zones of inherent high stress noted zones of alteration major faults and shears systematic structural features porosity and permeability the presence of any swelling clay or shale interbedding rqd throughout the various zones in and around all of the mineralized area to be mined out the host rock mass classification rock mass rating rmr or barton s q system temperature of the zones proposed for mining and acid generating nature of the host rock the structure of the mineralized material including all of the factors listed previously as well as the tendency of the mineral to change character after being broken e g oxidizing degenerating to all fines recompacting into a solid mass becoming fluid etc siliceous content of the ore fibrous content of the ore acid generating nature of the ore and systematic fault offsets planning related to physical properties the physical nature of the extracted mass and the mass left behind are very important in planning many of the characteristics of the operating mine four aspects of any mining system are particularly sensitive to rock properties 1 the competency of the rock mass in relation to the in situ stress existing in the rock determines the open dimensions of the unsupported roof unless specified by government regulations it also determines whether additional support is needed 2 when small openings are required they have a great effect on productivity especially in harder materials where drill and blast cycles will be employed 3 the hardness toughness and abrasiveness of the material determine the type and class of equipment that can extract the material efficiently 4 if the mineral contains or has entrapped toxic or explosive gases the mining operation will be controlled by special provisions within government regulations in countries where appropriate regulations do not exist best practice guidelines must be sought the need for a test mine from this long list of essential information required for serious mine planning it becomes evident that not all of this information can be developed from the exploration phase nor is it likely that it can all be obtained accurately from the surface if this is the first mine in this mining area or district then what is probably needed during the middle phase of the mine feasibility study is development of a test mine while this may be an expense that the owners were hoping to avoid up front the reasons for a test mine are quite compelling they include opportunities to

confirm from a geologic point of view the grade ore continuity ore configuration and mineral zoning confirm from the engineering viewpoint rock strengths and mass rock quality verify mining efficiencies confirm water inflows and demonstrate waste characteristics pilot test the metallurgical process enhance the design basis for cost estimates improve labor estimates build more accurate development schedules and lower the mining investment risk having a test mine in place will shorten the production mine development and will serve as a training school for the production mine land and water considerations information needed about the property includes details on the land ownership and or lease holdings including royalties to be paid or collected identified by mineral zones or areas availability of water and its ownership on or near the property quality of the water available on or near the property details of the surface ownership and surface structures that might be affected by subsidence of the surface the location of the mining area in relation to any existing roads railroads navigable rivers power community infrastructure and available commercial supplies of mine and mineral processing consumables drill steel bits explosives roof bolts steel balls and rods for the mills chemical additives for processing steel liners for all mining and processing equipment etc and the local regional and national political situations that have been observed with regard to the deposit other factors influencing early mine planning many operational decisions must be made in planning the mining operation none is more important than sizing the mining operation however it is not an easy or obvious decision to make and several factors must be considered sizing the production of a mine a considerable amount of literature is available on the selection of a production rate to yield the greatest value to the owners including works by tessaro 1960 mccarthy 1993 runge 1997 and smith 1997 basic to all modern mine evaluations and design concepts is the desire to optimize the net present value npv or to operate the property in such a way that the maximum internal rate of return is generated from the discounted cash flows anyone involved in the planning of a new operation must be thoroughly familiar with these concepts equally important is the fact that any entrepreneur who is planning a mining operation solely from the financial aspects of optimization and who is not familiar with the issues associated with maintaining high levels of production at low operating cost per metric ton over a prolonged period is likely to experience disappointment in years with low or no returns other aspects of the problem for optimizing mine production relate to the effect of npv when viewed from the purely financial side of the question i e producing the product from the mineral deposit at the maximum r

ate optimization that yields the greatest return is the case often selected this is due to the fixed costs involved in mining as well as to the present value concepts of any investment still there are practical limitations to the maximum intensity of production arising out of many other considerations to which weight must be given hoover 1909 there can be many factors limiting mine size some of which are listed here market conditions current price of the product s versus the trend price grade of the mineral and the corresponding reserve tonnage time before the property can start producing attitude and policies of the local and national government and the degree of stability of existing governments and their mining policies taxes and laws availability of a source of energy and its cost availability of usable water and its cost cost and method of bringing in supplies and shipping production physical properties of the rock and minerals to be developed and mined amount of development required to achieve the desired production related to the shape of the mineral reserve amount and complexity of mineral processing required availability of nearby smelting options if required size and availability of the work force that must be obtained trained and retained availability of housing for employees in remote locations potential instability of the government in the future which might cause a company to develop a smaller highgrade mine in the beginning until they have received their objective return then use the income from the existing property to expand and mine out the lower grade ores not only does the resource s tonnage affect the mine size but the distribution of ore grade can certainly affect the mine planning unless a totally homogeneous mass is mined it may make a considerable economic difference as to which portion is mined first or later furthermore no ore reserve has an absolute fixed grade to tonnage relationship trade offs must always be considered in most mineral deposits lowering the mining cutoff grade means that more metric tons will be available to mine but the mine cutoff must balance the value of each particular block of resource against every type of cash cost that is supported by the operation including all downstream processing costs as well as the amortization of the capital that was used to construct the new property even in bedded deposits such as potash or trona the ability or willingness to mine a lower seam height may mean that more ore can eventually be produced from the reserve in such cases the cost per unit of value of the product generally increases similarly narrow seam mining greatly reduces the productivity of the operation compared to high seam and wide vein or massive mining systems in evaluating the economic model of a new property after the physical and financial limits have been considered all of the variables of grade and tonnage with

the related mining costs must be calculated using various levels of mine production that in the engineer s judgment are reasonable for that particular mineral resource at this point in the analysis the various restraints of production are introduced this will develop an array of data that illustrate the return from various rates of production at various grades corresponding to particular tonnages of the resource at a later stage probability factors can be applied as the model is expanded to include other restraining items timing affecting mine production for any given ore body the development required before production start up is generally related to the size of the production as well as the mining method obviously the necessary stripping time for most large porphyry copper deposits is very large compared to the stripping time for even a large quarry for an underground example very high production rates may require a larger shaft or multiple hoisting shafts more and larger development drifts the opening of more mineable reserves as well as a greater lead time for planning and engineering all aspects of the mine and plant the amount of development on multiple levels for a sublevel caving operation or a block caving operation will be extensive compared to the simple development for room and pillar operation in combination all of these factors could amount to considerable differences in the development time of an operation in the past the time for mine development has varied from 2 to 8 years two indirect economic effects could result from this 1 capital would be invested over a longer period of time before a positive cash flow is achieved and 2 the inflation rate to time relationship in some countries in the past has been known to push the costs upward by as much as 10 to 20 per year thereby eliminating the benefits of the economy of scale of large size projects to aid the engineer in making approximations of the time it takes to develop shaft and slope entries related to the size and depth of the mine entry the reader is referred to bullock 1998 for example sinking and equipping a 6 1 m 20 ft diameter shaft ranging in depth from 305 914 m 1 000 3 000 ft will take approximately 64 113 weeks depending on the production capacity that the shaft must be equipped for driving a slope to the same depths will take approximately 52 151 weeks respectively for mines that are primarily developed on one or two levels and have extensive lateral development the speed of the development can vary considerably the lateral development on each level of a room and pillar mine opens up new working places and the mine development rate can be accelerated each time a turnoff is passed provided that there is enough mining equipment and hoisting capacity available this is in contrast to a vein type mine that has a very limited number of development faces as the vein is developed for each level and each level requires dee

pening the main hoisting shaft or slope experience has shown that the mine s development only progresses deeper at the average rate of about 50 m a chapter 4 7 of this handbook discusses several examples of the timing of mine developments the timing of a cost is often more important than the amount of the cost in a financial model timing of a development cost must be studied in a sensitivity analysis in this respect any development that can be put off until after a positive cash flow is achieved without increasing other mine costs should certainly be postponed government considerations affecting mine size government attitudes policies and taxes generally affect all mineral extraction systems and should be considered as they relate to the mining method and the mine size one can assume that a mine is being developed in a foreign country and that the political scene is currently stable but is impossible to predict beyond 5 8 years in such a case it would be desirable to keep the maximum amount of development within the mineral zones avoiding development in waste rock as much as possible this will maximize the return during a period of political stability also it might be desirable to use a method that mines the better ore at an accelerated rate to obtain an early payback on the investment if the investment remains secure at a later date the lower grade margins of the reserve might later be exploited care must be taken however that the potential to mine the remaining resource which may still contain good grade ore is not jeopardized whatever the stability situation there is merit in carefully planning to mine some of the higher grade portions of the reserve while not impacting the potential to mine the remaining reserve some mining methods such as room and pillar mining allow the flexibility of delaying development that does not jeopardize the recovery of the mineral remaining in the mine in contrast other mining systems such as block caving or longwall mining might be impacted by such delays similar situations might arise as a result of a country s tax or royalty policies sometimes established to favor mine development and provide good benefits during the early years of production in later years the policies change again as in the preceding case the flexibility of the mining rate and system must be considered but not to the extent of jeopardizing the remaining resource preplanning based on geology and rock characteristics in the previous section the influence of geology from a macro point of view was considered in this section the geology from a micro point of view of the rock and mineral characteristics must be considered in mine planning geologic data using geologic and rock property information obtained during preliminary investigations of sedimentary deposits isopach maps should be constructed for all potential underground mines to show the horizons to be mined and those that are to be left as the r

oof and floor such maps show variances in the seam or vein thickness and identify geologic structures such as channels washouts called wants and deltas where differential compaction is indicated associated fractures in areas of transition should be examined areas where structural changes occur might be the most favored mineral traps but are usually areas of potentially weakened structures where possible locating major haulage drifts or main entries in such areas should be avoided if intersections are planned in these areas they should be reinforced as soon as they are opened to an extent greater than that necessary elsewhere further details on development opening in relation to geologic structures and other mine openings appear in spearing 1995 reproduced by bullock 2001 again referring to flat lying type deposits extra reinforcing or reducing the extraction ratio may also be necessary in metal mines where the ore grade increases significantly since the rock mass usually has much less strength where the pillars were formed prior to discovering the structural weakness it will probably be necessary to reinforce the pillars with fully anchored reinforcing rock bolts or cables it is advisable to map all joint and fracture information obtained from diamonddrill holes and from mine development attempting to correlate structural features with any roof falls that might still occur characteristics of extracted material the hardness toughness and abrasiveness of the material determine whether the material can be extracted by some form of mechanical cutting action by drilling and blasting or by a combination of both methods the mechanical excavation with roadheaders of the borate minerals from the death valley california united states open pit mine operated by tenneco and the later use of the same type of roadheader at the nearby underground billie mine california illustrate the point bucket wheel excavators have been used extensively in germany and australia for stripping overburden from the brown coalfields technological advances in hard metal cutting surfaces steel strengths and available thrust forces allow increasingly harder and tougher materials to be extracted by continuous mining machines the economics of continuous cutting or fracturing as compared to drill and blast gradually are being changed for some of the materials that are not so tough or abrasive however for continuous mining other than tunnel boring machines to be competitive with modern highspeed drills and relatively inexpensive explosives it appears that the rock strengths must be less than 103 400 124 020 kpa 15 000 18 000 psi and have a low abrasivity rock that is full of fractures however is also a great aid to mechanical excavation in one case that was covered in an article about the gradual trend toward mechanical excavation in underground mining bullock 1994 a roadheader was being used in a highly fractured welded volc

anic tuff even though the rock strength was well over 137 800 kpa 20 000 psi at times reasons other than the cost of extraction favor one mining system over another using a mechanical excavation machine is nearly always advantageous in protecting the undisturbed nature of the remaining rock where blasting might be prohibited likewise the continuous nature of mechanical excavation can be used to speed mine production this was seen in the openings driven by magma copper in developing the kalamazoo ore body in arizona united states chadwick 1994 snyder 1994 and by stillwater mining company in developing their original ore body as well as for their east boulder ore body in montana united states tilley 1989 alexander 1999 a continuous boring tool may also be desirable for extracting an ore body without personnel having to enter the stoping area where their use is applicable continuous mining machines are certainly much easier to automate than the cyclical drill and blast equipment the automation of the 130 mobile miner at the broken hill mine in australia is one case in point willoughby and dahmen 1995 another more dynamic case is the complete automation of the potash mines of potash corporation of saskatchewan canada fortney 2001 planning the organization the amount of equipment or numbers of personnel required to meet the needs of all mines cannot always be given in absolute and precise terms the purpose of this discussion is to mention some of the general problems that may be encountered and actions to mitigate those problems work force and production design when planning the details of a mining work force it is necessary to consider several factors is the supply of labor adequate to sustain the production level dictated by other economic factors if not can the needed labor be brought in and at what cost what is the past history of labor relations in the area are the workers accustomed to a 5 day work schedule and if so how will they react to a staggered 6 or 7 day schedule are the local people trained in similar production operations or must they be trained before production can achieve full capacity will a camp have to be built and the workers transported in on weekly schedules if long commutes are required to reach the property daily are company contracted busses for all shifts an option if long commutes are required are shifts longer than the normal 8 hours an option can people with maintenance skills be attracted to the property or will the maintenance crew have to be built up via an apprenticeship program apprenticeship programs are very slow in terms of producing results accordingly some state laws in the united states restrict the annual number of people who can be trained in such programs this one item could cause a mine designed and equipped for a very large daily production to fall far short of its desired goals work force issues and local community resources need to be

investigated at the same time that the property is being evaluated and designed this will provide adequate time for specialized training minimize unexpected costs and also prevent economic projections based on policies that if implemented could negatively impact employee morale or community relations the productivity and profitability difference between an operation with high morale and good labor relations and an operation where such parameters are poorly rated can be drastic such matters can make the difference between profit or loss of all the items involved in mine design this one is the most neglected and can be the most disastrous equipment selection field tested equipment the equipment selected should be produced by manufacturers that field test their equipment for long periods of time before they are marketed to the industry too many manufacturers build a prototype machine and install it in a customer s mine on the contingency that they will stand behind it and make it work properly eventually after both the end user and the manufacturer redesign rebuild reinforce and retrofit this model a workable machine is obtained the cost in lost production however is imposed on the mine operator not the manufacturer the manufacturer can then proceed to sell the field tested retrofitted model to the entire industry including competitive mines beware of taking on equipment in underground mines that has not been thoroughly field tested in severe applications by the manufacturer unless it is going to be used for an isolated research application equipment versatility the equipment selected for the mining operation should be as versatile as possible normally for large surface mines this is not much of a problem for small quarries and for most underground mines however it can be a problem if the equipment can only be employed on a limited number of mining operations for example in one room and pillar mining operation bullock 1973 the same high performance rotary percussion drilling machines used for drilling the bluff or brow headings were then mounted on standard drill jumbos for drilling holes for burn cut drifting and stoping rounds and for slabbing rounds in the breast headings because the drills on these jumbos have high penetration rates they were also used to drill the holes for roof bolts and in some cases the holes for the reinforcing pillars the same front end loader was used to load trucks in one stope and to perform as a load haul dump unit in another stope by switching working platforms the same forklift tractors served as explosive charging vehicles and as utility service units for handling air water and power lines they also served as standard forklifts for handling mine supplies this equipment philosophy results in the following advantages less equipment must be purchased and maintained less training is required for operators and maintenance personnel in addition all personnel have

a better chance of becoming more efficient at their jobs having fewer machinery types means having a smaller equipment inventory possible disadvantages of this approach are the following a more efficient machine may be available to do the task currently being performed by the versatile but less efficient machine the mine may become too dependent on a single manufacturer to supply its equipment needs equipment acceptance equipment selected should have a very broad acceptance and must be in common use throughout both the mining and construction industries however since underground mines impose a headroom restriction not encountered on the surface this is not always possible nevertheless where headroom is not an issue selecting a standard piece of equipment means that the components will have endured rigorous testing by the construction industry furthermore equipment parts are normally off the shelf items in distributors warehouses application flexibility the selected equipment should be flexible in application for example the equipment should be able to accelerate and move rapidly have good balance and control at high speeds be very maneuverable and have plenty of reserve power for severe applications both trucks and loaders should have ample power to climb every grade in the mine and be able to accelerate quickly to top speed on long straight hauls planning the underground mine service facilities underground facilities such as underground pumping stations power transfer and transformer stations underground shops storage warehouse space ore storage pockets skip loading stations lunchrooms refuge chambers shift and maintenance foremen s offices and central controlled computer installations and communications all extract a strong influence on engineering study and design more detailed information appears in chapter 12 5 where these types of facilities are described and their designed size is discussed whereas underground equipment should be selected to suit the mine it is equally true that the mine should be designed to suit the equipment it is impossible to begin mine planning and scheduling without a good idea of the type and size of equipment available and how it might be applied design of stopes or panels mine accesses development cross sections ramp gradients electric power reticulation ventilation circuits and so on must be based on an assumed type and fleet of equipment general approach in practice equipment selection for a new hard rock mine starts with the stoping operation stoping equipment should be appropriately sized for the characteristics of the ore body and the stope dimensions stope development openings should be sized to suit the stope dimensions and the stoping equipment capabilities and requirements equipment should then be sized to mine the required size of stope development separate development fleets for access and stope development may be required if this logic is not ap

plied big efficient development equipment may be acquired for access development and then used in the stoping areas the result is that the stope development is sized to fit jumbo drills rock bolter booms or even the haul trucks and must be far larger than the optimum causing inefficiencies in stope design poor geotechnical conditions or excessive overbreak and development cost of course it is often necessary to select new equipment for an existing mine where the constraints including access dimensions ventilation and so forth have to be identified and managed in a shaft mine for example the machine size may be limited by the dimensions of shaft compartments and the component sizes in which the machine can be broken down for transport proposed or possible changes to the mine plan such as a move from a 5 day working week to continuous working or a change in the decline gradient will affect the equipment duty and must be considered sometimes a new machine may allow a new way of working for example replacement of rear wheel drive trucks with four wheel drive trucks would allow ramps to be steepened this could reduce overall costs and it is this larger context that must be analyzed machines should be compared in terms of capital cost operating cost specifications performance availability ease of maintenance service and parts backup and service life fleet description the first step is to describe the key elements of the equipment fleet for example in a hard rock ramp access mine this might include two boom jumbo drills one boom jumbo drills rock bolting machines long hole drills large load haul dump units lhds small lhds haulage trucks integrated tool carriers and road graders many other machines such as personnel transporters shotcrete machines air compressors raise drills and others may be needed but the items in the list will be considered in determining stope and development layouts and dimensions equipment matching individual items of equipment considered together form a development system or a production system and thus they must be physically compatible jumbo drills must be capable of reaching the limits of the largest excavation required usually determined for the production drills and the trucks lhds must be able to reach and fully load the tubs boxes or trays of the trucks in three to five passes sometimes a workaround will enable incompatible units to be used for example a step in the floor may enable small lhds to load large trucks electric machines should all operate from the same electric supply voltage a decision must be made about whether compressed air will be reticulated and if not each pneumatic machine must be self sufficient for compressed air when items are obtained from a single manufacturer the manufacturer will be able to determine how well matched they are if they come from different manufacturers some research may be needed includ

ing detailed study of the specifications and discussions with staff of other mines about using the units together choice of manufacturer or supplier manufacturers of underground equipment range from large international firms with worldwide representation to small manufacturers that generally serve a local market or a specialized niche in general the international firms offer robust reliable and well proven equipment at the upper end of the price range equipment is usually priced in euros or u s dollars which may appear prohibitive when converted to a local currency small regional manufacturers can offer attractive pricing and may offer a viable alternative to the name brands because their level of research and development is limited and because they are building to a price machines from such manufacturers generally have lower availability and may have reduced performance these factors should be assessed carefully in deciding whether lower capital cost outweighs higher operating costs including the impact of servicing and breakdown delays on the mining system as a whole other manufacturers offer specialist equipment that is not available from the major suppliers examples of the products of international niche manufacturers include narrow vein mining equipment from france and canada rock bolting platforms from canada backfill slinger trucks from germany raise climbers from sweden and so on equipment with similar specifications from the major manufacturers will generally give similar performance this is often compared with the choice of buying a car from ford or from general motors both will do the job it is worth enquiring at other mine sites that use the same equipment because specific models sometimes have problems that are not really resolved during the model run whereas competing machines are free of those problems apart from operating performance equipment analysis should include the ease of maintenance some equipment models are easier to maintain which can have a positive impact on the availability of the equipment the most important factor in choosing between otherwise similar machines is the level of backup support provided in a region by the manufacturer appropriate questions include the following what training will the manufacturer offer to operators and maintenance personnel does the manufacturer have service facilities with trained personnel within an hour or two of the mine by road if not is it willing to establish such facilities on the basis of an order will the manufacturer stock a comprehensive range of spare parts in country for any specific model what is the likely delivery time will the manufacturer provide spare parts in the mine store on a consignment basis so that payment is made only when they are used which major components are not stocked in country how long would it take to obtain them if needed and what would be the cost of urgent air freight will the manufacturer

offer maintenance and consumables on a fixed cost per hour or per meter basis and establish its own maintenance work force on site what is the reputation of the manufacturer for service and spares backup in the region are components interchangeable between different pieces of equipment such as the same axles in lhds and trucks are the driveline components from the same manufacturer after these questions have been addressed it may be advantageous to source the key fleet items such as trucks lhds and jumbo drills from a common supplier it is also attractive to minimize the number of different models of equipment on the mine site even if from the same manufacturer to minimize duplication of stores inventories or maintenance skill sets mixed manufacturer fleets especially trucks may result in unexpected operational conflicts such as different truck speeds with slower trucks delaying faster trucks especially when hauling loads up a ramp appropriate technology any new machine should incorporate a level of technology that can be understood and supported by the operators and maintenance team although it may be appealing to engineers to acquire the latest technology the potential performance will not be delivered if the machine spends much of its time waiting for repairs at the crudest level there are mines that operate compressed air rock drills and do not have the workshop facilities or the skills to maintain hydraulic rock drills similarly many mines do not have the electronics capability to maintain the latest fully computerized jumbo drills and prefer to buy the versions with manual controls in general the longer the supply lines to the mine and the less educated or skilled the work force the more robust and less innovative should be the technology it is also difficult to mix different technologies within one operation for example a mine developed some years ago in southern africa tried to combine handheld drills and slushers in the stopes with rubber tired jumbo and truck development and haulage the two systems were not compatible large numbers of walking stope miners impeded the movement of trucks the rock broken in development was too large for the conveyor feeders and the maintenance workshops had to service large numbers of stope drills and scraper winches with which they were familiar and smaller numbers of hydraulic jumbos and diesel trucks which were new to them the mine failed for these and unrelated reasons mine systems should be designed to accommodate the technology to be used for example if a conventional mine is to be mechanized not only must the required infrastructure be provided but materials handling logistics need to be thoroughly planned and assessed for example this could include transportation and storage underground of diesel spare parts and hydraulic oils innovative models innovative new models of machines even from the most reputable manufacturers may present operationa

l problems it often takes 2 or 3 years of operation to resolve design problems in trucks loaders and jumbo drills with trucks giving the greatest problems usually the machine gets a bad name in the industry and a revised model with a new code name is released after the problems are resolved the problems are not predictable and could be as simple as shearing of wheel studs on a high powered truck the promised performance of a new model may be appealing and some mine has to be the first to try it when considering such a machine it is usually possible to negotiate a performance guarantee with the supplier this should take the form of an operating availability guarantee possibly with a minimum performance expressed as metric ton kilometers per hour meters drilled per hour or similar financial compensation applies if the machine does not deliver the guaranteed performance it is usual for the manufacturer and supply agent to be very interested in the success of a new product and they may provide high standards of operator training and assume responsibility for maintenance for an extended period the duration of these above normal levels of service should be ascertained examples of new equipment that have proved disappointing in the past include some high powered 50 80 t metric ton capacity trucks some combined drilling and cable bolting machines some rock bolting jumbos many models of scaling machines undersized impact breakers and novel surface or soft rock crushers and sizers when applied underground in hard rock mines equipment replacement all equipment has a service life after which it becomes uneconomical and should be replaced for some items the service life may be equal to or exceed the mine life in which case replacement is not necessary generally major items such as headframes winders hoists crushers ball mills and so on are designed to last the life of the mine and in practice may outlast several mines worldwide a number of operations are still getting satisfactory performance from crushers and winders that were built in the 1930s most portable and mobile equipment has a finite service life typically new mobile equipment will operate at low maintenance costs for approximately 10 000 hours or 2 years after which the cost increases steadily if the use of the machine is lower than around 5 000 h yr this honeymoon period may be extended this cycle may be repeated following one or two major rebuilds although the subsequent maintenance costs are rarely as low as for a new machine if the working environment changes a machine may be deemed obsolescent and need to be replaced examples include a change to the stoping method requiring a larger machine or a change in the ventilation standards that requires a cleaner engine if there is a substantial improvement in the available technology a machine may be replaced even though it is still delivering to its original specification for

example new rear dump haul trucks with much improved power toweight ratios may displace older slower units a decision to replace an existing machine must consider the entire physical and financial environment for example in a struggling mining company capital may simply be unavailable no matter how well a replacement can be justified in such circumstances the equipment life can be extended almost indefinitely but with a much increased operating cost to extend useful life many operations cycle their equipment through two or more duties for example haul trucks may be converted to water trucks or production lhds become development lhds where they are used more intermittently a particular problem arises in shaft access underground mines the difficulty involved in dismantling and reassembling large mobile units for shaft transfer leads to inflexibility in maintenance and major overhaul periods may be extended operators tend to retain old equipment as backup units long after they have reached the end of their economic life this creates an environment of low usage and excessive maintenance costs for example one australian shaft mine was able to reduce its mobile equipment fleet from 50 to 11 units after establishing a service decline from the surface service life and maintenance costs also depend on a company s maintenance policy and the level of onboard monitoring installed making financial comparisons the benefit of investment in a new machine should be quantified in advance and in most companies this is a requirement of the capital approval process the economic benefit is best measured by comparing the net present values npvs or usually net present costs of the alternatives if the expected service lives of the alternatives are different the analysis becomes complicated for example two machines with service lives of 5 and 7 years would need to be compared artificially over a period of 35 years to get a meaningful comparison or an accurate estimate be made of the residual value of the longer life machine after 5 years to overcome this problem costs should be annualized using an approach known as the annual average cost or the equivalent annual value method equipment replacement decisions are not exclusive to mining but apply throughout all business activities detailed discussions of the methods can readily be found in textbooks on engineering economics and on the internet and suitable financial modeling software can be downloaded and purchased for a small sum many mining companies require that the internal rate of return for the investment exceed a hurdle rate a minimum acceptable rate of return that is set from time to time by the administration and is based on the company s cost of capital or desired rate of return on all of its investments although this may simplify decision making it often leads to suboptimal decisions because it prohibits many investments that add some value to an operation and in par

ticular it ignores the many benefits that cannot be quantified in a simple financial analysis companies may also limit the timing of capital approvals within the annual budget cycle and take a dim view of out of budget applications nonfinancial measures of performance may be important for example improvements in workplace safety or environmental quality may not be economically measurable but may be highly desirable provision of a new machine with an airconditioned cab may improve industrial relations even if the machine specifications suggest there will be no increase in productivity at many mines particularly at remote sites there is a large benefit in reducing labor numbers increased mechanization will aid this and can become a matter of policy even if it is difficult to quantify the gains from a particular decision on an npv basis labor costs are typically 50 of total costs but when the cost of extras such as airfares and site accommodation are included they may be higher yet conversely the social political climate in some areas may dictate that reduction in labor by introduction of more productive machines is politically undesirable an example is the large amount of hand labor on ramp roadway maintenance seen in many mexican mines when a grader would be more efficient information required for financial analysis the capital cost of a new machine should be estimated including the cost of spares which can be 20 of the purchase price for a mobile unit this will depend on whether spares are available on consignment that is they can be held in the store and paid for when used estimators should use the final or likely negotiated price rather than the manufacturer s list price which is usually somewhat higher if an item is sourced overseas estimators should check what currency will be used for the transaction and what protection if any is available against exchange rate variations any payable import duties and the transport and unloading costs including in country transport to the mine should be checked for a shaft mine the transport cost will include shaft transfer and reassembly which can be substantial for bigger items operating costs can be estimated from detailed cost records and performance of similar equipment cost records for the new item from another operation manufacturer s cost and performance estimates and a database of costs from a range of operations to support ongoing decision making good record keeping and cost allocation are required over the life of each unit the mine accounting system must be able to track each piece of major equipment and not report a combined cost for all lhds trucks jumbos and so on records should include engine hours and for rock drilling equipment percussion or hydraulic pump hours as well as outputs such as metric tons loaded metric ton kilometers hauled or meters drilled caution must be used when comparing costs from different operations where di

fferent operating conditions may apply and the basis of reported costs may not be well understood engineers should be particularly wary of manufacturer s estimates as they are inevitably optimistic this is an area where consultants can be very helpful as a good consultant should have an extensive cost and performance database the salvage value may sometimes be important for mobile equipment it will depend on the age of the unit its condition and the operating hours since a major rebuild a rebuilt 10 000 hour underground mobile item could be worth 60 70 of new cost the rebuild itself might cost 20 30 of the original purchase price an older unit will generally be saleable for about 20 of new cost if it is operable machinery merchants are generally willing to provide an estimate of value for a used item and for common items may not require an inspection the cost of removing the old item from the mine particularly if it is a shaft mine may negate the salvage value in which case the unit should be pushed into a stope and backfilled taxation effects and the method of financing the new item may affect the analysis leasing and contracting besides outright purchase other forms of equipment finance are available an operating lease is similar to rental is usually for shortterm requirements and can be cancelled without penalty it is relatively expensive because the lessor assumes the risk of re leasing and technological obsolescence but is fully deductible as an operating expense it does not affect the lessee s balance sheet a finance lease is a contract whereby payments are made over much of the useful life of the asset during this time the lease cannot be cancelled although it may be paid out if the item is disposed of or destroyed the cost includes capital less salvage cost interest a risk premium and the commercial cost of providing the service typically 60 70 of the purchase price is financed with the balance payable as a residual the item appears on the lessee s balance sheet as an alternative to leasing a contractor and contractor s equipment could be used applications include for example truck haulage production drilling surface crushing and the entire mine development or production program contracting is often appropriate where a small additional increment of capacity is required for 1 or 2 years it also makes sense where specialist skills such as cable bolting are required for a limited time or where the necessary expertise is not available in the vicinity of the mine and there is insufficient time to train a work force estimating the average annual cost the average annual cost is the sum of depreciation interest and operating cost consider a truck that has an initial cost of 1 0 million and a salvage value of 0 2 million after an estimated 4 year service life the depreciation is thus 0 8 million over 4 years or 0 2 million yr the interest expense arises because owning an a

sset ties up capital this is true whether all or part of the purchase price is borrowed or existing cash is used in the latter case there is an opportunity cost because that money could have been invested elsewhere the average investment is 1 0 million 0 2 million 2 0 6 million if the company has a cost of capital or cost of borrowing of 10 yr the annual interest expense is 0 1 0 6 million or 60 000 for each of the four years the annual operating cost may be calculated from the estimated hourly operating cost of 200 h which includes operator fuel tires maintenance etc and the cost of a major rebuild at 12 000 hours and the expected use of 5 500 h yr which is a total of 1 1 million yr from this the average annual cost is 0 2 million 0 06 million 1 1 million 1 36 million a similar calculation can be made for each of the trucks under consideration the equipment selection decision can then be based on these results together with all of the other important considerations outlined in this chapter hard rock equipment selection the key items of mobile equipment are loaders trucks jumbo drills and production drills each requiring a different approach to selection specific selection criteria apply to each item of ancillary equipment in all cases the considerations discussed in the previous sections should be addressed selecting a truck trucks used underground fall into three categories rigid body rear dump trucks adapted from surface mining units articulated rear dump trucks that may be adapted from surface mining units but usually are purpose designed to have a low profile figure 12 2 1 and tractor trailer units some with separately powered trailers usually side dumping which can be assembled as an underground road train figure 12 2 2 all have diesel engines except for trolley electric trucks which require special infrastructure where trucks are required to travel through development openings the truck cross section will fix the dimensions of the opening or vice versa allowance must be made for clearance from the load to any pipes ventilation duct or other services and clearance to side walls which may be a statutory requirement a typical design profile is shown in figure 12 2 3 the clearance to ventilation ducts is influenced by the following the shape of the tunnel profile whether flat arched or shanty backed which is determined from the geology and ground support requirements the duct diameter which depends on the airflow required sometimes oval shaped ducts are used to improve clearances but at the cost of greater resistance and fan power consumption the type of duct whether rigid or flexible flexible ducts may be more mobile and require greater clearances the size number and position of ducts within the cross section the position could be in the crown or in one corner of the profile in wider headings vent ducts could also be along a sidewall the s

ize and number of ducts depends on the planned distance of forced ventilation before connecting to a return airway and the ventilation volume required to service the heading a forced ventilation heading may typically require one loader and one truck the ventilation volume is usually a relationship based on a factor relative to the engine power duct diameter is then related to required ventilation volume and duct resistance duct diameter may become quite large 1 2 m diameter for significant forced ventilation lengths which may require duplication or a mix of types forced and push pull systems the expected load profile in the truck sometimes trucks are not fully loaded during the development phase but can be fully loaded after the development is completed when a ventilation circuit is established and the ventilation ducts have been removed the initial thickness of road base material and any subsequent buildup of road base dressing material that is not removed by grading and allowance for road spillage that a truck may accidentally ride over the layout of turnouts from the main drive and the way ventilation ducts make the transition through corners sidewall clearance is required to minimize impact damage and for safety reasons there must be room for pedestrians to walk past a stationary truck and for recovery work to be done in the event of a breakdown or tire failure it is not acceptable for trucks to pass pedestrians trucks require a minimum turning radius which generally will not be 20 m due to truck requirements capability and maintenance and development issues on grade overbreak on development rounds widening of the heading and or flattening the gradient may be necessary on bends to maintain required clearances or reduce longitudinal torsion of the truck frame road trains require additional attention to radius and wall clearances due to the variable tracking path of the powered trailers development loops are also required for the road trains as they are not designed for reversing and tipping is usually sideways the operating cost per metric ton kilometer decreases as the truck capacity increases thus there is an incentive to maximize truck capacity with diesel trucks of up to 60 t capacity being available for underground applications a trade off study is required to assess the benefit of the low truck operating cost against the disadvantages that come with size factors to be considered include the following the volume of waste rock broken increases with the cross sectional area of the heading for example a 5 5 5 5 m drive will produce 53 more waste rock than a 4 5 4 0 m drive if the cost of development breaking and disposal is 130 m3 and 10 km of primary development is planned the cost difference of more than 1 2 million must be recovered from a reduced trucking cost for all of the ore and waste rock from the mine ground support costs generally increase with the square of span lar

ger more powerful trucks require more ventilation per truck but not per ton moved a larger mine opening with larger trucks will have a greater total production capacity in metric tons per day or per year equipment manufacturers provide comprehensive manuals of specifications of their products and many provide detailed approaches to equipment selection the key considerations for an underground truck are physical dimensions and load capacity exhaust gas and exhaust particulate quality safety features including falling object protection system and rollover protection system and whether these comply with local standards air conditioned enclosed cabin ergonomic cabin and seat power to weight ratio and hence speed on grade engine braking capability or hydraulic retarder whether fitted with spring activated hydraulic release fail safe brakes whether the torque converter locks up to provide better power transfer and longer transmission life centralized lubrication and centralized fire protection system the truck load can be calculated from the truck body volume the broken density of the rock to be carried and a fill factor that typically varies from 85 to 95 the stated tonnage capacity of the truck is typically quoted as sae heaped where sae refers to the society of automotive engineers which has the top surface of the load rising one length unit vertically for every two length units horizontally from the top edge of all sides of the tray this is rarely achieved in practice particularly for up ramp haulage it is not unusual to find that a 50 t capacity truck achieves an average payload of 43 t particularly if there is any problem with back clearance on the haul route capacity is maximized by using a well matched loader and with the use of an onboard weightometer the manufacturer s tables will provide torque speed curves speed on grade curves and other information to allow cycle times to be estimated in general it is sufficient to determine the cycle as the sum of loading time including any wait for loader which will depend on the loader capacity and the distance from the muck pile to the loading point travel time at an average speed on grade from the loading point to the dump point usually including level and uphill segments dump time including any delay waiting for other trucks to dump and travel time at an average speed on grade back to the loading point the average speed when loaded will depend on the condition of the roadway including the presence of spillage or water the clearance to the tunnel walls the road gradient the number of curves or bends and any mine imposed or statutory speed limit the empty travel speed if downhill may be limited for safety reasons to less than the general mine speed limit similarly there may be load or gradient limits for loaded travel downhill traffic congestion occurs when the number of trucks required is high relative to

the length of haul so efforts should be made to provide effectives traffic management through the use of radio communications block lights and well spaced efficient passing bays the problem is exacerbated for mixedfleet systems different capacity different loaded speeds the difference between the efficiency factor introduced within the cycle time calculation and the job or efficiency factor discussed here is that the former increases the operating engine hours required and hence the annual operating cost whereas the latter reduces the operating engine hours available a useful check is to see that the annual hours required from the truck are consistent with experience at similar sites there can be a large difference in annual operating hours for underground haulage trucks ranging from around 2 000 hours up to 5 600 hours depending on the management strategy and mine layout because haulage routes vary it is often useful to compare truck performance on a metric ton kilometer basis this assumes that hauling 1 t for 10 km requires about the same truck capacity as hauling 10 t for 1 km in the previous example a 50 t truck hauled 43 7 t h over a distance of 3 5 km an output of 153 t km h actual truck performance as recorded from 32 operating mines is shown in figures 12 2 4 and 12 2 5 the metric ton kilometer approach should be used cautiously because it ignores roadway gradient which can vary substantially whereas some costs such as tires vary with load and distance traveled others such as fuel are more closely correlated with engine hours and engine power setting when calculating the number of trucks required allowance must be made for the haulage and disposal of waste rock from development often this rock will be backfilled into stopes requiring a shorter cycle time than a haul to the surface however backfilling often causes delays while waiting for a loader or bulldozer to clear the dump point so that the advantage may not be great where trucks must dump fill rock into stopes under a limited back height ejector tray trucks may be used these have an ejector blade something like a bulldozer blade which pushes the load from the truck body without having to elevate it into the tipping position ejector tray trucks have a reduced payload compared to the equivalent standard model and because of the increased weight in the truck body they are less fuel efficient they also have more moving parts to maintain and can suffer damage if large rocks wedge between the sidewall and the ejector trolley wire electric trucks have been available for many years but have limited application because the system has a high capital cost when the trolley wire installation is considered although there is no diesel exhaust they require substantial ventilation for cooling if the mine generates its own power from diesel on the surface there is no operating cost saving a very high standard of road maintenance is required

and without auxiliary power trucks cannot travel beyond the fixed trolley wire into actively developing areas despite these limitations trolley wire trucks may offer advantages for a mine using low tariff grid power with a fixed loading point such as the bottom of an orepass system selecting a load haul dump machine lhds are known variously as loaders muckers or boggers and differ from front end loaders in that they have a lower profile and are designed to carry a load in the bucket efficiently for up to 200 m between the muck pile and the truck or orepass that they are loading into thus an lhd spends more time traveling than does a front end loader and is engineered accordingly a mine may use different sized lhds for stoping and for development often two or more sizes of lhd are used for stoping to suit the ore body width and mining method at different locations within the mine standard models are available in bucket capacities ranging from around 3 to 11 6 m3 special machines designed for narrow vein mining can be as small as 0 5 m3 corresponding payloads range from 1 to 25 t extremely low profile machines are also available for working in flat dipping narrow vein ore bodies with as little as 1 6 m of headroom lhds may be diesel or electric powered the diesel units are versatile and can tram quickly from one location to another figure 12 2 6 the electric units figure 12 2 7 carry a cable drum and rely on a trailing electric cable so they are tethered to a location during normal production electric lhds have low noise levels and zero emissions so they require less ventilation they are highly productive in situations such as block caving where ore is transported from a series of drawpoints to a fixed orepass location the ergonomic and safety considerations for selecting an lhd are similar to those for trucks as listed previously in the past some operators sustained back injuries due to poor operator location and poor seating the best operating position is midway between the front and rear wheels and particular attention should be paid to the quality of the seat suspension lhds often operate in situations with limited forced ventilation so an enclosed air conditioned cab is desirable and the engine should be the latest generation low emission type fail safe brakes are highly desirable lhds can operate effectively in headings as steep as 20 1 in 5 although they perform best if the gradient is limited to around 14 1 in 7 or less the size of a selected lhd must fit within the planned development and stope openings and the bucket must be able to reach above and fill a truck efficiently the bucket load can be calculated from the bucket volume the broken density of the rock to be carried and a fill factor which depends on the fragmentation of the rock and may be different for development rock and for broken ore in stope drawpoints as with trucks the nominal capacity quoted is typically for sae

heaped loading which is also typically difficult to achieve in practice however a high fill factor can be achieved with well blasted rock for example an 8 m3 bucket with a 90 fill factor and broken ore of bulk density 2 2 t m would carry 15 5 t the manufacturer s tables will provide speed on grade and other information to allow cycle times to be estimated some manufacturers provide estimates of productivity in metric tons per hour for various haulage distances in general it is sufficient to determine the lhd cycle as the sum of the following loading time which depends on the power of the lhd the muck pile condition the floor condition and fragmentation of the broken rock the presence of large rocks which require individual attention can slow the loading cycle considerably travel time at an average speed on grade from the loading point to the dump point dump time including any delay waiting for trucks which is usually faster when tipping into a pass because no attention to the shape of the load is required however depending on the fragmentation of the rock and the size of the pass grizzlies or other methods of size restriction to limit the size of rocks in the pass may constrain the tipping rate into a pass travel time at an average speed on grade back to the loading point remote control and automation of lhds and trucks remote control of loading units is desirable where conditions at the face or in the stope are hazardous or where the ground overhead has not been secured radio remote control of lhds has been available for many years having evolved from cable remote control of compressed air powered loading units line of sight remote control has proven to be hazardous and has fallen out of favor despite training and strict operating rules operators tend to move into the path of the loading unit and hence are exposed to injury teleremote control using television cameras mounted on the lhd allows the operator to be seated in a safe airconditioned environment underground or even to operate the machine from the surface current developments include employing the operator in a city office remote from the mine thus making the job more attractive while greatly reducing onsite costs the productivity per hour of an lhd is typically less under teleremote control than with an operator on board because the teleremote operator does not have the full range of audible visual and physical inputs to enable optimum performance however most teleremote operations currently use two dimensional video screens for providing visual information to operators such things as three dimensional 3 d visual presentation and virtual reality helmets available in other industries have been trialed in mining operations but have not come into widespread use it is conceivable that in the future provision of real time 3 d virtual reality visual and audio information to remote operators may reduce the difference in productiv

ity between remote control and direct operation also teleremote operation offers the possibility of operating for a greater part of the working shift because delays such as travel time do not apply and because operators can be rotated in a surface office to keep the machines at maximum use the major manufacturers now offer autonomous lhds that can operate with minimal or no human intervention these machines have sophisticated onboard computers and typically use lasers to scan their environment and determine position such a machine can load tram and dump autonomously so that a human supervisor can manage two or more lhds the loading part of the cycle requires the muck pile to be well broken the current generation of autonomous machines is not as effective as humans at assessing the position of large rocks and maneuvering the bucket to displace and lift them the lhd can be damaged if used autonomously with large rocks so a common compromise is to load the lhd on teleremote and then switch to autonomous for the remainder of the cycle the use of surface teleremote and autonomous lhds requires a good broadband communications backbone to be installed in the mine and a high standard of training for maintenance personnel for safety reasons these machines can only be used in an area that has been locked off from human entry so the opportunities for their use are limited they are suited to block caving and some sublevel caving operations where production activities can proceed for a significant time without the need for other activities around them in many other mines activities such as development geological sampling surveying and so on interact with production loading so that it is impractical to create a nonentry zone for autonomous operations the mine must be designed to accommodate the level of technology where automation has been adapted to an existing mine layout the results have generally been unsatisfactory design of operations needs to include interaction between activities dealing with spillage drainage and road maintenance blast damage planned and unplanned maintenance activities surveying geological mapping and sampling and so on autonomous haulage trucks are also available but are not yet widely used autonomous trucks must operate on a dedicated haulage route that is locked off from other traffic and from pedestrian traffic which can rarely be arranged in existing mine layouts in most cases duplicate development of a dedicated haulage roadway would be required and may not be justified selecting a jumbo drill a modern jumbo drill has an articulated diesel powered carrier and an electrically powered hydraulic pump which operates the boom rams and hydraulic rock drill s when the machine is in drilling mode the jumbo may have one two or three booms depending on the size of openings the diesel engine enables the jumbo to tram between working places it is rare for the diesel engine to provide hydraul

ic power for drilling because it is inefficient and because there is usually insufficient ventilation to operate the engine at the face some models have a fast tramming option that is useful in a large mine where the work places are far apart single boom jumbos are suited to face areas from 6 to 30 m2 whereas twin booms may fit in an opening as small as 8 m2 but can extend 100 m2 or more two booms are sufficient to cover the face area in most mines with three boom jumbos being more popular in civil engineering works where a three boom jumbo would in theory provide a faster drilling cycle in practice the operator is fully busy managing two a valuable application of the center boom is drilling the large diameter reliever holes in the cut many jumbos have a third boom with a basket a single boom is used in small headings or narrow veins where there is insufficient room to operate a two boom machine figure 12 2 8 the manufacturer s data will show the capability of the jumbo to cover the proposed face area the blast design will determine the hole diameter generally in the range of 33 to 89 mm although larger holes can be drilled at greatly reduced penetration rates depending on the duty rock drills drifters with impact power in the range of 10 to 30 kw are available the length of round to be drilled and fired will determine the length of the drill steel and the boom in general longer heavier booms require a larger carrier it is important that the jumbo has adequate reach to enable back holes to be drilled parallel with the axis of the excavation if an undersized machine is used to drill large profile excavations by angling the back holes out not parallel to the direction of advance the result may be a sawtooth profile excessive overbreak and unsatisfactory conditions for ground control high powered drifters may require a high voltage system 1 000 v to deliver sufficient energy to the hydraulic pumps the voltage also impacts the trailing cable diameter and hence trailing cable length and power extension requirements this is of relevance for countries where lower voltage systems 450 v are the norm an important consideration is whether the jumbo will also be used for rock bolting and meshing either consistently or occasionally and the type of rock bolts to be used for rock bolting it must be possible to turn the booms at 90 to the axis of the carrier and to drill holes into the walls and back the boom length cannot exceed the available width and height in these situations a split feed boom may be required which effectively telescopes the boom to full length as required this adds complexity maintenance and weight and may lead to inaccuracy as parts wear so the benefit must be weighed against these considerations a jumbo may also be used for scaling or rattling the tunnel walls and back as part of the development cycle in lieu of barring down or using a dedicated scaling machine this is ne

ver recommended by the drill manufacturers because it can cause a lot of damage to the booms from falling rock nevertheless it is a common practice as operators have learned that the convenience and safety benefits can outweigh the increased maintenance for this application a robust and well protected boom should be selected the face advance that can be achieved from a jumbo drill depends on the number of faces available to it in a singleheading situation such as during the development of a main decline from the surface a twin boom jumbo can typically average 45 50 m of advance per week including the main heading and stockpile bays this means that the main face will advance on average about 40 m week this rate can be exceeded for periods of several months if in good ground but most mines present a range of conditions including zones of poor ground or faults which slows progress so that over a 2 year development program the 40 m week rate would apply a higher advance rate can be achieved using the latest long round technology for example during the development of newcrest mining s cadia east project in new south wales a record for 6 0 m 5 5 m mine development with full ground support was set with an advance of 283 m in a single month including stockpile bay development willcox 2008 long round development has limitations on curve radius and turnout or crosscut development so is not generally applicable to underground mine development higher development rates exceeding 400 m month have been achieved in civil tunneling works but the cadia east performance is a good example for mine development the process of estimating the performance of a two boom jumbo drill is set out in tables 12 2 1 through 12 2 3 which are based on actual measurements from a mine where each jumbo had at least five faces available to it for drilling and so could be kept continuously at work causes of delays were recorded in detail on shift reports the jumbo drills were conventional units without computer alignment and were not the latest generation of high frequency drifters after this analysis is completed using performance parameters for the jumbo drill under consideration the required number of jumbos can be estimated if the target is 800 m of development per month then in this example three jumbos would be required of course if sufficient faces cannot be made available to fully use the drills this problem is not solved by adding more drilling capacity instead the materials handling capacity must be addressed selecting a rock bolting machine rock bolting may be done using a handheld rock drill from the muck pile a work basket or a scissor lift platform purposebuilt articulated four wheel drive scissor lifts are available with or without a bolting boom attached alternatively the development jumbo drill can be used to install rock bolts and mesh with minimal exposure of the operator to the area being supported purpose de

signed machines are available for drilling holes and installing both conventional rock bolts and cable bolts a rock bolting machine resembles a jumbo drill and may use the same carrier the working boom carries a rock drill and a carousel that can handle a range of rock bolt lengths and typically 10 rock bolts a separate hydraulic arm may be fitted to maneuver and place screen mesh a rock bolting machine may be capable of installing mechanical or friction anchored bolts as well as cement or resin grouted bolts machine selection should address the following points is a remote control option available for hazardous situations the rock bolting machine size should be matched to the required heading size which in turn depends on the purpose of the heading it is unfortunately quite common to find that limited availability of rock bolting boom lengths forces headings to be larger than they need to be for the purpose for which they were designed at best this might merely produce more development waste than necessary at worst it can make the use of the heading for its intended purpose more difficult more costly or more time consuming than it would have been if correctly sized rock bolting machines are subject to impact damage from falling rock and require tight tolerances on working parts so that the rock bolt can be indexed to the drilled hole it is useful to enquire from other operators about their experience with maintenance on the model being considered rock bolting machines tend to be slower and less mobile than jumbos a check should be made to ensure that the rockbolting machine can travel between development faces within the cycle time required selecting a production drill production drills are used for drilling stope blastholes and service holes top hammer drills are generally available in sizes from 5 to 127 mm in diameter whereas in the hole hammer ith drills range from 95 to 178 mm in diameter the drilland blast design will determine the optimum diameter and length of the hole the selected drill must be able to quickly drill the hole at minimum cost with acceptable deviation deviation will limit the smaller hole sizes to around 12 15 m in length whereas holes of around 100 mm diameter with tube drills can be up to 40 m long longer holes require larger diameters favoring ith drilling skid mounted crawler mounted and trolley mounted rigs have generally been replaced by highly mobile dieselpowered rigs top hammer rock drills are hydraulic using electric hydraulic power packs whereas ith drills use compressed air from a stand alone or booster compressor highpressure water powered hammers are also available for narrow vein mining production drills that can operate in vein widths of 3 0 m or slightly less are available in the larger sizes fully automated operation is possible so that a complete ring of blastholes can be drilled without human intervention considerations in selecting a production drill in

clude the performance including penetration rate of the drifter at the desired hole diameter whether the desired hole inclinations can be drilled parallel hole coverage capability feed pivoting system rigid with sliders more accurate holes or totally variable highly variable hole alignment whether reticulated compressed air or an onboard compressor will be used whether centralized lubrication is fitted whether automated rod handling is available whether carousel capacity is adequate for planned hole lengths whether remote control is available umbilical cord system for misfire drilling management what onboard surveying system is used for hole positioning whether cab is closed and air conditioned the level of automation required for increased use operator visibility of hole collar and lighting conditions and the ability to collar the hole close to the wall particularly in narrow vein applications drill string selection rods in hole stabilizers tubes and feed stabilization front and or rear feed stingers are also important considerations for managing drill hole deviation figure 12 2 9 shows production drilling rigs with hydraulic top hammer rock drills both are designed for long hole ring and parallel hole drilling upward or downward in figure 12 2 9a the drill unit faces the canopy and is suitable for holes in the range of 89 to 127 mm the drill in figure 12 2 9b is boom mounted for maximum reach and flexibility in both production drilling and bolt hole drilling and is suitable for holes in the range of 51 to 89 mm other underground equipment the considerations outlined in the opening sections of this chapter apply to a wide range of underground equipment specific considerations include the following integrated tool carriers are based on wheel loaders for greater load carrying capacity versatility and stability than conventional forklifts operator visibility and automated quick coupling are important features robust scalers are becoming more effective in headings between 4 and 6 m high backhoes fitted with a scaling pick or lightweight impact hammer and very large scalers have proved less successful some limestone mines use a rotary head on an excavator very effectively shotcrete or fibercrete machines may be designed for wet or dry mix material mobility boom reach and remote control capability are important in underground operations the machine should have a remote controlled boom nozzle to remove the operator from the unsupported area explosives charge up machines may be used for cartridge emulsion or ammonium nitrate and fuel oil and the charging process may be manual or mechanized as with jumbo drills and rock bolting machines the chargeup machine must be sufficiently mobile to service the faces required within the development and production cycle the reach and load capacity of the basket and the explosives carrying capacity must be suited to the propos

ed duty road maintenance equipment such as graders with or without rippers bobcats water trucks and rollers need to be fully assessed for the suitability of the road surface medium and dimensions of the underground workings a wide range of four wheel drive vehicles are suitable for underground mining although many of the lower cost models may not stand the rough conditions of a typical mine and may have a very high maintenance cost it is a good strategy to buy the best rugged model available secondhand with around 100 000 highway kilometers in corrosive or rough mine environments the best vehicle life may be around 2 years estimators should not underestimate the cost of fitting a vehicle for underground operation which might include crash bars engine isolator and removal of unwanted interior fittings handheld airleg or jackleg drills rope scrapers slushers and rail equipment may still be used in some circumstances although their use is rapidly diminishing proper selection of rail equipment in particular requires careful consideration based on cycle times gradients and locomotive capability it is best to consult old textbooks including earlier versions of this handbook to gain an appreciation of the selection process soft rock mining predominantly applies to coal mining but is also applicable to other bedded soft mineral deposits therefore although this chapter primarily focuses on underground coal mining certain aspects are relevant to mineral deposits such as potash underground coal mining methods currently fall into the following principal categories longwall full extraction operations room and pillar partial extraction operations room and pillar first workings operations these methods are discussed in the following sections of this chapter figure 12 3 1 provides examples of the three categories longwall mining methods because of its mechanized process longwall mining generally achieves the highest productivity and reserve recovery and is typically the primary choice for new mines being developed unless geological surface or capital constraints dictate otherwise in longwall mining a set of roadways is driven out from the main entries down each side of the longwall panel to block out a portion of reserves the mechanized longwall equipment is then retreated through the panel to extract the entire panel of coal the roads used for traveling and the roads used for coal clearance during panel extraction are termed the maingate or headgate entries whereas the tailgate entries are generally used for return air where possible a set of parallel adjacent panels is extracted sequentially to allow reuse of one gate road i e the maingate for the first panel becomes the tailgate for the next panel and so forth working in this way maximizes coal recovery while keeping the quantity of roadway drivage to a minimum the number of roadways that comprise a maingate varies around the world

with single entries being common in europe two roadways common in australia and china and three roadways common in the united states the number of roadways influences the style of coal clearance and the mobile equipment used longwall mining methods vary slightly according to the seam height which can range from as low as 1 0 m to in excess of 15 m the height influences the longwall equipment used single pass seam extraction is undertaken in seam heights of up to approximately 5 5 m for which conventional longwall equipment is used for seams in excess of 5 5 m either multislice or top coal caving methods can be used top coal caving requires a different style of roof support and an armored face conveyor afc configuration in very thin seams the longwall plough system can be used which uses a different coal cutting machine a plough as opposed to a shearer the longwall panel width can vary up to more than 400 m and this influences the longwall equipment quantity of roof supports quantity of afc pans hydraulic power electrical power etc panel length can be dictated by geology surface subsidence restrictions equipment limitations or other mining constraints very short panel widths in the order of 25 to 50 m are typically termed mini wall longwalls and these use a specific type of afc arrangement room and pillar mining methods until the advent of high production longwall techniques the majority of underground coal extraction used room and pillar r p also termed development or bord and pillar mining techniques in many parts of the world this is still the case with many mines achieving high production rates from r p methods the use of the r p method is often favored because of its low capital requirements compared to the longwall system the r p method is also used in small or irregular shaped deposits deposits with surface subsidence restrictions and deposits where geological constraints such as faulting preclude the economic use of longwall mining r p methods range from first workings where only roadways are driven forming pillars to second workings where the pillars are then later either fully or partially extracted r p methods are generally deployed in seam heights ranging from 1 5 to 4 5 m in seam heights of less than 2 1 m the roof is often excavated periodically or continually to provide an adequate working height low seam heights can restrict equipment options particularly bolting equipment with greater seam heights some mines extract the floor as a second working thus allowing up to 6 m seams to be partially extracted in these cases consideration is required for roadway gradients and equipment capabilities in regard to negotiating steep grades in first workings a benefit can be gained by increasing the roadway width to up to 7 5 m or more depending on the spanning ability of the roof in potash mines roadway spans typically exceed what can be achieved in coal with the roadways fo

rmed by wide head continuous miners cms taking several passes a cm in an r p operation is shown in figure 12 3 2 secondary extraction methods generally fall into pillarsplitting or stripping methods or fall into split and fender methods for example wongawilli extraction in australia figure 12 3 3 in the standard pillar splitting technique the main development roadways are driven toward the boundary of the mining area forming coal pillars to support the roof when the limit of reserves or the mining boundary is encountered the coal pillars can either be split or stripped thus forming smaller pillars that are then left to support the roof or that collapse in a controlled manner on retreat this method of mining requires specific equipment that can extract coal and retreat in a nimble and rapid manner split and fender mining involves the formation of main access roads into a particular district and the subsequent blocking out of an area of coal typically 60 to 80 m in width and up to 1 5 km in length a roadway or split is driven through to the end of the block using a cm leaving a thin coal fender that is extracted as the cm retreats back toward the access roadways allowing the roof to collapse behind when the fender is fully extracted the mining equipment is retracted and the entire sequence is repeated until the whole coal block is mined out although similar equipment can be used for this method the use of powered roof supports mobile breaker line supports is often used to increase safety and allow for a greater level of roof control longwall mining equipment this section begins with a discussion of standard longwall equipment followed by discussions of longwall equipment variations selection specifications and sizing standard longwall equipment the standard longwall face equipment consists of roof supports a shearer an afc a beam stage loader bsl including crusher and boot end a monorail a pump station and system electrics figure 12 3 4 is a schematic of a longwall system showing the relative locations of each piece of equipment roof supports roof supports sometimes referred to as shields or chocks hold the exposed roof as the coal is mined allowing the roof strata to cantilever over the support fulcrum and then break off into the goaf or gob in this manner the operators are shielded from falling goaf as the face is advanced and the immediate face area is destressed as the shearer cuts along the face the afc is snaked across the new face line using the double acting sequence rams fitted between the supports and the afc pans as the afc is snaked the roof supports are lowered and pulled across to the advanced position by the sequence ram and reset to the roof referred to as the lower advance set cycle in this way the face is automatically advanced as each longwall slice is taken through the panel the immediate roof falls behind the support forming a goaf roof supports can be operated i

n either conventional mode or in immediate forward support ifs mode conventional supports cannot be advanced to support the roof until the afc is advanced i e the roof is exposed and unsupported for some distance behind the shearer whereas in ifs mode the supports are advanced immediately after the shearer has passed ifs support is generally needed in weak roof conditions and is the most popular mode of support currently used ifs supports require longer canopies than conventional supports with corresponding larger leg cylinders and hydraulic systems modern longwall supports have been built to operate to approximately 5 8 m in height support capacity depends largely on support configurations although the largest supports available have a capacity of approximately 1 750 t metric tons equating to a support density of approximately 125 t m2 and weighing up to 65 t the roof supports used at the gate ends are configured slightly different to the run of face supports in that they have longer canopies so as to extend over the maingate drives in addition the gate end supports have higher rated capacity relay bars to enable pushing the maingate drive frames and bsl typically there are five gate end supports at the maingate and three at the tailgate originally operated by valves fitted to each support lower advance and raise modern supports are controlled by electrohydraulic systems whereby solenoids operate the valves and the supports can be operated automatically in sequence mode or from adjacent or nearby supports this automatic sequence also can operate with a positive set mode whereby each support is set against the roof to a minimum pressure often between 70 and 80 of the yield pressure thus guaranteeing effective support when in automatic mode when specifying roof supports various options are available including the following four leg versus two leg supports because of advancements in cylinder technology and capacity and because of lemniscate design two leg supports are now more popular front walkway rear walkway or both rear walkways are often used in thick seams to offer more protection against face spall side shields on one or both sides used to push supports uphill face sprags or flippers used to prevent face spall in thick seams base lift systems used to raise shield prior to advance when in boggy floor a longwall roof support with face sprag and base lift fitted is shown in figure 12 3 5 shearers modern longwall double ended ranging drum shearers have two cutting drums one at each end mounted on hydraulically raised ranging arms that are in turn connected to the shearer body figure 12 3 6 the shearer is driven by onboard traction units that drive a sprocket wheel that connects to rack bars mounted on the walkway side of the afc the face side of the shearer slides along the top or toe of the afc sigma section the shearer operators control the speed of the shearer and

the position of the cutting drum using handheld controls either radio remote or tethered cable to maintain a suitable cutting horizon within the seam and minimizing dilution modern heavy duty shearers are fitted with up to 2 000 kw of available power at the cutting drums enabling maximum cutting rates of more than 5 000 t h to be obtained shearers can be up to 14 m in length and weigh up to 100 t armored face conveyor the longwall afc is used to convey cut coal along the face to the maingate where it is connected to the bsl and outby conveyor system steel afc pans are constructed in sections typically 1 75 or 2 0 m long to match the width of roof supports flexible dog bone connectors are used to join the pans together a clevice bracket is fitted to each pan by which the roof support relay bar hydraulic ram attaches as each slice is taken on the longwall the afc pans are advanced or snaked into the newly cut area behind the shearer by activating the relay bars in sequence the afc supports the shearer body as it runs along the face and the cut coal is conveyed along the afc by flight bars attached to afc chains driven by motors at both the maingate and tailgate face ends figure 12 3 7 inspection pans that feature a removable decking plate will typically be fitted as every fifth pan to allow for access to the bottom chain transition pans are used at the gate ends to transition the afc onto the drive frames and bsl maintaining tension in the afc chain is important with a chain tensioning device fitted at the tailgate drive frame often automated using an appropriate control system reed rods etc the tensioner operates by extending a hydraulic ram 1 m stroke fitted to the tail sprocket assembly while the tensioner provides fine adjustment on a day to day basis coarse adjustment is achieved by removing links from the afc chain important aspects of the afc include the following the drive motors must be able to start the afc when fully loaded the afc speed must be sufficiently faster than the shearer s cutting rate to allow coal to be cleared when cutting from tailgate to maingate the afc chain must be of sufficient strength to handle the applied tensions modern afcs can be supplied with up to 3 000 kw of installed power in afc pan sections with 1 300 mm raceways coal can be transported at rates in excess of 4 500 t h the current limiting factor for face length is the afc chain strength versus installed power modern chains are being developed so as to allow longer faces currently the two main afc drive systems are controlledslip transmission cst systems and fluid coupling systems cst systems incorporate clutch systems in the output stage of the afc gearbox figure 12 3 8 with the motor directly coupled to the gearbox fluid coupling systems are fitted between the motor and gearbox and incorporate valves to change the fill volume of the coupling thus changing the level of torque both systems allow

for controlled starting and load sharing between drives an afc maingate drive frame is shown in figure 12 3 9 beam stage loader crusher and boot end the bsl transfers the coal from the afc to the conveyor belt and includes a chain conveyor running over deck plates a crusher a gooseneck and a boot end the bsl is capable of handling a higher volumetric capacity than the afc to ensure that coal removal from the face occurs without accumulation at the maingate corner afc bsl intersection the crusher sizes the cut coal to ensure that oversize lumps are not carried onto the conveyor and they consist of a high inertia rotating drum fitted with large hammers to break up large lumps of coal that may have slabbed from the face or large pieces of stone from the roof the gooseneck is an elevated section that raises the coal so it can be transferred onto the belt the boot end consists of the tail pulley for the conveyor belt fitted to a sliding frame so that the bsl can be retracted a certain distance up to 3 m without moving the tail pulley location the boot end is fitted with tracks or pads for movement and features hydraulic rams so that the tail pulley can be leveled and adjusted for belt tracking monorail the monorail system provides a flexible services link between the operating longwall face and the pump station and transformer equipment services including electricity hydraulic fluid water compressed air communications and control are carried along the monorail system as either hoses or cables the purpose of the monorail is to allow the longwall to independently retreat typically 200 m two pillars before the need to relocate the services equipment further outby the monorail equipment comprises long length runs of hosing and cables mounted to trolleys supported by roofmounted monorail beams located adjacent the gate conveyor in the belt heading pump station the pump station provides the high pressure hydraulic fluid required to operate the longwall face equipment including the hydraulic emulsion supply for the roof supports and water for shearer cooling dust suppression and other auxiliary systems as required the pump station consists of sleds containing typically up to three pressure pumps including one high pressure set pump if used a high pressure water pump for the shearer a reservoir for the closed emulsion fluid system and the pump control units the pump station is typically located in the mine roadways on a steel framed sled or in cut throughs trailer mounted to assist movement around the mine hydraulic fluid can be either fed from a surfacemounted emulsion mixing farm via poly pipe to the pump station or from mixing tanks located on the sled electrical system a track mounted transformer located outby with the pump station reduces the panel s incoming power to the required face voltage prior to supplying the distribution and control board dcb equipment which is normally located on the bsl

the dcb provides isolation control and monitoring capabilities to the longwall face electrical equipment the face voltage is typically specific to the country e g 3 3 kv in australia and 4 4 kv in the united states the control and monitoring equipment which is the electrical control center for the longwall system is typically located on the maingate corner and provides an interface for operators with the longwall equipment longwall equipment variations alternate equipment configurations are available for alternate mining methods and or methods of operation top coal caving the roof supports for top coal caving feature a hinged rear canopy shield that when hydraulically lowered will allow coal that has fallen from the roof to flow onto a rear afc and thus be collected figure 12 3 10 although this feature can successfully allow for the recovery of coal from thick seams the caving methodology and sequence lowering of the canopy requires careful integration into the overall longwall sequence to ensure that the productivity of the entire operation is not compromised top coal caving requires a rear afc which is a second afc located behind the roof supports along the edge of the goaf and protected by the hinged rear canopy the rear afc is connected to the roof support by a hydraulic relay bar and chain the rear afc transfers the coal onto the front afc at the maingate end of the face hence it requires a short transfer loader that crosses the main front walkway in addition the rear afc drive unit must be located in the maingate area to fit all this equipment into the maingate area requires a special roof support configuration figure 12 3 11 longwall ploughs longwall ploughs are used in place of shearers in thin seam longwall applications a plough consists of a vertical steel frame housing a series of cutting picks that sits on top of the afc and is driven at high speed 2 5 to 3 6 m s the length of the face by a chain drive system figure 12 3 12 the plough cuts a thin slice of coal 150 to 250 mm in either one or both directions with the coal falling onto the afc to be cleared the longwall supports and afc are advanced after each cut in a staged manner because of the low seam height applications plough faces are generally fully automatic with the equipment being operated and monitored from the maingate longwall automation the longwall mining method is highly mechanized which lends itself to various aspects of automation this has been progressively developed over time a longwall face in operation is shown in figure 12 3 13 until approximately 2003 the level of longwall automation was limited and consisted primarily of inner equipment automation e g roof support sequencing and intra equipment automation e g shearer initiated support advance recent australian industry funded studies have advanced the capability of longwall automation to now include the following capabilities face alignment control

use of an inertial navigation system horizon control improved memory cut shearer state based automation shearer speed control based on position automated fly and straightening cuts automated gate road support operation open communication systems more user friendly software systems several mines are now testing these capabilities and they have been reporting improved productivities longwall equipment selection the detailed specification of a longwall system needs to be determined at the time of tender preparation and is based on the results of detailed studies carried out during the feasibility studies these studies include detailed mine planning geotechnical characterization caving studies and engineering analysis key specification aspects include face length web depth nameplate capacity npc and support density face length because of improved technology longwall face lengths have progressively increased from 200 m in the mid 1980s to 440 m today long faces in the world include 400 m at ulan in australia 385 m at cumberland in the united states and 440 m at prosper in germany the current limiting factor for face length is the afc chain strength versus installed power hence there is a trade off between face length and npc modern chains are being developed so as to allow longer faces current technology indicates that a 400 m face length with 4 500 t h npc is the current limit web depth the depth of cut made by each pass of the shearer can range up to 1 200 mm and commonly ranges from 800 to 1 000 mm there is a significant productivity benefit from having a deep cut as this translates to more coal per pass of the shearer however the greater the web depth the greater the required shield canopy length to achieve the required tip to face distance hence the cautioning aspect to web depth is the risk of face instability a detailed geotechnical study is required to establish an appropriate web depth that balances the need for high productivity with the need for face stability another approach is to use a half web or partial web system whereby only half the web depth is cut during each pass which can lead to better ground control in particular situations a half web system requires faster shearer speeds to maintain productivity however this has been very successful e g twenty mile mine in the united states if a half web operation is envisaged then the longwall specification will need to include partial push capability of the relay bars half stroke nameplate capacity npc is the rated capacity of the afc and is a measure of the coal output during the main cutting run of the shearer the most important aspect of npc is that it directly impacts the capacity of the entire underground coal clearance system the longwall belt conveyor must match the longwall npc or the longwall will not be able to operate at designed capacity this is commonly seen in practice for various reasons consequently the

shearer speed is commonly restricted to compensate similar to face length nameplate capacities have steadily increased in line with increasing technology up to the maximum of 4 500 t h in australia and 5 200 t h in the united states although high npcs can theoretically directly translate to higher productivities this is not always seen in practice with many of the highest producing longwalls in the world rated at 3 500 t h this is because npc is not the only aspect of the determination of productivity in general terms an npc of 4 500 t h will require 6 000 to 6 500 t h trunk belts 2 000 to 2 200 mm belt widths while an npc of 3 500 t h will require 5 000 to 5 500 t h belts 1 600 to 1 800 mm belt widths support density support density is the rated capacity of the roof shield and is expressed in units of metric tons per square meter of canopy area usually after the shearer has made its pass after cut again improving technology has allowed this capacity to increase over time with capacities up to 130 t m2 now available typical support densities are between 90 and 110 t m2 variables such as support width 1 75 m versus 2 0 m shield geometry and leg cylinder diameter have an impact on support density the required support density is determined using geotechnical techniques such as ground response curve modeling or fast lagrangian analysis of continua flac modeling modern supports typically have a support density rating between 95 and 115 t m2 typically the ratio of setting pressure to yield pressure will be specified at 80 however there is an increasing trend toward 90 set which requires use of a high pressure set hydraulic supply system longwall equipment specifications there are several longwall equipment suppliers throughout the world many of which offer complete systems the typical process of selecting suppliers is to conduct a formal prequalification process involving the following expression of interest all suppliers are requested to formally express their interest in supplying the equipment the expression of interest may be advertised in a suitable media however in most countries equipment suppliers maintain close relationships with organizations so they are aware of forthcoming tenders technical prequalification an indicative technical specification is issued to interested suppliers who are requested to prequalify by submitting a response as to whether they can meet the equipment technical requirements provide examples of where they have supplied before and provide sufficient equipment description to satisfy the prequalification commercial prequalification often attached to the technical specification the commercial prequalification requests answers to various questions of a commercial aspect relating to security of supply financial strength of the supplier warranty provisions offered and a budget price for the supply of the equipment short listing following the prequalifica

tion process a short list of two to four suppliers can be selected these suppliers would then be subjected to a formal tendering procedure for the supply of the equipment longwall equipment sizing longwall equipment size is generally referred to as its nameplate capacity selecting the longwall equipment size relates to the required levels of productivity hence it is important to understand the relationships between the longwall system its npc and its productivity the relationship between the longwall system and annual productivity is shown in figure 12 3 14 two important productivity key performance indicators kpis are the average metric tons per operating hour and the operating hours per week for an operating mine these statistics can be collected as part of the production monitoring data for a new longwall operation these kpis need to be derived which requires a combination of modeling and benchmarking to model longwall productivity the following definitions are used nameplate capacity npc npc is the rated instantaneous capacity of the longwall e g 2 500 t h this is the load that is carried by the afc during its main cutting run and will occur somewhere in the cycle for between approximately 10 and 20 minutes in some instances the npc is determined restricted by the coal clearance system process cycle capacity pcc pcc is the average capacity achieved during one full cycle at maximum efficiency this is equal to the metric tons produced in cutting maingate tailgate and tailgate maingate divided by the time taken e g 1 800 t h actual productivity this is the actual operating rate achieved at the mine over an extended period as determined from the statistical reporting system e g 1 000 t h depending on the analysis period the actual productivity will fluctuate from very low levels up to but not exceeding the pcc productivity reduction factor prf prf is the difference between the actual operating rate and the pcc it is a measure of efficiency and reflects various issues such as adverse mining conditions operator skills motivation and organization for new mines this can be derived by benchmarking against similar operations and will generally range between 55 and 75 operational availability this is a measure of the average number of hours in a shift that the longwall is producing divided by the total hours in a shift e g 56 it is calculated by dividing the actual operating hours recorded over an extended period by the planned operating hours the difference between the two measures is the amount of unplanned downtime that occurs either due to equipment breakdowns or process delays for new mines the operational availability can be derived by benchmarking against similar operations and will generally range between 50 and 70 the longwall model will calculate the pcc assuming everything operates to its design capacity however in reality the longwall does not always

operate in this manner numerous operational issues such as operator skill and attention the presence of geological anomalies taking fly cuts straightening cuts slowing for lumps on the afc and gate end cleanups all slow and impact the cycle the prf takes this into account and is a measure of how efficient the longwall cycle is performed the prf will fluctuate from very low levels up to very high levels but can be measured over an extended period for use as a benchmarking tool in addition numerous unplanned breakdowns mechanical electrical and operational delays occur which reduce the total amount of operational time available within a shift this is termed operational availability also called utilization pcc is calculated as follows pcc equals the cycle metric tons per cycle time where cycle metric tons equals metric tons cut from the maingate to the tailgate and the tailgate to the maingate this is a function of face length extraction thickness web depth and wedge cuts at gate ends cycle time equals the time taken for the shearer to cut from the maingate to the tailgate and the tailgate to the maingate this is a function of npc shearer speed shearer acceleration snake length cutting mode bidirectional and unidirectional and face length when calculating the pcc it is important to conduct some operational reality tests regarding issues such as the following excessive shearer speed sufficient shearer underframe clearance excessive drum rotational speed adequate ranging arm power adequate afc power typically the engineer should have some operational experience with the longwall method to perform these calculations longwall top coal caving calculating the productivity of the longwall top coal caving ltcc method creates an additional level of complexity however the same principle can be applied it is important to accurately model the ltcc process particularly the caving cycle as to when this occurs in the process typically the main cut and the cave cannot occur at the same time as the bsl and the section conveyor cannot handle the dual output therefore the cave either occurs while the shearer is stationary at the gate end or while the snake is occurring therefore the caving cycle is generally the rate determining step for the process cycle with several factors to consider including the capacity of the rear afc the number of caving drawpoints the time required to cave each support and the method of caving manual or automatic the key benefits of ltcc are resource recovery more of the coal seam is recovered and improved development ratio more longwall metric tons per development meter however this can often be at the expense of productivity when compared to a conventional high capacity longwall and care is required when designing and modeling the process other longwall equipment considerations other engineering considerations when selecting a longwall system include the f

ollowing afc power demand calculations these calculations are performed to ensure that sufficient afc drive power is installed so that one can be able to start a fully loaded afc with sufficient reserve chain pull electrical load flow study this study is performed to ensure that adequate electrical power is available to cater for both normal loads and for starting under adverse conditions hydraulic flow simulations these simulations ensure that the hydraulic system is sufficient to meet the roof support requirements including routine las cycle up to three supports moving at once high pressure set gate end advance and to cater for adverse loading events room and pillar mining equipment r p mining can be divided into two categories first workings development and second workings extraction both processes require equipment to cut and gather the coal at the mining face convey and discharge the cut coal onto a section conveyor and install ground support at the mining face ancillary support equipment is required to provide ventilation and power there are numerous options of equipment that can be used as shown in figure 12 3 15 standard r p mining equipment standard r p equipment and systems include cms roadheaders haulage units feeder breakers mobile bolters electrical power ancillary equipment and mobile roof supports continuous miners a cm is a large electrohydraulic machine that extracts the coal to form a rectangular profile roadway or tunnel it features a rotating cutterhead drum laced with rock picks at the front the cutterhead is driven into the coal face thus breaking out the coal the broken coal falls to the ground and is loaded onto a centrally located chain conveyor using a loading apron and gathering arms spinners or east west conveyor the coal is conveyed through the body of the cm and loaded into coal haulage units typically shuttle or ram cars which come to the rear of the cm to be loaded modern cms are fitted with hydraulic drill rigs that drill and install the primary ground support as the roadway is formed roof and rib bolts platforms are provided for the drill rig operators to stand on while the cm is cutting coal commonly the drill platforms feature hydraulically operated temporary support mechanisms to protect the operators from roof or rib collapse the cm is fitted with caterpillar tracks to allow it to be propelled forward and backwards skid steering cms are also commonly operated by handheld radio remote control units although older units have pendant controls cable connected or operator cabins the three main types of cms are as follows 1 simultaneous cut and bolt this type features a full roadway width cutting head e g 5 2 m that can sump into the coal face while the body of the cm remains stationary resulting in the ability to use the drill rigs at the same time as cutting coal figure 12 3 16 2 sequential cut and bolt this type features a full roadway

width cutting head however the body of the cm moves forward when cutting coal and thus drilling cannot occur at the same time 3 place change or cut and flit these cms are not fitted with drill rigs often have a narrower cutting head e g 3 5 m and are used for the place changing method of mining other machines are used to install the ground support in these applications figure 12 3 17 roadheader a roadheader can be used as an alternative to the cm a roadheader is a large electrohydraulic machine that extracts the coal to form an arched profile roadway often extracting a portion of the stone roof roadheaders are fitted with a pineapple shaped cutting head attached to a slewing boom similar to the cm the cut material falls on a loading apron equipped with gathering arms or spinners that directs the material to a central chain conveyor that discharges from the rear of the roadheader coal haulage coal haulage units are required to convey the cut coal from the cm to the gate conveyor loading point key requirements include a large payload and rapid and flexible maneuverability three types of units are available namely electric powered shuttle cars scs with a tethered trailing cable batterypowered ram cars and diesel powered ram cars cable electric scs figure 12 3 18 are commonly used for longwall roadway development and offer the benefits of being fumeless do not require refueling or battery replacement do not generate large amounts of heat and do not require turning around shunting to discharge their disadvantages include attention to cable handling cable damage while in operation and a limited travel distance of approximately 200 m the length of the trailing cable ram cars offer unlimited travel distance however they require regular refueling and require shunting to discharge if battery ram cars are used figure 12 3 19 then a battery charge station is required which is commonly placed some distance from the working face because of their construction and ventilation requirements a ram car in operation is shown in figure 12 3 20 several scs can be used behind each cm however most australian mines use a single car in longwall gate roads because of the short shuttle distance the congestion caused by two scs and the need for an additional person in the crew if ram cars are used then at least two units are required to provide service while one is being refueled for pillar extraction methods several scs need to be used between two and four to rapidly clear the coal when a pillar split has commenced as an alternative to coal haulage units various techniques of continuous haulage have been developed over time including flexible conveyors bridge conveyors and addcar systems careful panel design is required for continuous haulage to gain the benefits of this technology consideration is required for roadway width and height cutting sequence clearances face ventilation cut through angles

access to the face by vehicle and conveyor extension process continuous haulage has been successfully used in several non longwall coal mines and potash mines throughout the world feeder breaker feeder breaker units are typically required for large capacity coal haulage units these consist of coal hoppers that accept the rapid discharge from the haulage units and then size and discharge the coal at a steady rate onto the section conveyor belt feeder breakers are fitted with crawler tracks so that they are self advancing during panel extensions however they are typically left stationary during the mining process an advantage of feeder breakers when using large haul units is their ability to protect the gate conveyor boot end from knocks collisions which otherwise will misalign the conveyor for a high capacity development system a basic requirement is the ability to quickly drive the haulage unit up to the discharge point and rapidly discharge the load feeder breakers can provide this requirement whereas boot ends require a slower approach to avoid damage some suppliers offer mobile boot ends that are a feeder breaker with a tail pulley and sufficient power to extend the conveyor belt pull belt out of the loop take up when they advance this can assist the belt extension process however care is required to ensure the mobile boot end can withstand the belt tension as the panel extends staking props are often fitted to the mobile boot ends to anchor against the roof so that the belt tension does not pull them back mobile bolters when the place changing cut and flit method is used roof support is installed by an independent bolting unit that trams into the heading when the cm is flitted out modern mobile bolting units are electrohydraulic powered are fitted with up to four roof bolting rigs two rib bolting rigs temporary roof and rib protection canopies and have provision for the storage of consumables electrical power electrical power to a development panel is reticulated at the country specific mine supply voltage e g 11 kv in australia to a section transformer via an isolator switch located at the start of the development panel the transformer located within 600 m of the working face steps the voltage down to typically between 950 v united states and 1 000 v australia and is then fed to a distribution control box located within 200 m of the face the distribution control box features up to seven outlets to which trailing flexible cables feed the various devices at the face cm scs fan feeder pump etc cable sizes reflect the various power demands of the equipment with the cm cable typically 120 mm2 and other cables typically 35 mm2 the section feed cables are typically 150 mm2 ancillary equipment auxiliary fans are required to provide immediate face ventilation typically these are centrifugal exhaust fans mounted on relocatable sleds or trailers placed in close proximity to the operating face

in conjunction with either lightweight fiberglass or steel ducting 600 to 750 mm in diameter these fans effectively place the fan s negative pressure suction inby of the cm operators minimizing dust and methane buildup at the face the fans commonly feature variable inlet vanes to control airflow and methane degassing valves to prevent high gas levels from passing over the fan blades the number of fans required for each working face depends on the number of roadways being developed the method of working and the level of activity commonly between one and three fans are required an alternative method of face ventilation is by forcing ventilation in conjunction with dust scrubbers mounted on the cm in this case relocatable axial or centrifugal fans are located in close proximity to the face with air directed using flexible or rigid ducting or brattice sheeting panel dewatering is typically through pneumatic diaphragm pumps as these pumps are able to operate in this arduous environment with minimal attention because of their relatively low output flow rate and their pressure head capability they normally report to a nearby electric pump pod the electric pump pod units are capable of the greater pumping duties required to reach the central mine dewatering station before delivery to the surface because of the high solids content associated with mine dewatering system design and pump selection requires sound detail to avoid the settling of solids in the reticulation system causing pipe blockages mobile roof supports for some of the secondary extraction methods e g wongawilli mobile roof supports that are also called breaker line supports are used to provide a level of protection along the goaf gob edge these are electric powered four legged hydraulic roof supports mounted on caterpillar tracks with pendant controls typically three units are used in a section breaker line supports are often used during longwall takeoffs to provide roof support as the roof supports are removed r p development equipment selection the detailed specification of the r p equipment needs to be determined at the time of tender preparation and is based on the results of detailed studies carried out during the feasibility studies these studies include mining method analysis panel configuration design development process studies productivity requirement studies ground support specification and engineering analysis key specification aspects include mining strategy development for longwall first workings r p secondary extraction pillar splitting etc face mining method place change in place sequential in place simultaneous type of cm coal clearance method continuous haulage or cars bolting method mobile bolters cm mounted rigs and panel configuration development equipment specification the key aspects in selecting the most appropriate development arrangement and equipment specification include the following mining

method development or extraction panel design number of headings pillar width angle of cut throughs and roadway width cutting method in place or place change cutting machine roadheader or cm type of cm sequential simultaneous or place change system of coal haulage sc ram car or continuous haulage number of scs one or two number of bolting units one two or three method of face ventilation method used table 12 3 3 summarizes a typical development arrangement for a longwall mine and table 12 3 4 summarizes a typical general specification for the development equipment for a first workings development system r p development equipment sizing development productivity is fundamentally measured in meters advanced per week or in metric tons per week for which the conversion is relatively simple therefore development equipment is typically sized to match a desired advance rate e g 8 m h or output e g 250 t h unlike a longwall where the equipment is completely integrated into a highly mechanized system development is a collection of equipment that requires the mining engineer to integrate it into a process the key process requirements are cutting coal clearance and ground support and these are conducted by separate and independent items of equipment development productivity can be calculated in a similar manner as longwall productivity as shown in figure 12 3 21 however for development calculating the process advance rate is more complex as itvneeds to take into account that the cm is continually moving from a fixed conveyor discharge point two example development productivity cases are given in the following discussion the first example shown in table 12 3 5 is for a conventional two heading gate road entry for a longwall panel using a simultaneous miner bolter with a single shuttle car 100 m cut through spacing the second example shown in table 12 3 6 is for a three heading development using the place change method 60 m cut through spacing this example compares three mines with slightly different plunge depths definition of the terms used in development productivity modeling include the following nameplate capacity npc npc is the peak cutting and loading rate of the cm e g 950 t h it is often quoted by the equipment supplier and can be used to determine the time taken to load a shuttle car or to load onto a continuous haulage system nameplate advance rate nar nar is a measure of the peak advance rate of the cm it is the time taken to cut and load one car drive the car to the boot end discharge its load and then return divided by the cut out distance of the cm e g 0 5 m nar is a useful factor when sizing the section conveyor belt process advance rate par par is a measure of the maximum advance rate during the cycle assuming that all activities are completed at their peak rate with no delays the importance of this factor is that it can be determined by

modeling and using actual measurements observed at the mine or from benchmarking therefore it is a scientific means of calculating productivity the par will be less than the nar as it takes into account bolting extending the ventilation ducting and any other process related activity such as repositioning the miner and lowering the canopy the par will only be evidenced in practice over short durations at the mine e g 2 to 6 hours as it is rare that the process flows at their peak rate over an extended time in the development process actual advance rate aar aar is the advance rate achieved at the mine over an extended period e g 3 6 m h aar is measured from the mine statistical reporting system for an operating mine for a new mine this would be calculated from the modeled process advance rate aar par prf process reduction factor prf prf is the difference between the aar and the par it is a measure of efficiency and reflects various issues such as adverse mining conditions operator skills motivation and organization and the complexity of the mining process for new mines this can be derived by benchmarking against similar operations and will generally range between 45 and 70 operational availability this is a measure of the average number of hours per week that the development section is operating divided by the total planned operating hours in a week e g 56 it is calculated by dividing the actual operating hours recorded over an extended period by the planned operating hours the difference between the two measures is the amount of unplanned downtime that occurs either due to equipment breakdowns or process delays the time to extend the panel is not included in planned operating hours for new mines the operational availability can be derived by benchmarking against similar operations and is generally between 50 and 70 coal clearance equipment the coal clearance system provides the means for transporting run of mine rom coal from all mining units operating in the underground mine to the surface rom stockpile the coal clearance system consists of various types and sizes of belt conveyors which are designed specifically for the various duty requirements associated with each mining unit the loading pattern from each production face varies depending on the mining method for example longwall mining will produce a fairly constant flow for 15 to 20 minutes while the main cut is taken punctuated by 10 to 20 minutes of irregular flows as the snake is taken first workings development will produce very short duration high flows of 1 to 2 minute duration as each sc is discharged with a frequency between cars of 10 to 20 minutes occasionally underground surge systems are used bins bunkers etc to smooth the flow and allow for lowercapacity outby conveyors to be used the required duty of a coal clearance system depends on the mining system for example if the mine employs a 3 500 t

h npc longwall system and up to three development units operating simultaneously i e one unit completing mains development and two units completing gate road development then the coal clearance system volumetric capacity duty would be as follows simple method longwall gate conveyor 3 500 t h matched to npc of longwall development gate conveyors 1 000 t h loading rate reduced to approximately 500 to 600 t h by use of feeder breaker main headings trunk conveyor 5 000 t h to allow for longwall plus three development units detailed conveyor calculations are required to refine these duty requirements which are undertaken during the mine feasibility study these calculations take into account the panel gradients and panel lengths over the mine life and provide key specification data such as belt rating power requirements and the need for tripper drives mid panel booster drives specialist conveyor duty software is available to perform these calculations however some practical experience is required to audit the outputs and to ensure that realistic specifications are produced where conveyors are required to run downhill braking units are required because gate road conveyors are extended as the panel is developed and then retracted as the panel is extracted care is required in the calculations to ensure that the maximum duty case is calculated this is especially prevalent where undulating seams are encountered both up and downhill sections because the development conveyor duty is commonly significantly lower than the longwall conveyor duty the following options are available when developing the gate roads install a small development conveyor e g 1 050 mm for development and then fully replace this conveyor prior to longwall extraction install the longwall conveyor drive head and structure e g 1 600 mm but with a narrow belt 1 200 mm and minimal drive units and then replace the belting prior to longwall extraction and fit additional drive units install the longwall conveyor drive head structure and belting e g 1 600 mvm but twith minimal drive units and then fit additional drive units prior to longwall extraction preliminary specification table 12 3 7 is an example of the general requirements for the coal clearance equipment for a new longwall mine specific details would also be required in regard to requirements for related aspects including the following dust control and belt scrapers transfer chute design detection systems for belt wander belt alignment belt tear chute blockage belt slip and belt tension belt monitoring emergency stop communication systems gas fire monitoring e g carbon monoxide detectors ancillary equipment ancillary equipment used in underground coal mines must comply with the local statutory regulations which are intended to ensure that these vehicles are safe to operate in potentially hazardous environments the following comments relate to

australia which has strict vehicle requirements mobile equipment mobile equipment refers to the diesel vehicles that are typically used in underground coal mines they are categorized as personnel carriers material transporters and utility vehicles or special purpose vehicles typically these vehicles were developed over the years into purpose designed vehicles to suit the rugged environment of the underground coal mine statutory regulations require that these machines comply with diesel engine exhaust emission limits be flameproof have wet braking systems and be fitted with various shutdown monitors high temperature low scrubber water etc recently nonflameproof personnel carriers modified toyota land cruisers have been successfully introduced into some queensland mines however these suffer some restrictions in their ability to travel into working faces depending on methane levels and statutory zones personnel carriers transport of personnel from the surface to the work site and back is through the use of personnel carriers having typical carrying capacities of up to 14 people including the driver several models are available from suppliers with some companies requiring additional safety features such as forward facing seats and seatbelts some personnel carriers are converted into maintenance utility vehicles with a rear flat tray for carrying tools and spares material transporters and utility vehicles material transporters and utility vehicles include a range of vehicles for duties such as equipment and materials transport and other work duties around the mine typical configurations include the following load haul dump vehicles lhds lhds are used with attachments including fork tines loading plates buckets jib cranes elevated work platforms augers hydraulic drilling rigs stone dusters pipe installation platforms cable reelers and trailers lhds are typically the workhorse of the modern mechanized mine and are commonly fitted with a quick detach system typically there are one or two of these vehicles with a capacity between 7 and 10 t at every panel the mine will also have some higher capacity lhds with capacities of between 10 and 15 t these are used for roadwork and other heavyduty requirements a utility loader with a 10 t capacity is shown in figure 12 3 22 multipurpose vehicles multipurpose vehicles are specifically designed to pick up and carry modular pods and tubs bobcats occasionally mines will use small lhds such as bobcats for light cleanup duties special purpose vehicles these are purpose built vehicles that perform only one or a few specific duties on a periodic basis with typical configurations including the following heavy duty lhds 40 to 50 t used for transport of longwall equipment and some belt conveyor equipment figure 12 3 23 roadway graders and rollers used for road construction and maintenance roof support trailers these are used in conjunction with

an lhd to transport longwall roof supports specific shield haulers used for larger roof supports shearer transporter used to transport the longwall shearer mine dozer diesel powered heavy lift machine for moving longwall equipment mule electric powered heavy lift machine for moving longwall equipment mobile generator used for cm flits mobile bolters used for outby roof support drilling rigs used for gas drainage and or exploration some of this equipment is used infrequently and there are often a number of leasing companies available to service the mine requirements the majority of mines in australia rent this equipment during longwall moves although if the mine purchases equipment that is not typical e g roof supports weighing more than 35 t then the mine may be required to purchase its own longwall move equipment the general ewatering pump stations area lighting table 12 3 9 provides an example of the various power supplies utilized in the underground electrical power distribution system for an australian mine for other countries the voltage and frequency will differ e g the united states uses 60 hz and 950 v for the development face and 4 2 kv for the longwall face the applicable standard may revert to international standards as opposed to australian standards all underground power circuits operate using an impedanceearthed system it earthing system code in accordance with as 3007 2 2004 the purpose of this is to limit the magnitude of an earth fault and hence the resultant touch potential thus reducing the risk of electric shock from indirect contact also because most electrical faults occur as a result of cable damage cables used underground are generally constructed with an earthed screening around each phase conductor so that the initial type of fault is a low energy earth fault rather than a high energy short circuit controlling the energy dissipated during an electrical fault in this way reduces the risk of injury to personnel fire ignition of methane or dust or damage to equipment as a result of an arc flash the performance of the underground electrical powerdistribution system has a significant impact on the performance and reliability of the electrically powered mining equipment consequently the design for the underground power supply system needs to account for the following key power supply performance parameters fault level thermal capacity voltage regulation transient motor starting capability particularly with respect to the large longwall afc motors other factors that may influence the quality of the power supply and require detailed study include transient stability and harmonic distortion all electrical equipment must comply with the relevant parts of as nzs 4871 2002 in addition equipment for use in an area designated as an explosion risk zone must be certified as being explosion protected by complying with the relevant parts of as nzs 60079 10 2004 un

derground reticulation the underground reticulation system is typically comprised of the following components power supply 11 kv from the surface to underground via either cables in the mine entry roadway or via cased boreholes underground switchboard 11 kv underground reticulation cables 11 kv mounted along the roof of the development roadways section circuit breaker isolators for a new mine the power supply from the surface to underground typically occur in the following two stages the initial development phase and the operational phase the power supply for the initial underground development is typically via the mine access portals and is achieved by means of an overhead power line installed overland from the surface main mine substation e g 132 11 kv to a site in close proximity to the portals this could be to the top of a highwall if the entry is via an open cut or trench a cable then connects to switchgear located at the portal entry switch room which will control the power supply entering the mine and will be interlocked with the mine ventilation fans to trip the underground power in the event of a ventilation failure this feeder will initially provide a power supply for the following loads initial underground power supply development equipment temporary drift conveyor temporary ventilation fans pumps and other ancillaries in the longer term this feeder will also provide the permanent power supply for the drift conveyor pumps and other loads located in the portal area prior to the commencement of the longwall the power supply to the underground operations will require upgrading typically by using borehole feeder cables originating from the surface main mine substation 132 11 kv two 300 mm2 12 7 22 kv cross linked polyethylene xlpe single point suspension borehole cables provide the power supply to an underground switchboard 11 kv located in the main headings at the bottom of the boreholes the underground switchboard is configured with two incomers one bus tie and four outlets for the control and distribution of the power supply to the following loads longwall feeder development and conveyor feeder pit bottom feeder one spare the underground cable selection process is a balance of the following considerations achieving acceptable performance parameters minimizing the number of different types and sizes used underground practical sizing for the purposes of handling the cables underground all 11 kv reticulation cables must comply with australia new zealand safety standards as nzs 1972 2006 while cable couplers must comply with as 1300 1989 load flow modeling of the underground reticulation system is required to accurately specify the required cable types and sizes a typical specification is listed in table 12 3 10 although paper insulated lead covered pilc cable has traditionally been used for underground reticulation cabling the use of xlpe cable is oft

en recommended because it offers superior current carrying capacity for the same size cable approximately 50 increase in capacity over the pilc type this is important to provide adequate current rating while keeping the cable size as small as possible for cabling handling purposes section circuit breakers are used to sectionalize the underground reticulation system 11 kv and to control the power supply entering the various areas of the mine section circuit breakers are required at each gate road entry to control the power entering that development or longwall panel the section circuit breakers incorporate the following features incoming and through going supplies two switched outlets circuit breakers and protection relays for each outlet isolation and earth switches for each outlet programmable logic controller plc network and control equipment ip56 rated minimum enclosures ip56 is an international electrotechnical commission weatherproof standard arc fault rated enclosures using arc fault control methods base frame wheels draw bar and stabilizing leveling legs development panels the electrical equipment for the development panels includes the following items panel substations panel distribution and control boxes and panel trailing cables the power supply equipment up to the input connection of the development substation which includes the 11 kv cables and section circuit breakers forms part of the underground reticulation system the panel substations transform the incoming power supply 11 kv to a 1 000 v nominal supply for distribution to the development machines and ancillary equipment the current growing trend in the industry is for 2 mva megavoltampere rated substations to accommodate the increased power requirements of the development equipment generally these substations are not flameproof as they are located far enough outby of the working face to be in a nonhazardous zone the panel substations typically incorporate the following features incoming and through going high voltage supplies high voltage circuit breaker and protection relays isolation and earth switches for the transformer section 2 0 mva 11 1 05 kv transformer four outlet low tension end 1 000 v switchgear and protection relays plc network and control equipment communication components for the cm ip56 minimum rated enclosures arc fault rated enclosures using arc fault control methods base frame wheels and draw bar longwall panel the electrical equipment for the longwall panel includes longwall substations longwall dcb s monorail and face trailing cables and control monitoring and signaling equipment the power supply equipment up to the input connection of the longwall substation which includes the 11 kv cables and section circuit breakers forms part of the underground reticulation system the longwall electrical equipment is normally included in the scope of supply for the longwa

ll equipment package underground conveyors the electrical equipment for the conveyor systems includes conveyor substations conveyor starters power and control cables and control and signaling equipment the power supply equipment up to the input connection of the conveyor substation forms part of the underground reticulation system 11 kv the conveyor electrical equipment is normally included in the scope of supply for the respective conveyor system package dewatering pump stations the electrical equipment for the dewatering pump stations includes substations pump station starters power and control cables and control and monitoring equipment generally the dewatering pump stations will obtain their power supply from a substation conveniently located nearby such as a conveyor substation or starter however dedicated substations may be required for more distant pumping locations such as the pit bottom area to provide a power supply for pumping an equipment tramming facility and general power for the area underground lighting underground lighting is typically required in the following locations electrical equipment locations such as switchboards section circuit breakers substations and dcbs conveyor drives loop take ups and transfer points including the walkways in between pump stations transport access portal other areas where a particular risk exists and which can be mitigated by lighting initial 50 m approximately of the mine access drifts to assist the transport vehicle operators in the transition from daylight to darkness and vice versa two types of development exist in mining primary and secondary the main difference between these two types is the life expectancy of the development primary implies long life and includes such items as shafts declines and other mine accesses main tunnels haulages crosscuts ore and waste passes crusher chambers and other large chambers thus primary development is the development of the more permanent features of a mine these do not have to be developed early in a mine life but they need to have a relatively long life span secondary development is more temporary in nature and tends to be associated with the needs of a particular production unit or units in many cases this type of development is consumed as production proceeds thus a typical lifetime for this type of development is 1 or 2 years and is often less than this development is a necessary part of mining as it provides the infrastructure with which production of ore can be undertaken development is essential but represents an additional cost in most situations this cost is not recoverable until mining operations such as preparing to excavate secondary development undertaken within the ore body itself begin however in some situations e g coal mine development of revenue generating material and some payback is obtained during development primary development is normally almost entirely undertake

n in waste external to the ore body the exception to this is the mining of bedded deposits where some ore in primary development can sometimes be taken development requirements are influenced by a number of parameters the following list is far from exhaustive ore body type ore body depth mining method selected economics ore body extent the mining method selected is a significant influence especially with regard to secondary development but also with respect to primary development as it is important to the layout and size of the primary mine infrastructure to be able to handle the production of waste and ore over the lifetime of the mine it is also important to account for possible variations in production over the mine life as well given the cost of development and the usual need for mining companies to generate income relatively rapidly due to the long lead times between exploration and eventual production the aim in most cases is to bring production on stream as quickly as possible however the long term capabilities of a mine are dictated by the capabilities of the primary mine infrastructure the need for accurate geological information environmental considerations and numerous other factors during the mine planning stage the primary infrastructure and secondary development are investigated it is essential at this stage that a thorough economic technical and sensitivity analysis of all possible mine production systems be undertaken to evaluate the project on a logical basis mine development serves a number of purposes which include accessing the mineral deposit provision of accurate geological information to aid the production process and providing accurate information on other technical needs such as geotechnical characteristics of the rock water inflow rates ventilation pollutant inflow rates and so on figure 12 4 1 illustrates some of the openings whose construction will be discussed in this chapter excavation methods development openings in mines can be excavated in one of two ways either by drilling and blasting or by mechanical excavation both techniques are used and are discussed in the following sections for various types of development openings both have their advantages and disadvantages mechanized mining has the following advantages compared to drilling and blasting methods higher advance rates development rates smooth development profile and superior rock stability tunnel boring machine and mobile miner circular profile in particular minimized ground support reduced overbreak less surface disturbance less noise lower running costs continuous haulage system such as conveyor automated haulage systems less labor required automation precise direction control laser guidance safety no explosives however there are disadvantages also including cost difficultly dealing with variable ground conditions and in hard abrasive rocks high rates of wear on

the cutters can lead to high operating costs and long periods of downtime for cutter replacement and other maintenance shafts according to the mining glossary ausimm 2007 a shaft is defined as a primary vertical or non vertical opening through mine strata used for ventilation or drainage and or for hoisting of personnel or materials connects the surface with underground workings although this is a good definition shafts can also be internal within a mine connecting vertically separate workings as well as levels in some ways a better definition of a shaft is a vertical or subvertical opening in a mine that is excavated from the top downward as opposed to a raise which will be discussed in a later section of this chapter shafts are the most important capital openings for deep mines shafts provide all the services for underground operations including fresh air transport of ore and supplies personnel transport power communications water supply and drainage the depth of the shaft has a direct influence on the total time to develop a mine prior to first production in some cases shaft sinking can account for up to 60 of the total development time consequently proper selection of the shaft sinking method to minimize sinking time and ensure uninterrupted operation is of great importance determining the shaft diameter and hoisting depths requires consideration of future mining needs it is generally better to overdesign the shaft in the initial stages of the project life span than prevent feasible increases in production later by the creation of a bottleneck designing the excavation requires data to be gathered and solutions to elements within the design to be determined in order to design the excavation the following need to be taken into account site geology including a description of the geological column usually in a table form identifying the rock strata their geotechnical parameters groundwater levels including water heads inflow rates and chemical contamination if any determination of the shaft diameter choice of shaft sinking technology description of the shaft lining should include a list of lining sections with their thicknesses detailed shaft foundations including their location and dimensions shaft collar depth foundations thickness and materials to be used the type and number of openings with a description of their function dimensions and elevations should also be detailed shaft sump including depth structural characteristics pumping arrangements and cleaning system surveying data calculations including ground and water pressure acting on the shaft lining the resulting lining thicknesses shaft insets and their dimensions and airflow capacity timetable of construction cost specifications engineering drawings shaft shape and diameter numerous factors regarding the shape and diameter of a shaft need to be considered such as shaft duties and purpose

such as rock hoisting ventilation and worker and materials transport is the shaft to be used for a single purpose or for a variety of purposes what are the quantities involved the amount of water to be handled number of pipes and the diameter of the pipe are important parameters as is the method of hanging these within the shaft ground conditions size of mining equipment to be handled by the shaft is the equipment to be handled in single or multiple parts for transportation of oversize elements is suspension underneath the shaft conveyance possible the expected lifetime of the shaft the majority of shafts in modern mining are circular in shape however rectangular cross sections were common in the past although rectangular shafts were more popular within metalliferous mines most modern shafts tend to be circular other shapes are used such as square or elliptical circular shapes have inherent strength and have the advantage that for a particular cross sectional area the minimum perimeter results elliptical shafts tend to be used in areas where a significant horizontal stress field acts in a particular direction the major axis of the ellipse is aligned to this direction to enhance the shaft strength shaft linings shaft linings basically serve to support the shaft equipment and the walls of the excavation the minimum lining usually required to support shaft equipment is about 200 mm thicknesses required for wall support are generally higher in modern projects where round or elliptical shafts are employed concrete lining is used almost exclusively such shapes reduce airflow resistance make sinking easier and allowing the full advantages of concrete to be employed the placement of concrete can be mechanized allowing high sinking rates as well as lower costs the strength of the concrete can be adjusted as required and watertightness can be ensured within aquifers of moderate head where heads exceed the strength of concrete cast iron linings are employed shaft collars the shaft collar is the upper portion of a shaft which is anchored to the first footing as a minimum and to the bedrock if required overall the dimensions of the collar depend to a large degree on its depth cross section thickness the function of the shaft the overburden rocks hydrologic conditions the ground pressure on a diurnal and annual basis sinking method and additional loading conditions a simple collar design is illustrated in figure 12 4 3 the size of the shaft collar depends to a large extent on the loads in the immediate surrounding areas from the shaft and associated equipment examples of these loads include winding equipment and its associated dynamic loading loading impacting on the shaft collar area from other surface facilities and so on the shape thickness and choice of shaft collars depend on numerous factors but in essence depend on structural factors some typical collar shapes are illustrated in figure 1

2 4 4 shaft insets and sumps shaft insets should be designed to minimize airflow resistance model and or computational fluid dynamics models can be used to ensure this sumps at the base of a shaft and at intermediate levels to minimize pumping heads must be of sufficient depth or size to cope with the water flow requirements obviously suitable dewatering arrangements must be provided in addition periodic cleaning must be designed into the system to account for sedimentation which will reduce the holding capacity of the sump technology of shaft sinking the main classification between traditional shaft sinking and other specialist methods is based on the engineering and hydrogeological conditions of the individual site the conventional method of sinking is summarized in figure 12 4 5 and involves repeating cycles of face advance and lining erection without any previous ground stabilization such methods are applicable to good to fair quality rocks with limited water inflow into the shaft 500 l min specialized methods of shaft sinking include those applied to strong rocks that have an extensive fracture pattern resulting in a strong water inflow such as the south african goldfields coherent weak plastic flowing ground such as permafrost areas and loose water bearing grounds conventional drill and blast shaft sinking the main components within a traditional drill and blast shaft sinking operation are illustrated in figure 12 4 5 additionally the following classifications of traditional shaft sinking can be made based on the order of the main components of the operation series system in which sinking occurs with nonsimultaneous face advance and erection of the permanent lining the shaft is sunk in sections the length of the section depends on the rock quality in a particular section of the shaft the face is advanced first with a temporary lining when sinking is suspended the permanent lining is constructed fast sinking rates are not ensured however high capital expenditure is not required this method is commonly applied to the sinking of small diameter shafts at shallow to moderate depths parallel system in which sinking with simultaneous face advance and lining erection at certain distances behind the face occurs in this method a temporary lining is required as with the series system it is applied to the sinking of large diameter and deep shafts where the high capital cost can be paid off over the life of the project simultaneous system in which advancing the face and erection of the lining in the same shaft section occurs this method can be further subdivided into series simultaneous and parallel simultaneous methods the series simultaneous system involves face advance and lining completed in one cycle with each being done one after the other the lining is erected either downward or upward and temporary lining is eliminated the parallel simultaneous system involves simultaneous face advanc

e and lining erection shaft buntons beams guides and so on are usually installed following the advance of the permanent lining however in some methods of shaft sinking e g shaft boring they are erected following the completion of the shaft lining shafts can be sunk to the final depth or to depths required for certain stages of the mining process shaft extensions to subsequent levels need to be synchronized with reserve depletion in the upper levels to ensure uninterrupted production shaft extensions can be undertaken by either sinking or raising before sinking a shaft preparatory work needs to be done this includes determination of geological and hydrological conditions preparation of technical documentation and construction work associated with site preparation and construction of the shaft head the shaft site should be secure from flooding from a design perspective this would be based on a 100 year flood level numerous activities associated with the preparation of a shaft site include but are not limited to construction of access roads and materials storage areas grading provision of a water supply provision of electrical power and other power sources and temporary and permanent buildings drilling and blasting when undertaken correctly drilling and blasting will ensure correct size and shape of the excavation an even face surface for simpler mucking and drilling of the next round uniform sizing of the broken rock for effective mucking and safe and economic operation shaft blasting uses a variety of high explosives the correct product has to be used given site specific conditions such as explosive gas emissions reactive ground conditions and the presence of water electric detonators with millisecond delays are also employed the drilling pattern has a strong influence on sinking performance it should ensure accuracy in the shape of the excavation clean separation of rock from the bottom of the excavation even and sufficient comminution of rock avoiding cutoffs of neighboring active holes and avoiding damage to nearby shaft installations the drilling pattern depends on such factors as the shape of the shaft section rock strength cleavage dip of strata water inflow and hole loading structure typical drilling patterns for circular shafts are shown in figure 12 4 6 with small and moderate strata dip holes are placed on concentric circles drawn from the center of the shaft the number of circles is three to five depending on the shaft diameter the central hole is vertical and shorter about two thirds the length of the other holes this hole initiates conical excavation by cut holes and is especially recommended in strong rocks holes of the first central circle wedge holes are inclined toward the shaft center and are fired first to open an additional free surface and to make the work of the remaining external holes easier and more efficient they

are drilled with an opposite outward inclination of 65 75 four to ten holes are drilled for the first circle the ends of four holes are intended to create a square with sides of 180 300 mm for a larger number of holes a 400 600 mm circle is used in general the harder the rock the more first cut holes are required holes of the next two three and sometimes four circles crush and lift the excavated material the inclination of holes varies from 65 to 85 the farther they are from the center of the shaft the more vertical they are the distance between adjacent holes on a circle should not exceed 0 8 to 0 9 m for 32 mm and 36 mm cartridges for a 45 mm cartridge it should lie in the range of 1 to 1 2 m the last circles of perimeter holes are supposed to leave the shaft wall smooth they are holes drilled toward the shaft wall 65 85 less inclination for weak rocks and more for stronger ones in very weak rocks they can be slightly inclined toward the center of the shaft perimeter holes are placed in a circle with a diameter of 0 6 m less than the shaft diameter in rocks with a coherence factor of f 2 8 and 0 3 m for rocks with f 9 20 the distance from the shaft wall has to be such that in rocks with f 8 holes should not project from the design shaft section and in rocks with f 8 they should protrude about 100 200 mm outside the shaft contour the distance between rib holes in moderate strength rock is 0 9 1 2 m and in strong rocks 0 7 0 9 m drilling of shot holes blastholes can be drilled with sinker percussive drills hand held or mounted on a special shaft jumbo sinkers are self rotating driven by compressed air or hydraulically with air or water flushing two systems exist 1 series in which the drilling starts from the shaft center toward its perimeter after completion of mucking 2 parallel where mucking and drilling are done at the same time drilling starts from the perimeter toward the shaft center a shaft jumbo can be used only in the series system a series system is used in smaller shafts when a cryderman or similar type of mucker is used this system ensures precision in blasthole drilling the parallel system is sometimes chosen in large crosssectional shafts where cactus grab cabin muckers start to clean the sides of the shaft moving toward its center drilling is done in the cleaned area owing to parallel drilling and mucking some time saving can be achieved however it is at the expense of drilling accuracy blastholes can be drilled individually or by groups of drillers in the first case approximately 25 of the time is lost in changing bits in the second one shaft area is divided according to the number of drillers each group drills in one sector holes of a certain depth only bits are changed only when they get dull then the drilling rod is changed lengthened and work is done in the next sector wooden sticks are used to prevent holes from being plugged by muck drilling equip

ment is delivered in a special bucket known as a kibble or basket to the shaft bottom an elastic hose with quick coupling is connected to the main steelcompressed air tubing at the end of the elastic hose there is a distributor with multiple outlets and valves for particular drills drilling is a labor intensive operation that takes about 16 of the cycle time shaft drill jumbos have been developed to mechanize blasthole drilling these permit the drilling of larger diameter holes more precisely faster and deeper mucking in shafts mucking in shafts accounts for as much as 50 60 of cycle time table 12 4 1 details the different factors influencing mucking efficiency figure 12 5 2 in chapter 12 5 shows a schematic of the sinking phases with an eimco 630 loader mucking efficiency is also dependent on the degree of rock fragmentation and the height of the bucket for mechanical mucking medium sized fragmentation approximately 125 mm is most efficient generally in a muck pile after a shot 70 80 of the material is fine fragmentation and the rest is composed of larger fragments several types of muckers are in use depending on shaft shape size depth and location scrapers used in rectangular shafts and very largediameter shafts underground bunkers overshot loaders on caterpillars e g eimco 630 cryderman muckers cactus grabs light cactus grab heavy on central pivot a variety of systems are used some are driven by an operator e g eimco some are hand guided at the face when hanging from a platform and some are attached to the shaft wall cryderman and operated from a control cabin heavy cactus grab units have a rotating arm on a central pivot below the stage muckers can be electric pneumatic or hydraulically driven most shafts intersect one or more water bearing horizons so shaft dewatering is an essential part of sinking the characteristic features of shaft dewatering consist of changing water inflow rates changing the pump head filtering water to eliminate rock and mud particles and periodic changes in positioning dewatering devices dewatering can be considered in two ways either as limiting water inflow into the shaft or pumping water out of the shaft water inflow into the shaft can be diminished by initial drainage of the rock mass sealing of the water flow routes or by tapping the water the following are used to dewater the face rock kibbles and special water kibbles face pumps combination of face and stationary pumps combination of hanging and stationary pumps submersible pumps airlifts shaft lining construction the type of lining is determined by the following factors hydrogeological conditions shaft function planned lifetime shape of shaft section and its depth availability of construction materials construction cost in choosing a shaft lining the geotechnical properties of the rock and hydrologic conditions are decisive in most cases the chemical act

ivity corrosiveness of the water can also be an important consideration since modern new shafts often have automatically operating hoisting gear moisture sensitive they should be dry main shafts are usually planned for the entire mine life span so they should be constructed to minimize repairs and maintenance time temporary support is placed to protect the crew and equipment against falling rocks from the exposed shaft wall before a permanent lining is placed numerous systems of temporary lining are in use forms of temporary lining include rock bolts and mesh and other traditional methods of temporary support used in other mining operations depending on the shaft design and environmental conditions several different permanent lining systems can be used timber brick or concrete blocks monolithic concrete reinforced concrete tubing shotcrete anchor bolts shaft sinking stage sinking stages are an important element of any shaft sinking program they provide a working platform from which to perform the sinking operations they also ensure the safety of the personnel working at the shaft bottom an example of a sinking stage can be found on the andra web site 2009 this web site details the construction of test facilities in france for nuclear waste emplacement with access by shaft it shows the sinking stage being installed in the shaft following construction of the collar and gives technical details of the shaftsinking process as well as the development of the test galleries underground these can vary considerably in design from a simple single deck working platform to complicated multideck arrangements with facilities for drilling mucking shaft concreting or timbering and storage of hoses pumps and all other equipment associated with shaft sinking sinking stages have been used since the earliest shaft sinkings and have developed to maximize efficiency of the operations while minimizing delays the primary purposes of sinking stages are to enhance safety and maximize efficiency stages are suspended in the shaft by wire ropes and winches that are independent of the main hoisting system stages are raised prior to blasting and then lowered to their new position at the end of each sinking cycle sinking cycles are usually based on a certain height of concrete pour or timbering specialized methods of shaft sinking special methods of shaft sinking are used when the following conditions are encountered incompetent water bearing rocks e g quicksand weak unstable soil type rocks competent fractured rocks with high water inflows above 0 5 m3 min numerous special methods exist that encompass many variations in method and application because of environmental conditions and range of depths the characteristics of these are given in table 12 4 2 the following methods will be described in subsequent paragraphs depression of groundwater level grouting methods freezing method and shaft drilling an alternati

ve to these methods can be applied in some cases e g moving the shaft to another site depression of groundwater level prior to designing a method for depressing the groundwater the following decisions must be made choice of the most suitable method for depression of the water table under site conditions determination of the number spacing diameter and depth of the wells calculation of the radius of the depression cone amount of water to be pumped out and the height of the remaining wet portion of the water bearing strata evaluation of equipment hardware and pumping facilities choice of the auxiliary method to go through the wet remainder of the water bearing strata choice of a suitable water discharge plan to lower the water table in the area of shaft sinking the following systems can be applied pumping water from three to six depression wells drilled around the planned shaft perimeter placed on a circle about 4 m larger than the diameter of the shaft excavation this system is used most often draining water through drill holes to existing openings in the mine below draining water through drill holes to water absorptive strata e g dry sands under water bearing strata depression wells are drilled with 300 500 mm diameters without drilling mud they are spaced uniformly around the future excavation within the range of the water bearing strata they are equipped with filters perforated by round or elongated holes 5 10 mm in diameter spaced in a checkered pattern with an outside screen having eyes 0 4 0 9 mm in diameter sized against entry of sand particles generally the construction of the filter must be suitable for the character of strata to be drained submersible drainage pumps have a widecapacity range varying up to 350 m3 h calculations for well efficiency are based on principles of flow through a porous media the objective of the design is to determine a pumping regime to achieve a dry out zone surrounding the shaft this is created by the combined effect of the depression cones around particular wells the movement of water through a porous media is given by the darcy formula application of grouting based on geologic data the following decisions have to be made prior to the start of the project choice of the grouting system including the decision as to whether grouting will be done from the surface or from the bottom of the shaft determination of the grouting depth and determination of the length of the grouting sections determination of the number of grouting holes and their respective distances diameters depth and angle of dip and applicable drilling method construction and calculation of the grouting plug grout type to be used its density and admixtures injection system and required equipment determination of working pressure to be applied in particular sections and approximate calculations of grout consumption number of control holes and their locations set of

drawings corresponding to this list of considerations necessary hydrogeological information should contain number and characteristics of the water bearing water horizons chemical analysis of water flows and determination of their degree of corrosiveness flow velocity and coefficient of filtration properly performed grouting seals the rock mass in the area of the shaft making sinking possible by a conventional method grouting is performed from the surface when the water bearing strata are relatively thick and close to the surface or from the shaft face when the strata to be grouted are located at such a depth that it is not feasible to drill long holes from the surface grouting is most successful in waterbearing competent but fractured rocks grout fills the fractures and cracks creating a sealed zone around the shaft to be sunk and protecting it against water inflow grouting for sealing is often applied to improve tightness of the lining in existing shafts and to eliminate the voids behind the lining as well as to seal rocks in the immediate vicinity of the lining petterson and molin 1999 when grouting from the surface grouting holes with an initial diameter of 75 150 mm are drilled vertically around the future shaft on a circle with a radius 3 4 m larger than the excavation diameter holes should be straight and with their greater depth any deviation should be recorded grouting from the shaft face is illustrated in figure 12 4 7 and is recommended when water bearing strata occur below 100 m and when their thickness does not exceed 70 m the length of sections varies between 12 and 25 m holes are drilled along the length of the grouting section usually by means of heavy drill hammers or small rigs when space allows holes are directed at an angle in the vertical plane to get the hole bottoms 1 2 m away from the future excavation to increase the probability of encountering fractures by the hole they are also tilted in a tangential direction as shown in figure 12 4 7 the distances between the hole entries are 0 8 1 5 m spacing between the circles of holes is 1 5 3 m the number of holes drilled depends on the rock character shaft diameter and hydrogeological conditions practically it varies from 10 to 30 the initial drill hole size is 50 75 mm the end size is 30 50 mm the sequence of drilling and grouting is as follows 1 the first one half of the total number of holes is drilled and grouted during hardening of the cement the second half is drilled and grouted 2 all holes are drilled and grouted one after another starting with the hole with the largest water inflow 3 a pair of holes lying on opposite sides is drilled and grouted then the pair lying in the plane perpendicular to the first one is processed and so on this system is the one most often used being recognized as the best because the initial maximum distance between the holes lowers the probability of grout penetration from one hole to

another grouting plugs are put in the shaft bottom to prevent water and grout inflow into the shaft entrance pipes to the holes are secured firmly by grouting them in the plug the plug can be either natural or artificial a natural plug is composed of a protective layer of grouted rock 3 5 m thick depending on the strength of rock and expected pressure water and grout artificial plugs are used more often and are constructed of concrete or reinforced concrete they are constructed on the shaft bottom 2 4 m from the water bearing strata when grouting is completed the plug is demolished with explosives and sinking proceeds through the cemented section but again is suspended 2 4 m above the end of the grouted zone the shaft lining is constructed and the procedure with the plug is repeated until the shaft passes the whole water bearing zone details of the use of grouting to sink a shaft through waterbearing ground at the longos gold mine in the philippines are described by ayugat et al 2002 freezing method this is a method commonly applied where a relatively large thickness of water bearing strata has to be sunk through some good examples are presented in harris 1995 it is commonly applied when shaft sinking to the coal measures in the united kingdom due to the thickness of the water bearing bunter sandstone for a freezing process a suitable design procedure requires solution of the following problems calculation of the wall thickness of the frozen rock cylinder calculation of the diameter of the circle where freezing holes are located their number depth and spacing and the number and distribution of surveying and control holes selection of the drilling system for freezing holes specification of equipment choice of freezing pipes and their diameter and assessment of time required for drilling design of a temporary shaft head with a freezing basement planning of the freezing procedure in active and passive periods with specifications of freezing temperature heat balance required efficiency of freezing installation and freezing time determination of the system and time of thawing and liquidation of freezing holes initial planning of freezing technology shaft sinking and lining and complementary groutingthe cylinder of frozen rock when properly created allows the sinking of a shaft of moderate depth even in long sections using temporary lining this is desirable because there are fewer seams between lining sections the most important problem during the drilling of freezing holes is to ensure verticality of holes every hole is equipped with two columns of tubing as follows 1 outside the freezing pipe itself has a diameter of 102 152 mm and is constructed of steel with inside connectors so it is smooth on the outside and has a welded bottom 2 inside the inner tubing has a smaller diameter of 32 51 mm is made of steel or polyethylene pipe and has inside diameter and wall thickness of 75 43

mm respectively polyethylene pipes are used because they have several advantages over steel they are cheaper and easier to install by unwinding from a reel have lower thermal conductivity and the specific gravity of the pipe is close to the specific gravity of the brine resulting in less suspended weight freezing brine flows at high velocity down the inner pipe it rises more slowly in the annulus absorbing heat from the surrounding rocks methods of excavation in frozen ground are similar to those under normal conditions however when explosives are used special care is required because of the proximity of the freezing holes and the potential danger of their destruction when the core of a shaft is not frozen a cactus grab or other mechanical mucker can dig the shaft bottom when the rock is partly frozen or too hard to be excavated handheld drill hammers with chisel picks are used however these working conditions are very unhealthy in hard frozen rock the drilland blast method is used with limitations the hard rock miners handbook de la vergne 2009 provides some good examples of freezing calculations and rules of thumb for ground freezing as shown in table 12 4 3 shaft drilling shaft drilling is a commonly applied technique for the development of small diameter ventilation shafts in the united states it has also been applied with moderate success in australia shaft boring uses a mechanical borer similar to a large diameter drill bit to excavate the shaft a typical 4 25 m cutterhead is illustrated in figure 12 4 9 and further details of shaft drilling technology including technical details on equipment methods of shaft drilling and some case studies are provided on the shaft drillers international web site n d zeni and maloney 2006 identify the following hazards with a conventional drill and blast shaft sinking operation falling of persons falling objects methane and other gases hoisting and communications failures equipment operating in a small work area respirable dust water inundation back injuries ignitions in the shaft concrete burns blasting escape in an emergency noise zeni and maloney also state that many of these hazards can be eliminated using blind boring of shafts because blind boring is probably safer than conventional sinking shafts in the range of 4 3 4 9 m in diameter are routinely bored using this technology with larger diameters achievable the blind boring machine is rotated from the surface using a rotary table in addition the shaft is also completely filled with a fluid such as water therefore the boring element is completely submerged the cuttings from the boring operation can be lifted from the bottom of the shaft in a number of ways however hydraulic and airlift systems are most common the most common problem associated with blind drilling is lining the shaft the most common lining for shafts fewer than 4 6 m in diameter is the use of steel linin

gs placed from the surface with concrete or grout used to bind the lining with the rock by injection in the annulus between the steel and rock other techniques do exist and are discussed by zeni and maloney 2006 these include remote slipforming concrete lightweight composite linings made from extruded fiber reinforced polymer lightweight concrete combined with steel or polymer and totally waterproof castin place concrete the future of shaft drilling depends on technological advances in fabricating and installing linings remotely which would avoid the need for humans to work in a shaft environment another good example of the method is presented by bessinger and palm 2007 for the installation of a ventilation shaft at the san juan coal mine in new mexico united states the paper particularly discusses the options available for construction of the shaft the reasons why boring was chosen and a number of the technical aspects of the project such as collar construction pilot hole drilling shaft reaming lining installation and fitting of the ventilation and escape hoist system winzing winzing is the term for small scale shaft sinking usually it is a form of internal mine development used for example to connect levels winzing has the following characteristics usually small diameter no room for mechanization although in modern systems raise boring machines can be used to develop winzes however most are still sunk by hand the possible range of sizes and depths are limited due to problems with muck removal muck removal is usually by hand shoveling use of pilot holes makes winzing easier as deviation can be avoided if levels are connected by a pilot hole ventilation is highly important but usually poorly done secure ladders are essential for access small winches are used for raising and lowering small kibbles typically a bench cut is used water can be a problem but its impact can easily be overcome by using a pilot hole all services need to be removed prior to a blast ventilation ducting water pumps and hoses compressed air lines hard physical work especially in hot conditions vacuum lift devices have been used successfully for muck removal winzing is always a slower and more expensive undertaking than anticipated in modern mining it is avoided wherever possible and raising is used in preference raising a raise is a vertical or steeply inclined opening excavated from a lower level to a higher level i e inverse winzing the inclination is such that a ladder is required for passage raises are used to gain access to ore as traveling and supply routes and as ventilation routes raises are also used as orepasses waste passes or as slot raises to provide free faces for initial blasting of certain types of stopes raises can be constructed through conventional raising cage or gig raising alimak raising longhole raising and raise boring conventional raising this method is restricted to shor

ter raises of up to 100 m in length and involves drilling and blasting using handheld drills an important requirement in this type of raising is the provision of a working platform at the face to allow drilling and charging operations to occur this platform is dismantled prior to blasting raising operations can be dangerous if correct procedures are not followed the advantages of raising include the following raising is independent of mining production so its input on current mine performance is minimal loading of muck into buckets is avoided blasting effectiveness is greater no water pumping is required the disadvantages of raising include the following the height of a raise is limited to between 100 and 120 m ladder climbing is an inconvenience for personnel and the transport of materials to the face potential hazards of rock falling from the face exist potential threats of muck jamming in the chute compartment exist surveying is difficult deviation is possible time and cost implications are involved in preparing the opening for raising purposes cage or gig raising cage or gig raising is a mechanized raising technique and is illustrated in figure 12 4 10 a pilot hole is drilled from the upper level to the lower level a hoist rope is passed down the pilot hole and attached to a cage or work platform that can be raised or lowered in the raise to the face allowing work to proceed as normal by hand methods alimak raising alimak raising is a development of cage raising using a rigid steel guide rail attached to one wall of the raise no pilot hole is required and all the controls are mounted on the cage or at the base of the raise the design of the system incorporates provision for air and water hoses drilling equipment blasting requirements and accessories alimak raising has the capability of driving longer raises at faster rates than conventional methods the machine runs on a guide rail that incorporates a pin rack the segmented guide rail is fastened to the side of the excavation with rock bolts the alimak method has the following advantages over conventional systems it allows the driving of very long raises vertical or inclined straight or curved when traveling to the head of the raise personnel are well protected in a cage under the platform the miners work from a platform that is easily adjusted for convenient height and angle risks involving gases are reduced as the raise is ventilated after blasting with an air water mixture supplied in the raise climber all equipment and material required can be taken to the head of the raise in the raise climber timbering is avoided and if needed rock bolts with or without mesh are used for support the raise climber can be used for areas up to 8 m2 the disadvantage is the cost of the raise climber system which cannot be justified for short raises therefore the most frequent application is in mines where driving raises in

hard rock is part of the mining method the australian contract mining web site 2009 provides some excellent examples of the use of the alimak climber as well as many images of alimak climbers longhole raising in this method all operations are undertaken from the top level long holes are drilled from the top level to the level at which the raise is to start holes need to be drilled accurately and can be reamed out charging is undertaken from the top level and blasting carried out in a similar manner to that used in vertical crater retreat mining figure 12 4 11 illustrates the longhole raise method raise boring raise boring is the most common method of raising in modern mining it is a safe method for the driving of long raises but it requires a high capital expenditure numerous methods of using raise borers are available such as blind hole raising pilot hole methods and so on raise boring is one of the newest technologies in shafts and raises and has matured through improvements made during the decades since its initial application large diameter holes or shafts can be drilled vertically or inclined the costs of raise boring have been substantially reduced due to the improvement of cutters raises are used as a part of multilevel mining systems for orepasses and ventilation as coal mine ventilation shafts and also as full sized shafts primarily in coal mining excavating rock by a continuous boring system has several advantages over the conventional drill and blast methods especially in vertical excavations such as shafts and raises where it eliminates the hazards associated with the presence of workers at the face similar to shaft boring the continuous action of raise boring usually ensures faster completion of the project than conventional blast and drill methods boring does not disturb the surrounding rock and as such often there is no need for support or alternatively only minimum support is required for the finished excavation a perfectly round shape makes the hole inherently stable and its smooth walls reduce resistance for ventilation as well as for the passage of ore with little prospect of hangups provided conditions are amenable the cost advantage over other systems is clearly apparent for long raises and for shafts the disadvantages of a continuous boring system include a general requirement that rock conditions be similar along the length of the raise bore another disadvantage is that drilling in very hard rock is slow and costs rapidly increase also access to upper and lower levels is necessary a raise bored rock mass is preferably dry the initial cost of the drilling rig is also high a large variety of machines are available that are designed for certain sets of rock conditions and nominal sizes and lengths of the finished hole typical characteristics of the machine specify the pilot hole diameter drill pipe diameter reaming torque type of drive ac dc or hydraulic its rpm and power h

p or kw and type of gear reducer additional specifications are the pilot hole thrust and feed rates reaming pull size of drill pipe base plate and derrick dimensions with characteristics especially important for underground application the type of transporter erector if applicable and the water and air consumption raise boring machines operate on the principle of first drilling a small pilot hole and then reaming the hole in one or more stages to the desired size the modes of operation are shown in figure 12 4 12 the most frequently used is a conventional scheme with the pilot hole down and reaming upward however in some mining systems box hole drilling upward in one stage is chosen the raise boring machine consists of several parts the rig is composed of a rigid plate and a structure on which are assembled drilling hydraulic and electrical equipment the rig has its own crawler driven by compressed air or it is mounted on rail wheels for haulage the rig can also be disassembled and transported a pilot hole is drilled through the stem with stabilizers and a conventional drilling bit stabilizers are used to ensure directional accuracy of the pilot hole several types are in common use they are of a larger diameter than drill rods with strips of tungsten carbide welded vertically or spirally on the outside bringing the outer dimension close to the hole diameter these stabilizers are used for reaming up other types for pilot holes and reaming down have more clearance allowing cuttings to pass after the pilot hole penetrates to the target opening the pilot bit and stabilizer are removed the stabilizer is mounted on the drill stem and the reamer is put in the place of the small pilot hole bit starting the hole collaring is a critical operation and damage to the cutters or bit body can occur because of the uneven surface the tension on the stem has to be adjusted to allow starting the hole without highly variable dynamic loads acting on the drilling components the stem is the longest at this stage of drilling and is therefore prone to maximum elongation and twist the design of the reamer depends on the final size of the hole and the number of reaming stages in the past the primary body of the reamer was often supplemented with stages mounted below resulting in a christmas tree shaped tool as illustrated in figure 12 4 13 because of large torque requirements this system is presently used only for shaft sized holes together with large nonrotating stabilizers in another configuration extensions are attached to the primary body by doing this an increase of the final diameter is achieved by adding more cutters in the same planes as the cutters on the primary body itself the accepted measure of drilling performance is the rate of penetration which depends primarily on bit design the cutting action has to match the rock properties to optimize the drilling rate during the reaming operation in ad

dition to the cutters the design of the reaming head also influences the raise boring performance if no other source of information is available an analysis of the wear of cutters can help in the design of an improved cutting structure independent variables when drilling are rock properties which are usually characterized by compressive tensile indirect and shearstrength measurements one of the most important technological aspects of boring is the removal of cuttings standard drilling technology with compressed air air and foam or drilling fluid is used for cuttings removal during pilot holing during the reaming operation cuttings simply fall down the hole the access point usually determines whether to drill raises from the top or bottom if only one end of the raise has adequate room the machine must be put there the extent to which raise boring will interfere with production must also be evaluated locating the borer where fewer disturbances will be caused is preferred cost is also of major concern generally drilling down and reaming up is cheaper than other configurations rock quality and its predicted response to drilling can influence the choice of direction as well in strong rocks without fractures usually no support or lining is needed if support is required shotcrete rock bolts with or without wire mesh or steel lining can be applied shotcrete can be placed remotely or by workers from a cage rock bolts must be installed manually steel lining required in unstable formations is installed remotely directly behind the reamer table 12 4 4 compares the factors that influence the selection of vertical subvertical construction methods used in mining declines a decline is an inclined tunnel from a higher level to a lower level usually but not necessarily from the surface the purpose of declines is to access ore bodies or coal seams and provide use at a later stage as haulage routes in coal mining declines are known as drifts these either provide access from the surface or from seam to seam in the latter case they are known as cross measure drifts drifts in coal mining operations can be driven at steep grades up to 1 in 2 5 40 as the predominant mode of haulage of coal is by belt conveyor which can handle such steep grades steep grades shorten the length of the drift saving costs and giving early access to the coal seam s in metalliferous mining declines also allow access and may be used for truck haulage grades of 1 in 7 14 are common however steeper grades are possible in modern mining as more powerful underground trucks continue to develop various cutoff depths for declines can be found in mining literature as shown in table 12 4 5 beyond these depths shaft systems should be employed over the years the economic cutoff depth for declines has increased for individual mines the cutoff depth of declines is a purely economic decision and should be treated as such mccarthy and livi

ngstone 2009 medhurst 2001 shows that other factors also need to be considered in the determination of a cutoff depth including advances in technology such as the development of larger underground haul trucks production rate proposed mine life potential for mine life being increased and ventilation flow tables 12 4 6 and 12 4 7 compare and contrast decline and shaft access declines are excavated and supported in a similar manner to horizontal tunnels but drainage may require more attention details of tunneling techniques applied are described in the following sections tunneling after the site of the surface connection s has been selected the sites of tunnels or drives are often limited by the position of the ore body in relation to the surface connection s if choice is available the most competent ground should be chosen as with surface connections the drive size is determined by duties such as ventilation access haulage and so on another prime consideration is the mining method as this controls the size of mining equipment required in modern mining drives smaller than 2 2 m are not common a 2 2 m drive will accommodate a small loader and allows easy passage for people medium sized drives have a cross sectional area in the range 10 25 m2 and represent the most common size range of mining drives such drives can accommodate locomotives conveyors and rope haulage systems as well as allowing ventilation flow and access from a mining viewpoint large drives lie in the range 25 40 m2 this drive size is becoming increasingly popular especially in mines employing trackless equipment and mines employing decline access this drive size range allows for the use of off highway trucks and other larger items of trackless equipment drives have a limited range of sizes due to the need to maintain a level floor for haulage and access reasons common shapes are square rectangular and arched arched roofs have better strength characteristics than square or rectangular cross sections they are used for drives where long term stability is required square and rectangular section drives are very popular in mining especially in areas of competent ground or in production areas where the drives have a limited life span two methods of excavation exist traditional drill andblast and mechanized methods which are described in subsequent sections drill and blasting is a cyclical operation whereas mechanical methods offer a more continuous system of excavation drilling and blasting the majority of mining drives are excavated by traditional drill and blast however mechanical mining is becoming much more popular outside its traditional area of coal and other soft rock mining drilling can be done by hand using air legs the number of air legs in use will depend on the drive size e g in a 2 2 m drive only one drilling machine can be accommodated in larger drives drilling is mostly undertaken using mechanized drill jumbos

with twin or three booms common the most common form of drill jumbo is electrohydraulic but other forms of power are used in dry holes anfo ammonium nitrate and fuel oil is the most commonly applied explosive whereas in wet holes explosive emulsions or slurries are increasingly being used figure 12 4 14 shows the basic cycle of operations for a typical drill and blast tunnel a good example of a typical cycle of operations is provided by argyle diamonds 2007 argyle is developing a new block cave operation to mine diamonds in australia the cycle of operations during the mine development prior to establishing caving uses typical drill and blast tunneling with one of the key design elements being that at no time does the work force work under an unsupported roof a common element in australian mining for each heading under development the cycle takes about 18 hours to complete for a nominal face advance of 4 m cycle and an average heading advance rate of 32 m week in small drives mucking may be carried out by hand shoveling into mine cars this is really only a viable method when cheap labor is available in the past rail mounted shovel loaders were used extensively these were almost universally powered by compressed air and operated as a front end loader slinging the muck over the body of the loader into mine cars located behind generally bucket capacity was limited to 0 5 m3 the real revolution in mucking operations came with the development of rubber tired loaders which lead to the development of load haul dump units lhds the advantage of these units is their mobility and flexibility bucket sizes vary from 1 to 10 m3 the size selected depends on drive dimensions and production rate required motive power is either by diesel engine or in some cases electrical power either battery or trailing cable lhds have a limited economic hauling distance range if haul distances become too large a more economic intervening haulage system such as trucks needs to be added between the lhd and the transfer point to the main mine haulage system numerous other methods of mucking also exist such as front end loaders and gathering arm loaders numerous methods of haulage exist in small drives hand tramming of mine cars may be used other commonly applied methods include the following locomotive rail haulage motive power can be provided in numerous ways such as diesel engine battery locomotive electric trolley wire systems or rope haulage lhd with or without truck shuttle cars with or without trucks conveyor belt conveyor several methods of support are possible the support required depends on local conditions the main methods of support include rock bolts cable bolts steel sets timber sets shotcrete concrete combinations of methods mechanical shields and natural support to design a drive blast the method outlined by persson et al 1994 can be employed the method can be broadly

summarized as follows determination of the burden and spacing for the cut holes to provide the relief space required such that the main blast does not freeze for both angled and parallel hole cuts determination of the burden and spacing for the main blastholes or production holes also called easers for the holes that blast downward and sideways into the cut area determination of the burden and spacing for the main blastholes or production holes also called easers for the holes that blast upward into the cut area spacing of the trimming or perimeter holes in the backs and sides of the tunnel for both conventional and or perimeter blasting spacing of the lifter or floor holes mechanized tunneling three basic rapid mechanical tunneling systems exist fullface tunnel boring mobile miners and roadheaders or boomtype machines full face tunnel boring systems full face boring systems or tunnel boring machines tbms have been in common use in civil tunneling for many years but are used less frequently in mining projects nevertheless tbms are viable for mine development some good examples of tbm application and other mechanized tunneling techniques are shown in the video the burrowers harris 1994 constant development and improved tooling have resulted in machines that are capable of advancing large diameter openings in strong igneous and metamorphic formations at rates that compete favorably and in many cases exceed conventional drill and blast methods in addition to high advance rates tbms leave a smooth profile and usually minimal ground support is required a disadvantage of tbms is their wide turning circle although a range of mini full face machines are available that have smaller turning radii the high initial cost of these machines is balanced by low running costs compared to drill and blast excavation systems full face boring machines consist of a rotating cutting head fitted with disk cutters drag bits button bits or various combinations of these advanced machines are available on which the tool type can be changed and tool spacing varied the cutting head may be an open structure with spoke like cutting arms or it may completely conceal the face except for muck removal openings and accessways for tool maintenance the open type head gives better access to the face and tools and can be used with a forepoling arrangement cutting forces are provided by the head rotation whereas normal forces are provided by the thrust of the machine against the tunnel face reaction to this thrust is provided by grippers mounted on the tbm body which in turn react against the tunnel sidewalls mucking is performed by buckets mounted on the periphery of the cutterhead and muck is removed via a central conveyor system preparations for tbm excavation typically involve construction of a portal placement of a concrete pad on which the tbm is assembled and installation of support services and equipment as shown in figure

12 4 15 careful excavation including the use of controlled blasting techniques is usually required to mine the setup area to the tight tolerances required placement of a thin concrete layer against the start up face is recommended to reduce out of balance loads tbm excavation is a continuous process with cutting mucking and support installation proceeding concurrently as the cutting head rotates it moves forward reacting against the grippers the grippers are repositioned periodically when they reach the limit of their travel the grippers are also used to steer the machine in some cases in others a floating main beam is used to steer while boring mucking in the immediate vicinity of the face is done by buckets located on the head periphery and a central conveyor system moves the muck through the body of the machine to a bridge conveyor when the rock is wet or where water inflow is a problem earth pressure balance systems are used and the muck is removed in a coagulated form in a batch process the bridge conveyor allows access for track laying and service installation without disrupting the mucking operation a variety of mucking systems can be used to haul muck to the surface but usually conveyors or shuttle trains are used adequate muck removal rates are critical to optimum face advance rates rock support requirements for hard rock tbm drives are generally minimal and estimates have suggested that the savings in rock support costs compared to drill and blast can offset the cost of the machine in as little as 7 km of tunnel hard rock tbms are commonly equipped with a partial or slotted shield and when support is required conventional rock support methods are used both soft rock and hard rock tbms can be equipped with a full shield and segmental linings installed this equipment enables hard rock tbms to cope with localized occurrences of soft ground mobile miners mobile miners are continuous hard rock mining machines a prototype robbins mobile miner was used in 1984 for the development of a 1 150 m decline at mt isa in australia advance rates of 3 66 m shift were attained mining a decline 3 66 m high 6 4 m wide section through high strength quartzite boyd 1987 roadheaders roadheaders have been used in mining and tunneling for a number of years they are known by a number of names including boom headers boom type tunneling machines and selective tunneling machines figure 12 4 16 shows commonly used terminology for the various machine components originally developed as a means of advancing roadways in underground coal mines early roadheaders were limited to cutting relatively low strength strata development of these machines has greatly extended the range of applications and they are now used in a wide variety of mining and civil tunneling work improvements in cutterhead design and the increasing use of water jet assisted cutting will result in a further extension of the range of roadheader application

s in the coming years roadheaders offer a number of advantages over tbms chiefly related to flexibility they can cut a variety of cross sections limited only by the basic dimensions of the machine and are able to cut tight curves or corners these machines also offer advantages over conventional drill and blast methods one of the most important advantages of roadheaders is the avoidance of blast damage to the rock and the consequent savings in ground support costs in addition because mechanical excavation is a continuous process shift time is more effectively used roadheader cutting assemblies consist of a cutterhead on a movable hydraulically powered boom mounted on a rotatable turret attached to a track driven chassis in addition to the cutterhead the machine also incorporates either a gathering arm or chain conveyor mucking system to remove broken rock from the face the machine may be controlled by an operator seated on the machine or located some distance away inside a shield or beneath a supported roof section to improve machine stability during cutting many machines are equipped with hydraulically powered steering rams that are used to brace the machine off the excavation sidewall two types of cutterheads are available from several manufacturers and are interchangeable on specific machines these heads are termed transverse or ripping heads and in line or milling heads in line heads rotate coaxially with the boom these heads are best suited to cutting rock with an unconfined compressive strength of 80 mpa or less in line heads require less thrust when sumping and the head shape allows greater selectivity in cutting specific beds or bands transverse heads rotate at right angles to the boom axis advances in cutter booms have resulted in machines equipped with telescopic booms or booms with extended length for cutting high backs or crowns telescopic booms are useful for cutting on steep gradients or on weak floor materials where the thrust from the machine s travel system would not be adequate roadheader face mucking systems generally consist of an apron with either a scraper chain or gathering arm system without adequate mucking capacity the face will become muck bound and excavation will be delayed in low strength formations in which high cutting rates are possible the mucking system may become the limiting factor in controlling advance rates roadheader cutting booms are generally mounted on a custom chassis with track propulsion track dimensions control ground pressures and should be carefully assessed in conditions in which the invert rocks are weak or prone to slurrying in the local area around working roadheaders high concentrations of airborne dust generated during both cutting and transport of the muck and high ambient temperatures commonly occur in addition to exceeding statutory respirable dust limits excessive dust may completely obscure the face resulting in inefficient excavation

and increased overbreak high temperatures and humidity result in labor inefficiency and overheating of electrical motors the use of water jet assisted cutting leads to a reduction in dust levels but does not entirely eliminate the problem and may actually increase humidity at the face one of the primary advantages of using a roadheader excavation system over drill and blast methods is the elimination of blast damage and the consequent savings in rock support costs all types of rock support can be adopted for use in conjunction with roadheaders because it can be relatively difficult to reverse a roadheader away from the face the machine must be covered prior to shotcrete application to prepare a deeply buried ore body for extraction by underground mining methods the mine must be developed this means driving all of the underground openings for entry into the mine as well as all of the openings needed for exploration exploitation and services for the mine many openings must be driven before the mine can begin production and many more ongoing openings must be progressed as the ore body is extracted for some block caving mines this means driving an average of as much as 25 km yr 14 mpy this chapter discusses the types of openings that must be developed the purposes for developing them what they are called and how they relate to various mining methods a development in a hard rock mine is any initial opening driven to provide an accessway anywhere in the mine it may not necessarily be in the ore such accessways usually are driven as single openings and provide space for roads ventilation power air and water lines drainage storage areas shops dump stations offices and so on they also expose the ore so that stope exploitation can begin at the points of exposure the term development implies only that there is an opening for some mine service or where production might begin in contrast in coal mining the term development refers to that phase of the mining cycle during which multiple entries are advanced through a virgin seam forming pillars by breaking crosscuts or breakthroughs between the entries when the development has reached the extent of the advance the coalextraction process removes all or part of the pillar on the retreat or a longwall panel is set up for the extraction the only area of obvious similarity between the two meanings is where the development opens more extractable ore or coal all underground mining development openings can be classified into four categories 1 to open production accessways 2 to obtain information 3 to construct service facilities 4 combinations of the first three categories for each development decisions must be made pertaining to each of the five physical variables 1 length 2 direction 3 inclination 4 size 5 method of ground control each development must serve the purpose for which it is intended and usually named predevelopment studies should include a discounte

d cash flow analysis and ensure that all elements of each alternative are included and are indeed comparable spacing and alignment of development excavations there are other general rules for recommended or preferred location and direction of mine developments the following is taken from the works of spearing 1995 and has been distilled from many years working in south african deep mines while it may be true that all of this information applies to rock under considerable stress it applies equally to weaker ground under low to moderate stress these rules are based on the theory of stress concentrations around underground openings and the interaction of those stress concentrations the usefulness of these guidelines has been proven by experience obtained underground stress interaction between excavations can obviously be controlled by an increase in the installed support but costs will also increase significantly if there is adequate available space it is generally more cost effective to limit stress interaction between excavations flat and vertical development in flat development the spacing between two square crosssectional openings is spaced horizontally at three times the combined width of the excavations and spaced vertically at three times the width of the smaller excavation provided the area of the larger excavation is less than four times the area of the smaller the spacing between a rectangular cross section is spaced horizontally at three times the combined maximum dimensions of the excavations and spaced vertically at three times the maximum dimension of the smaller excavation provided the height to width ratio of either excavation does not exceed either 2 1 or 1 2 the spacing between two circular cross sectional openings is spaced horizontally at three times the diameter of the larger excavation and spaced vertically at three times the diameter of the smaller excavation provided the area of the excavation is less than four times the area of the smaller in vertical development e g shafts and raises the spacing of a square cross section is three times the combined widths of the excavation a rectangular cross section is three times the combined diagonal dimensions of the excavations and a circular cross section is three times the diameter of the larger excavation successive turnouts breakaways multiple turnouts from a given intersection should be avoided to reduce dangerously large roof spans at intersections the rule is that turnouts should be spaced at six times the width of the excavation between successive tangent points this rule does not apply to room and pillar stoping in this case the roof spans are safely designed acute turnouts less than 45 should be avoided because these result in pointed bullnoses which are unstable fracture and failure of a bullnose result in large unsupported hanging wall spans the breakaways for inclines are often brought too close to the conne

cting crosscut the length of the connection between an incline and the flat drifts should be three times the diagonal dimension of the flat end orientation of adjacent excavation developments the most highly stressed part of an excavation is the corner therefore positioning a development such as a decline in an unfavorable orientation too close to the corner of an adjacent opening should be avoided development influenced by geology in all cases the geology of the area should be taken into consideration in planning the development known weak geological horizons should be avoided even at the expense of longer developments in permanent excavations sumps settlers hoist chambers etc the position of faults and the orientation of joint sets are critical to the stability of the development layouts must therefore cater to such geological features all excavations should be kept away from high stress dikes where possible when a dike is traversed the most direct route should be used breakaways should not be sited in dikes even at the expense of extra development haulages drift opening should not be positioned in fault zones and development should not occur alongside a fault a fault should always be intersected at an angle as near to 90 as possible development influenced by stress condition when haulages are positioned beneath mined out areas consideration should be given to the 45 destressing guideline an overstoping angle of 45 is generally required to destress a haulage at an angle greater than 45 stress concentrations are higher this rule applies to reefs of 0 to 20 dips greater than this often call for computer modeling to show the extent of overstoping needed haulages should not be laid out too close to remnant pillars which are very highly stressed another case of this application is where the production level for block caving is developed beneath and well ahead of the extraction level to avoid the high stress conditions that will come once the extraction level is opened opening production accessways initial entry into the underground mine figure 12 5 1 shows various types of entries into an underground mine the initial entry may be for further exploration a test mine production openings or eventual service openings the opening may be a drift entry adit into the side of a hill a decline slope a vertical shaft an incline shaft or a multiplestage hoisting shaft the type of opening will depend on many factors the purpose of the opening the elevation of the terrain at the surface of the mine in relation to the elevation of the ore body the type of ore body or shape of the ore body how material workers and ore waste will be moved in and out of the mine the depth needed to eventually develop openings and the mining method that will be used to exploit the ore each of the types of initial openings is discussed in the following subsections drift entry adit an adit or drift entry is the mos

t economical approach when the extractable material is above the floor elevation of a nearby valley literally thousands of appalachian room and pillar coal mines have been opened this way to mine coal seams exposed on the hillsides these are known as punch mines most of the early western u s hard rock mines used this type of development as did most of the limestone and dolomite mines throughout the missouri mississippi river basin the exposed rock could be opened through horizontal adits into the sides of the mountain bluffs a few of the advantages of the drift entry are mine drainage by gravity level or downhill haulage for mined material easy access for equipment in and out of the mine and rail haulage can be used if desired adits are usually driven by drill and blast methods using trackless jumbos rail mounted jumbos jackleg handheld drills or with a mechanical excavation machine such as a tunnel boring machine tbm or roadheaders the grade of the adit should be driven at a 2 to 3 grade for good drainage but probably limited to 1 to 2 if rail haulage is to be used almost never should a drift be driven by drill andblast methods on a flat grade if the adit goes any significant length because there is no way to drain it unless the water rises enough to top the roadbed for nearly flat grades laserguided equipment should be used and if there is a lot of water it will likely be necessary to overexcavate the bottom build a flume type ditch u shaped on the side and build a roadbed up to the grade that is needed when the mineable material goes to a depth lower than the surrounding terrain either a slope decline or a shaft must be developed many mines start out as adit mines and then a shaft is sunk or a decline is driven to exploit the lower portions of the ore body the ore is then brought to the level of the adit where it is hauled out of the mine the same is true for ore lying above the adit the shaft or ramp will be excavated above the adit and the material will be dropped down to adit level for haulage to the surface some refer to these as upside down mines a few examples can be found at the pine creek tungsten mine bishop california united states the el soldado copper mine chile and the stillwater and east boulder platinum mines montana united states decline slope entry again there is a difference between what this development is called in hard rock mining and coal mining coal mines term such development as slopes and these slopes usually contain a production conveyor belt hard rock miners call them declines and unless it is a very high production mine more than about 4 540 t d 5 000 tpd they probably will not contain a conveyor belt slopes declines are usually limited to relatively shallow mines because for a given vertical depth declines require approximately five to six times the linear distance compared to a shaft thus to reach a depth of only 460 m 1 500 ft

an 18 slope would be approximately 2 540 m 8 333 ft long the declines will normally be driven by using drill jumbos and trackless loaders and trucks roof bolting and shotcrete will normally be used for support it is common practice to drive muck bays at regular intervals as the decline is being driven so the blasted round can be moved very quickly from the face to the muck bays leaving the face available for another drill round sooner the development muck can then be transferred to the surface while the next round is being drilled and charged the usual distance between muck bays is about 120 180 m 400 600 ft declines can in theory be excavated by tbms but rarely are there is an ongoing debate regarding what the proper decline grade for truck haulage should be however it depends on the length of the haul power of the truck s engine and drivetrain versus the size of the load and the surface condition of the decline normally for good efficient haulage the truck should be able to maintain a speed of 12 13 km h 7 8 mph hauling a full load up the ramp as an example hauling a 48 t 53 ton load up a straight well maintained road surface at 15 grade a truck with a 485 w 650 hp engine should be able to maintain 12 km h 7 mph for the full length of a 4 km 2 4 mi decline many operators try to drive haulage declines at 18 20 to save initial cost with a shorter decline but this usually increases the operating cost by increasing the cycle time of the haul this can be remedied by employing engines and drivetrains that are much more robust than those used for normal mine haulage many suppliers of haulage equipment today can supply the needed drivetrain for very steep grades even for large haul trucks gradients steeper than 15 are not practical for main production areas due to difficulties maintaining the road surface to an adequate standard the road surface tends to undulate and creep particularly when wet there are also issues with accelerated transmission wear on trucks the slope driven for conveyor belt haulage for most crushed ore and coal can be up to 30 however if rubber tired mobile equipment must also operate on the decline this may limit the grade to less than 20 for safe operation the advantages and disadvantages of declines are discussed later in this chapter see the slopes versus shafts section vertical shaft entry in north america for most deep hard rock mines over 450 m 1 500 ft deep vertical shafts are still the preferred method of opening the mine for production the purposes of vertical shafts are distributed as production 22 exploration and test mine development 18 service including ventilation 27 and production and service combination 33 dravo corporation 1974 these proportions have probably not changed since the 1980s for north america however in australia where declines are used more than shafts for production the primary use for vertical shafts is ventila

tion basic design considerations like all underground openings vertical shafts demand considerable study planning and design indeed this becomes even more important with shafts the detailed engineering includes seven specific areas that must be considered 1 size in the united states shaft size varies from only 1 2 m 4 ft to 8 5 m 28 ft depending primarily on the purpose of the shaft 2 configuration while most new shafts that are sunk from the surface today are circular many older shafts and those sunk from underground are rectangular 3 ground support for deepening old shafts timber supports are still used however semimonolithic concrete lining is the most important shaft lining used today concrete linings have many advantages good bond between the lining and the rock decreased labor intensity for permanent linings high strength resisting rock movement in some cases complete mechanization of lining placement the disadvantages of concrete linings are the inability to take load immediately after placement lower resistance to some corrosive waters and higher sensitivity to mass rock movement and more difficult to repair 4 internal arrangement of services various compartments for hoisting ore and waste which may or may not include timber or steel conveyance guides workers materials and utility services all have to be designed sometimes the utilities are incased within pipes embedded in the concrete lining 5 depth length of shaft for hoisting ore and waste rock the shaft needs to be sunk to a depth where there is adequate room below the haulage level for a large storage bin orepass between the rock dump and the feeder system that feeds the crusher the crusher system feeding the crushed material into the skip loading system accommodating the length of the skips and accommodating some sort of shaft spill rock cleanout equipment the practical depth for shafts sunk for hoisting in one pass is about 1 220 1 500 m 4 000 5 000 ft beyond that depth an underground hoisting and ore transfer system should be considered see figure 12 5 1 the notable exception to this is what is being done in south africa the constraints on the unique deep gold mines in south africa have led to improvements in the development of deep shaft hoisting these followed an industry research and development program undertaken in the last decade changes in legislation winder control and operating practice has allowed the depth limits of singlelift hoisting to be extended beyond a previous practical limit of about 2 400 m 7 900 ft several shafts have been commissioned where the winding depth is beyond 3 000 m 9 850 ft 6 shaft linings concrete lining strength is usually between 19 and 33 mpa 2 800 and 4 200 psi but may be as high as 50 mpa 7 200 psi type iii cement is usually used where the concrete is mixed on the surface and can be placed either by dropping batches down the shaft in a conta

iner and then distributing to the forms shutterings or by using slick lines and pots that distribute the concrete to the forms in either case the concrete must be compacted in place with pneumatic vibrators usually the shutterings are collapsible steel forms that can be quickly erected and torn loose curb ring forms are placed and poured first then the shuttering forms are placed on top of them shotcrete linings can be placed in a few variations depending on the shaft wall stability the shotcrete may be applied directly on the wall by itself it can be sprayed over woven wire fabric or the walls can be rock bolted before or after the shotcrete is placed some operators prefer using steel fibers in the wet shotcrete for the added yield strength rather than shotcreting over woven wire fabric shotcreting of very small shafts is particularly desirable where steel forms are difficult to work with using shotcrete over wire mesh is also fairly common in european coal mines for internal shafts where the permanent lining is only roof bolts wire mesh and shotcrete 7 shaft collars all hoisting shafts require a massive block of reinforced concrete designed to anchor the headframe and secure the top portion of the shaft in the unconsolidated material through which the shaft collar was excavated they must be designed to allow all of the mine utilities to pass through the collar into the shaft basic shaft sinking equipment most deep shafts are sunk using a multistage platform that is installed in the shaft and stays in the shaft until it is completed some call this platform a galloway stage the drill jumbos can pass through the opening in the stage the concrete forms can be stored and assembled on the stage and mucking out the bottom of the shaft can take place with the stage in place while buckets kibbles can pass through the opening most shafts use one of the following mucking lashing systems cactus type air operated grab pneumatic or hydraulic cryderman mucker backhoe type machine riddle clamshell type machine crawler type overshot loader also the first portion of any shaft the presink may be excavated with a clamshell or cactus grab bucket operated from a surface crane before the galloway stage is placed in the shaft the shaft sinking schedule using a multistage platform is shown in figure 12 5 2 shaft sinking is normally performed by specialized contractors another concept is blind boring or upreaming the shafts literally hundreds of blind bored shafts have been completed in the united states and elsewhere around the world slopes versus shafts no general rule governs whether a shaft or a slope is the proper type of entry to be developed each has advantages and disadvantages every major development decision should include a complete trade off study combined with an accurate discounted cash flow analysis that considers net present value npv and internal rate of return ror on capital investment alter

natives although the cost per unit distance for a decline may be somewhat lower than for a vertical shaft the overall cost usually is higher and the development time is usually longer as a result of the greater length required thus the cost per vertical meter foot for the development except for very shallow depths e g 150 m 500 ft is considerably higher for slopes than shafts figure 12 5 3 compares relative costs per vertical unit of distance with total depth for declines and shafts this chart was taken from work done many years ago on u s mines that produced between 1 6 mt a and 6 6 mt a 1 8 million tpy and 7 3 million tpy and while the actual cost is no longer accurate the relative costs are still approximately comparable declines slopes have specific advantages over shafts the technique of driving a decline does not require as much specialized technique as sinking a shaft declines can be used for trackless track or conveyor belt haulage when used for conveyor belt haulage the tonnage can be very large the ground support cost for roof bolts and shotcrete is usually less expensive than the concrete liner normally required in a shaft large equipment can be moved in and out of the mine without disassembly maintenance is easier and less costly if the production is from a fairly shallow depth and greater than about 5 000 t d 5 500 tpd the decline with a conveyor is likely to be lower cost no visible head gear is required i e no visual pollution no winding gear is required which in a remote location may need to be transported considerable distances along an unsuitable road network no surface winding facilities need to be constructed maintained or manned for major wrecks within the opening shaft or decline there are fewer production delays in declines for an inclined ore body and if only the top portion is initially developed declines incur a much lower initial cost than a shaft that would go to the depth to mine the entire ore body the disadvantages of declines over shafts are for an equivalent depth the slope will be four to six times longer the total cost and the time to construct for very deep declines will be greater but for inclined ore bodies production may be able to start early in the upper ore body if there is a thick very weak zone of rock to penetrate the shorter distance of the shaft to penetrate through this zone will probably be much easier and less costly than for a decline where substantial ground support may be required if the development opening is to be used for ventilation the length of the opening will be increased four to six times and the wall rock will be rougher than the concretelined shaft both of these issues will require more power to overcome the increased resistance for the same size development opening from depths of about 250 to 350 m 800 to 1 100 ft mines producing from 3 3 to 6 6 mt a 3 6 to 7 3 million tpy have been dev