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remove some or all pillars placing cemented rock fill around pillars all the way to the back to properly support the stope between the solid and the backfill creates a pressure arch over the intervening pillars which can then be removed economics permitting the area can be backfilled if necessary and the pillars that were encapsulated or trapped in the original fill can be mined from a sublevel beneath the pillar by long hole blasting of the pillar into the sublevel area lane et al 1999 the total backfill prevents any future subsidence all of the previous methods except sublevel extraction were used in the final mining of the old lead belt of missouri in its last 25 years and much more intensely in its last 10 years this r p mining district was mined for 110 years before finally shutting down sometimes when pillars are partially removed without proper planning or when pillars that are too small are left after first pass mining the pillars begin to break up and serious convergence accelerates uncontrollably one of two things must be done to save the area 1 massive pillar reinforcement if there is time before collapse or 2 massive backfill in the entire area among examples of massive pillar reinforcement fully grouted rebars were placed in more than 300 pillars in the one of the r p mines of the viburnum trend weakly 1982 has documented the method employed reinforcing pattern cement grout mixture convergence instrumentation and results in the old lead belt areas pillars to be reinforced were wrapped with used hoist cable with a load of 5 5 t 6 st placed on each wrap wycoff 1950 but this reinforcement method is not as fast economical or effective as use of fully grouted rebars in the pillars among examples of massive backfill are the leadwood mines of the old lead belt where this method was used only about 132 m 425 ft below surface however the back was thin bedded dolomite interbedded with shale and glauconite and badly fractured and leached the roof bolt was originally developed in these st joe mines in the late 1920s as a means of tying the layers together to act as a beam weigel 1943 casteel 1964 the roof bolts were used with channel irons to form a crude truss even though the rock in the pillars provided poor support because it contained bands of high grade galena for economic reasons pillar removal and slabbing took place over a period of 25 years occasionally local cave ins would occur after an area had been pillared but since the cave ins were beneath uninhabited st joe owned land they were of no real concern however when slabbed pillars between two smaller cave ins began to fail and a third and fourth cave in occurred in more critical areas there was a considerable amount of concern initial extraction of some of the area involved multiple pass mining and room heights were mostly 6 1 to 12 2 m 20 to 40 ft final mining of the area had resulted in ore extraction of approx

imately 95 to stop caving in the third and fourth areas in around 1962 more than a million tons of uncemented cycloned sand tailings were put into the mines filling the rooms nearly to the back the backfill was very successful in controlling the converging failing ground the end result of use of these methods is comparable to the overhand cut and fill practice of deliberately mining pillars very small and immediately filling in around them this method known as postpillar mining was used in the falconbridge nickel mines canada cleland and singh 1973 the elliot lake uranium mines canada hedley and grant 1972 and in some mines in san martin and niaca mexico and san vicente peru the end results for the two mining methods with very small pillars encapsulated in sand tailings are similar ventilation most of the literature on ventilation design for r p mining is written for coal mines however a few considerations are unique to metal and aggregate r p mines r p metal or aggregate mines differ from r p coal mines in the following respects everything is larger it is not uncommon to have entry drifts of 9 9 m 30 30 ft and rooms stoped out to 12 15 m 40 50 ft a lot more air or a stream of high pressure directed air is required to meet the required minimum velocity across the working face stopings are difficult to build and the total force against them can be enormous many r p mines rely on auxiliary fans to pick up air from the main ventilation drifts and carry enough air through vent tubing to serve the needs of the active face ventilation doors are similar in design to airplane hangar doors again the total force against these doors which must be controlled automatically is enormous air stratification in large stopes can be a problem diesel equipment is used extensively and the exhaust must be diluted to meet particulate standards ventilation fans can be placed underground if it is beneficial to do so the effect of changing market conditions on r p mine planning market conditions are important in planning for r p mining even during the stages of feasibility planning an advantage of r p mining for metal or any commodity that is gradational in value is that mining methods and approaches are flexible and can be changed quickly in response to changing market conditions more quickly in fact than is possible for other mining systems another advantage of r p mining is that if the ore body is continuous new faces are continuously being opened a stope only four pillars wide can have as many as 15 faces open simultaneously for drilling and blasting imagine how many faces would be exposed to drilling and blasting if a stope were 10 pillars wide for metal mines this offers a great deal of flexibility one can mine the grade of ore that is most desirable at the current market price for at least short periods of time in each stope unit it is usually possible to work only the higher grade

or lower grade faces of a particular mineral although doing so usually has a drastic effect on the grade within a few days during initial mining of the high grade fletcher mine in southeastern missouri united states after operation for only 3 to 4 years approximately 50 to 70 faces were open for mining on any given day although only 10 to 12 were actually worked one can then select the faces to be mined so as to maintain the grade of lead zinc or copper ore that can best be handled by the concentrator and still optimize the financial objective of the mine spare equipment can be put into reserve stopes to increase production if the remaining materials flow can handle the added capacity however if this is done too often or for too long mine development must also be accelerated old stopes if maintained can be reactivated quickly for mining of lower grade minerals that have become financially worth mining lower grade ore can be left in the floor or the back of stopes until price rises at which point the reserves are readily available for quick mining this technique is often overlooked by those unaccustomed to planning r p metal mining where mineral values are gradational the optimal technique is first to mine the better areas of the mineral reserve and maintain a grade of ore that satisfies the economic objectives at that time then when economics change the lower grade areas left as remnant ore reserves can be mined in spite of these considerations however even in r p mining drastic changes in the rate of mining momentum cannot be assumed to be free it often takes several months with an increased labor force to regain a production level that seemed easy to maintain before a cutback in mine production if spare equipment is used to increase production maintenance will probably convert from a previous preventive maintenance onshift schedule to a breakdown overtime schedule at least until permanent additional equipment can be obtained nevertheless the necessary changes can be made as discussed in the section on pillar robbing another method for achieving economic flexibility with r p mining is to slab or remove high grade pillars then even in the later years of the mining operation some sweetener is left to blend with the lower grade although not unique to r p mining this technique is easier to accomplish in an r p operation than in other more complex mining systems capital and operating costs for hard rock r p mining capital and operating costs for hard rock r p mining are usually considerably less than for other hard rock mining systems at the same daily production however panel caving is not comparable because r p mines are not normally built to produce the equivalent lower tonnage per day table 13 1 2 shows estimated preliminary capital costs for r p mines the original estimates stebbins and schumacher 2001 based on 2000 data included cost contingencies of 7 2 to 7 6 the contingencies were remo

ved the percentage of the remaining total was calculated for each element of capital cost elemental capital costs were escalated using the proper indexes for the time period and totaled and a 25 contingency was added to the total these costs include all equipment preproduction development mine surface facilities engineering and project management working capital and cost contingencies table 13 1 3 shows estimated direct operating costs for r p mines the estimates are based on 1998 data from six highly productive r p mines that vary in daily production from about 2 300 to 6 300 t 2 500 to 7 000 st per day the estimates were escalated to 2008 using the proper mining unit inflation index for each cost element and a 20 contingency was added it is noteworthy that the six mines have very large rooms are highly mechanized and are considerably more productive in tons per worker shift than is average for the industry the latter is about 40 50 t 44 55 st per worker shift early mines using the r p method were developed more or less at random with pillar sizes determined empirically and headings driven in whatever direction was convenient for example in australia in the early days the rooms were laid out systematically to suit hand shoveling and endless rope haulage in modern coal mines however rooms and pillars are typically laid out orthogonally in a regulated pattern figure 13 2 1 in addition to support coal pillars are also expected to restrict surface movement protect critical airways from the effect of high load distribution and abutment stresses form ventilation partitions for airflow distribution act as yield sacrificial pillars to provide temporary support in high lateralstress situations and as shaft pillars to protect air shafts and other critical mine structures barrier pillars can be used to form a load bearing framework to minimize the impact of ground movement from adjacent workings to isolate major faults and local joints to protect critical roadways to separate adjacent mine leases and most importantly to provide a barrier to large masses of water filled old workings r p mining has two basic operations entry development and coal production the former is to develop mine entries to access coal seams and construct the necessary support infrastructure to facilitate coal production although in inclined coal seams it might be necessary to excavate either above or below strata to access the coal seam most entries are in the coal seam itself which also produces coal in the process both horizontal and inclined r p operations require other support operations such as roof bolting ventilation face cleaning and transportation the mine layout has to be carefully planned to accommodate all functions for maximum efficiency development openings entries and production entries rooms closely resemble each other both are driven parallel and in multiples and connected by crosscuts to form pillars d

riving multiple entries simultaneously is necessary for high production and operational efficiency it also facilitates ventilation roof bolting and coal transportation routes in the face area usually a hierarchy of entries exists in underground coal mines main entries are driven mainly to divide the property into major areas and usually serve the life of the mine for ventilation as well as for worker and coal transport submain entries can be regarded as feeders from the mains that subdivide each major area from the submains panel entries take off to subdivide a block of coal into panels and rooms for orderly coal extraction it can be seen in figure 13 2 2 that the east main has a 3 2 2 or intake neutral return arrangement and the submain off the east main has the same arrangement intake neutral return but in 2 2 1 layout the number of intake entries exceeds the number of return entries indicating it is a blowing system with a fan on top of an intake air shaft where neutral air travels away from working faces the number of development entries in any area can vary between three and seven or as many as eleven in some cases the exact number is often a balance of the following factors expected load on the pillar due to its depth or other geotechnical concerns coal production the system that supports the expected production number of pieces of mining equipment that are operating labor and material transport ventilation etc it is critical that a proper balance be achieved between the optimum coal extraction ratio and the size of pillar necessary to support the overburden geotechnical design for r p mining because of the complicated and varying nature of rock strata it is customary to start the basic pillar design by assuming 1 rock strata to be homogeneous uniform and horizontally bedded deposits 2 a uniform pillar layout and 3 the rock materials are competent other factors such as pillar shape uniformity of distribution and cross section and height width ratio true pillar geometry outside dimensions minus an outer shell of distressed materials that in effect support very little load and true mechanical behavior of pillar materials also have to be considered continuous mining method with a continuous mining method separate unit operations of drilling cutting blasting and loading are replaced by a single high performance continuous miner cm as the steel drum cutterhead spins in a circle the tungsten carbide steel teeth or cutting bits break the coal from the face the coal falls onto a steel apron where moving arms load the coal onto a short conveyor on the back of the cm and transport it to a shuttle car waiting at the rear figure 13 2 5 the shuttle car is driven to a feeder breaker three to four breaks 180 280 ft depending on pillar sizes away and loads the panel conveyor which carries the coal to the surface after one working face is completed the cm then moves to the neighboring fa

ce and the process is repeated a typical cm measures approximately 33 11 4 5 ft length width height respectively the depth of each cut in a working face has increased today to 30 40 ft deep cut because of more powerful and reliable machines but this requires multiple passes and resetting of the miner to complete cutting in one face efficiency of the miner operator and coordination of other auxiliary functions such as face cleaning establishing ventilation and roof bolting are all key factors in determining the face advance rate and productivity remotecontrolled cms are now the norm in the industry and they have significantly enhanced face productivity and safety a typical standard continuous mining section uses one set of equipment with one crew operating in a panel of three to seven entries the section is equipped with one cm two to three shuttle cars or ram cars one roof bolter and a section scoop for auxiliary jobs in addition to a section foreman a cm section would have nine to ten workers two cm operators one operating and one assisting relieving two roof bolters three shuttle car operators one mechanic and one general laborer in addition to the standard unit u s coal mines have also been using either a double or a super unit arrangement the former uses two sets of equipment and two or nearly two crews operating side by side in a panel with 8 13 entries the latter uses two sets of equipment and one crew plus two extra workers operating side by side in a panel with 4 13 entries depending on the mine s specific condition either method has been used successfully in developing long distance mine mains with multiple entries conventional mining drilling and blasting method conventional mining is cyclical employing mobile mechanized mining equipment to carry out production unit operations in five distinct steps production cycle cut drill blast load haul a cutting machine cuts a 5 7 in slot or kerf at about 10 ft in length to improve breaking during blasting a drill is then used to drill holes into the coal face in a predetermined pattern which are then loaded with permissible explosives after the coal is shot and fragmented by a qualified shot firer following strict safe blasting procedures it is scooped up and loaded into a waiting shuttle car that transports the coal to a conveyor belt loading head feeder breaker for transport to the surface in some smaller mines the coal is hauled in rubber tired trailers drawn by battery powered tractors after the first working area is completed depending on specific practice this could be the first heading on the right in a panel if the mine works from right to left direction the crew moves to the adjacent workplace and the steps repeat as shown in figure 13 2 6 coal mined at the face is transported using a shuttle car or ram car to the breaker where it is transferred to the conveyor belt several other auxiliary functions must als

o accompany production functions face cleaning roof control proper face ventilation cables and equipment advances and movement of materials and supplies as a result multiple entries are necessary such that all functions can be performed at separate faces without delays depending on the face advance rate and traveling distance as well as the change out points due to multiple hauling units the location of the dumping point breaker needs to be located in a position that minimizes haulage travel time and congestion in the united states haulage entries where coal is transported are separated from both intake and return airways thus the introduction of neutral entry airways which should not be used in the face area but be channeled directly into the return airway increased safety concerns efficiency demand and the need to operate several pieces of mining equipment simultaneously have caused this conventional mining method to fall out of favor and it is slowly being replaced by the continuous mining method in 2008 only 0 6 of all coal mined underground in the united states was mined using this method key auxiliary functions to provide a highly productive mining environment while maintaining safety a well coordinated mining plan that incorporates several key support functions is critical the top two such functions are effective ground control and proper face ventilation ground control in coal mining ground control is concerned with both the short and longterm stability of mine accesses and entries as well as subsidence control although no method of roof control yet devised has proven to be universally acceptable for the wide range of strata conditions perhaps the most significant development in coal mine ground control during the 20th century was the introduction of roof bolting in the late 1940s and 1950s figure 13 2 7 being able to provide effective and efficient ground control is essential to ensure safety and productivity in underground coal mines the theory of the interaction between roof bolts and the rock strata is complex and several models are available beam theory roof bolts bind together several weak strata into one stronger beam suspension theory weak members of the strata are suspended from a strong anchor horizon and the keying effect theory roof bolts act much like the keystone in an arch all seem to be able to explain at least part of the support phenomenon under certain conditions from an engineering standpoint roof bolts are inherently more effective than the wooden timbers they replaced they greatly accelerated the transition to trackless rubber tired face haulage and they significantly facilitated coal production in underground coal mining steel bolts usually 4 6 ft long and 0 6 1 in in diameter are inserted in holes drilled into the roof strata by a roof bolter and are secured as primary roof support by friction or resin various types of bolt systems are commercially available

frictional tensioned nontensioned fully grouted tensioned and accepted by the united states and world mining industry the tensioned bolt system also referred to as an active system includes the conventional point anchored mechanical bolt a tensioned combination bolt partially grouted tensioned rebar bolt mechanically anchored resin assisted bolt partially grouted cable bolt and so on the nontensioned passive bolts include the fully grouted resin bolt and resin anchored cable bolt the fully grouted tensioned bolt combines the advantages of both the tensioned and nontensioned bolt roof channel roof mat roof pan wire mesh and shotcrete have also been used in combination with roof bolts for surface restraint they are used to support small pieces of loose rock or as reinforcement for shotcrete and the latter to provide support to the rock surface between bolts shotcrete provides the added advantages of reducing the rock exposure thereby minimize weathering effects today all major mining countries have mandatory safety and ground control standards for underground coal mines in the united states specific regulations on roof support standards are set forth in subpart c title 30 part 75 of the code of federal regulations cfr 2010 regulations stipulate that no person shall work or travel under unsupported roof underground and the roof face and ribs of areas where persons work or travel shall be supported or otherwise controlled to protect persons from hazards related to falls of the roof face or ribs and coal or rock bursts no person shall work or travel under unsupported roof additionally there are detailed regulations dealing with roof bolting including installations conventional roof support pillar recovery and automated temporary roof support systems additional supporting systems either temporary or permanent in nature to meet specific geological conditions are commonly used wooden cribs or posts trusses cable bolt steel set and wire mesh or shotcrete have all been successfully applied in underground coal mines wooden cribs figure 13 2 8 are prop timbers or ties laid in alternative cross layers i e log cabin style to effectively provide significant reinforcement capacity over a large deflection range however crib support systems restrict airflow generate floor heave and permit large roof deformations that can lead to failure in other words they have to be used carefully roof truss as a means of secondary or supplemental mine roof support has been receiving increasing acceptance in recent years particularly in bituminous coal mines in those entries where long life is required or doubtful performance from other forms of support seems likely trusses in many cases have provided the necessary protection for continuing operations steel trusses employ a combination of steel rods anchored into the roof to create zones of compression and provide better support for weak roofs and roofs

over wide areas the aforementioned roof bolts used to reinforce the roofs of mines although effective are relatively expensive and by definition inflexible cable bolts use strand cable coupled with properly designed couplings these are not only more economical they provide relative flexibility as compared to the steel bar this support system has been recognized as an effective ground control system in underground coal mines steel sets are also commonly used as long term support of underground openings such as inclined mine entries main travelways belt entries and so on the operator generally connects the steel set with tie rods inserts wood blocks between the arch sets as lagging and backfills the voids between the steel set and roof under some extreme geological conditions yieldable arches are used to provide effective ground support in areas of excessive ground movement or faulted or fractured rock they are designed such that when the ground load exceeds the design load of the arch as installed yielding takes place in the joints of the arch permitting the load to settle into a natural arch of its own and thus bring all forces into equilibrium the arch is often stronger after yielding than before due to increased joint overlap ventilation proper ventilation is critical in any underground coal operation because of the presence of methane and coal dust as well as diesel particulate matters from diesel powered equipment control of these hazardous pollutants has become increasingly challenging because of increasing mine sizes coal production and the use of diesel powered mining equipment in recent years other essential auxiliary services other auxiliary operations are essential to the performance of an operating unit coal haulage power supplies pumping materials and supplies stopping lines and overcast construction a cm is not truly continuous but often only operates for less than half of the time with the other half either employed for tramming from location to location or waiting for haulage vehicles to unload coal different methods have been used to minimize this nonproductive time in order to realize cms true production potential in the face area the proper number of haulage vehicles for the face and an optimum distance between the face and the feeder breaker are the most commonly used measures to minimize a cm s waiting time face haulage uses either an electric powered vehicle with an umbilical cable rubber tired shuttle cars or battery powered ram cars today it is common to use three haulage vehicles with well planned travel routes to maximize efficiency a compromise between optimum coal haulage and minimum congestion flexible conveyor trains fcts have also been used in some mines to minimize the cm s waiting time they are electric powered self propelled conveyor systems that provide continuous haulage of material from a cm to the main mine belt the fct uses a rubber belt similar to a standard fix

ed conveyor the fct s conveyor belt operates independently from the track chain propulsion system allowing the fct to move and convey material simultaneously available in lengths of up to 570 ft the fct is able to negotiate multiple 90 turns in an underground mine infrastructure thereby greatly facilitating face haulage in addition to face haulage proper planning is also critical to ensure an efficient haulage system to transport coal from the face area to the panel belt and ultimately to the surface through panel and main conveyor belts to prevent explosions from settled coal dust on the floor rib and roof areas of mine entries are covered with pulverized limestone inert dust or rock dust studies show that during an explosion the rock dust disperses mixes with the coal dust and prevents flame propagation by acting as a thermal or heat sink i e the rock dust reduces the flame temperature to the point where combustion of the coal particles can no longer occur under u s federal law cfr 2010 all underground areas of a coal mine except those areas in which the dust is too wet or too high in incombustible content to propagate an explosion must be properly rock dusted to within 40 ft of all working faces rock dusting is usually carried on a third or idle shift to not interfere with routine production in coal methane drainage is still an art but some of the drilling technology e g directional drilling from both surface and underground borrowed from the petroleum industry has been practiced at some gassy mines since the mid 1990s and has obtained quite satisfactory results its performance is still continually being assessed pillar extraction only after careful review of pillar performance to date and pillar extraction plans and upon approval from regulatory agencies can coal pillars can be systematically removed after primary mining to maximize coal production some of the commonly used extraction methods are the christmas tree or split and fender extraction methods the former is favored because it does not require place changes and bolting a third system is the pocket and wing method which can fully extract large pillars the general procedure is to extract one row of pillars at a time leaving the mined out portion or gob free to subside whereas extraction of all the coal in a pillar is a desirable objective partial pillar extraction schemes are more common these are usually accomplished by cms or hydraulic mining with augers driving into pillars according to predetermined extraction plans figure 13 2 10 pillar extraction plans must be properly designed according to site specific geotechnical and operational parameters equipment and timber availability and cost pillar dimensions major geological features structures roof competency seam thickness extraction span or cavability gob edge and depth of cover to ensure pillar stability during extraction case histories attributed to roof falls h

ave found that both weak rock and competent immediate roof strata cause pillar failures hazards associated with pillar extraction also tend to intensify with depth failures include both squeeze nonviolent gradual pillar failures and bumps both of which are generally more severe at depth design of pillar extraction schemes requires geotechnical input from specialists system rationalization and optimization traditionally coal productivity has always been measured by tons per workshift or similar measures based on total tonnage other factors such as size distribution dust generation and waste management also need to be considered at the planning stage to meet specific contractual demands and processing capacities at the surface preparation plant in the face area cm cutting loading and tramming capabilities miner coal haulage face layout and so on must all be carefully planned and matched in order to achieve optimal overall efficiency materials handing still represents a significant component of the operational cost a study on coal mining production costs in the united states chugh et al 2002 shows that of the total production cost labor with benefits constitutes 47 mining related costs supplies maintenance power and other direct costs uses another 29 5 with the remaining designated to depreciation taxes and other expenses of this total production cost half goes to face production outby construction uses another 25 coal processing and waste management uses 15 and the remaining 10 is allocated to transportation the same study indicates that face haulage is still the major bottleneck to productivity despite the use of more powerful high voltage cms the study compares a joy 14cm15 miner 995 v with a high voltage 14cm27 miner 2 300 v in the illinois coalfields results show that the high voltage miner outperforms its counterpart in cutting and loading rate by 29 89 tons per car by 8 64 and tramming rate 7 84 but they perform equally in tons per shift true productivity gains of the high voltage miner were not fully realized because of the batch haulage system delays in most countries the major impediment to productivity is still the time required to install roof supports summary with its low capital investment and great flexibility r p mining continues to be an indispensible method in underground coal mining in some cases it is the only method to access remnant parcels of coal that would otherwise be abandoned using longwall mining it is also essential in gate road development in longwall mining and under special circumstances with proper planning it may in cost terms even compete with longwall mining although the basic methods of operation in r p mining remain similar there are a variety of systems with minor variations from location to location depending on local specifics it is all a matter of system optimization what would be an optimal number of entries or haulage method

in one location may not be so in a different location depending on advance rates ground control concerns and dust and methane emissions that dictate ventilation practices the optimum plan for a site is a trade off of such factors for example the more entries the higher the coal production although at the expense of advance rate on the other hand fewer entries will cause congestion but will provide faster advances modern high speed computers and simulation packages are ideal for modeling these scenarios for optimum performances a cm must be matched with a compatible haulage system in order to realize its full potential options could be a highvoltage miner matched with a continuous haulage system a batch haulage system or a surge car another benefit is that the high voltage miner generates 43 less dust using surfactants reduces that by an additional 6 underground coal mining has always been a costcompetitive industry and will continue to be so in the foreseeable future results from continuing efforts to reduce production costs and or increase productivity to stay competitive have been quite impressive according to the u s energy information administration 2010 in the period from 1949 to 1999 the underground coal productivity per worker hour rose from 0 68 t h short tons per hour to 3 99 t h a 587 increase and to 4 04 t h in 2004 a 594 increase from 1949 and a historical high this rate decreased to 3 17 t h in 2008 the continuing improvement in equipment technology system monitoring automation and system optimization in r p mining make this method more productive and indispensible in underground coal mining shrinkage stoping is a vertical overhand stoping method in which most of the broken ore remains in the stope to provide a working floor for the miners the broken ore also provides wall support until a stope is completed and ready for drawdown of the remaining ore in general the method is used in steeply dipping 50 dip narrow ore bodies with regular boundaries ore and waste both the hanging wall and the footwall should be strong and the ore should not be affected by storage in the stope the method is labor intensive and cannot readily be mechanized it is usually applied to ore bodies on narrow veins or to ore bodies for which other methods cannot be used or might be impractical or uneconomical the method can be easily applied to ore zones as narrow as 1 m 3 1 ft but it can also be successfully used for extraction of much wider stopes the method is most efficient when drilling of the ceiling or back is done with uppers instead of horizontally shrinkage stopes are mined upward in horizontal or inclined slices usually about 35 of the ore derived from the stope cuts the swell can be drawn off shrunk as mining progresses as a consequence no revenues can be obtained from the ore remaining in a stope until it is finally extracted and processed for its mineral values logically ore broken i

n a shrinkage stope should be free flowing and broken ore should not pack or re cement itself together neither the ore nor the adjacent country rock should contain undue amounts of clay or other sticky material that would cause the broken ore to hang together in the stope and make it difficult or impossible to draw the ore should not readily oxidize which may cause the broken ore pile to recement itself and consequently hang up also oxidation of the ore minerals may have an adverse effect on recoveries of values in subsequent beneficiation of the ore to avoid mining excess amounts of waste as dilution from the stope back ore developed for shrinkage stoping should be fairly continuous along the strike of the vein or ore body in many cases however it may be possible to mine around waste areas and leave them in place as random pillars in designing for shrinkage stoping consideration must also be given to the plunge or rake of the ore body especially where an entire ore body or ore shoot may be mined as a single stope rather than as preestablished stope panels with defined vertical endlines a stope with a shallow plunge or rake 50 may be difficult to mine by shrinkage methods because the ore moves away too quickly from the predeveloped extraction system occasioning the need for additional development of the system further development especially raising is also often required when the ore abruptly extends beyond stope endlines following is a summary of the parameters for the application of shrinkage stoping to a particular ore body boshkov and wright 1973 lucas and haycocks 1973 morrison and russell 1973 lyman 1982 ore characteristics the ore should be strong and nonoxidizing should not pack or stick together and should not spontaneously combust host rock characteristics the host rock should be characterized by moderately strong walls that are free of clay or geologic structures such as faults joints and so on deposit shape the deposit can be almost any shape but it should have uniform definitive boundaries deposit dip the dip should be greater than the angle of repose 39 and preferably steeper than 50 deposit size width of the ore deposit should be narrow to moderate the deposit can be of any length and height ore grade average ore grade should be moderate to high production rates production rates should usually be small to moderate vertical longitudinal sections of typical shrinkage stopes are shown in figures 13 3 1 through 13 3 3 development and preparation areas selected for shrinkage stoping are generally developed by drifting in the vein or ore body on two levels that are spaced vertically at predetermined intervals that customarily vary between 30 and 180 m 100 and 600 ft after a viable ore body has been established the next phase is to develop one or more raises to establish vertical ore continuity and provide a ventilation opening and access to the stope in some ca

ses the raises are not developed and access is only from the sill level below the stope raising may be done conventionally with platform type raise climbers or mechanically with raise boring equipment drifting is normally done using conventional drill and blast track or trackless methods stopes may be prepared with extraction raises on 7 6 to 9 m 25 to 30 ft centers the full length of the stope block each raise is fitted with a chute except for manways and or service ways that is normally of timber construction extraction raises are belled out and hogged over mined over or excavated to form the undercut for the start of the first stope cut see figure 13 3 2 another method used for preparing a stope is to blast down at least two cuts from the ore zone clean up the broken ore and install stull timbers or timber sets in the sill drift below the stope timber chutes or chinaman chutes are installed at approximately 7 6 to 9 m 25 to 30 ft intervals as part of the timbering a chinaman chute is a timber platform constructed below the caps of the timber sets ore is shoveled from the platform into cars spotted below it both of the foregoing types of stope preparation are still used but on a very limited basis a more common method for preparing shrinkage stopes in modern operations is to drive a sill drift along the vein or ore body as well as an extraction drift in the footwall of the ore zone and parallel to it the footwall drift is normally situated about 7 6 to 15 m 25 to 50 ft from the ore body subsequently draw hole extraction crosscuts are driven from the footwall drift into the sill drift on 7 6 to 15 m 25 to 50 ft centers the back of the ore body sill drift is then drilled and blasted down and the swell is extracted via the draw holes either with rail mounted mucking machines or with load haul dump lhd units see figures 13 3 1 and 13 3 4 obviously other variations for the preparation of shrinkage stopes have been used successfully these may include preparation with short crosscuts into the waste footwall from the ore sill drift driving finger raises back up to the ore and preparing the first stope cut from the finger raises another variation might be the elimination of the construction of chutes or drawpoint crosscuts and drawpoints electing instead to extract ore from the ore sill drift with remotely controlled lhd units stoping operations when an ore block has been developed for shrinkage stoping a manway is usually constructed in the raise to the next level one or more manways and service ways are usually constructed with timber on one or both ends of the stope often a timber slide is installed in one of the manways for hoisting and lowering supplies and equipment into and out of the stope hoisting and lowering are often done with an air or electric single drum tugger hoist which is installed in the level below the manway after the manways ventilation raise s and service ways have b

een established for a stope mining can commence drilling of a shrinkage stope back is usually done with handheld stoper or jackleg drills although mechanized drill wagons or jumbos have been used in wide 5 0 m 16 4 ft stopes back stoping vertical or near vertical drilling is the preferred mode of operation but breasting down horizontal drilling is also common upholes are generally 1 5 to 2 4 m 5 to 8 ft in length and usually all drill holes are loaded and a complete cut is blasted simultaneously breasts are drilled with 1 8 to 3 0 m 6 to 10 ft holes or normally are blasted once per shift drill holes are loaded with anfo ammonium nitrate and fuel oil products water gels and slurry blasting agents primers are usually water gel sticks or dynamite and holes are initiated using nonelectric electric or fuse blasting systems after a cut or round has been blasted in a stope drawdown emptying out of the 35 swell is necessary leveling of the muck pile is required as a next step to provide a work platform and facilitate drilling the next vertical or horizontal cut leveling of the ore in the stope can be done with hand shovels for small stopes with two or three drum air or electric slushers for stopes that are longer and or wider and with lhd units for large stopes after leveling construction of manways and service ways and drilling of the next stope cut are done to continue the mining cycle variations for establishing openings for manways ventilation raises or service ways may include strategically cribbed timber raises steel culverts or rings or even timber sets within the confines of the stope these installations may be desirable during the mining phase but may create safety problems or nuisances during ore drawdown due to the collapse of the structures for these facilities pinning stulling or wedging these installations to the stope walls may prevent their destruction during drawdown however if the construction materials are dragged down with the broken ore they can cause hang ups in the extraction drawpoints or chutes a stope should have strong self supporting walls to permit the application of shrinkage stoping excessive dilution through scaling of walls can preclude use of the method good mining practice and government regulations dictate a minimum ground support program for stope extraction shrinkage stoping has inherent dangers merely because workers must be in the excavated area to drill load rounds and install ground support and stope services all while standing on a broken ore pile many serious accidents have occurred when voids in the broken ore pile suddenly collapsed dragging miners and equipment down into the void in other cases a chute or draw hole over which miners were working had mistakenly been pulled with the same result the requirement that miners use safety belts in a stope may reduce the danger but it also reduces the workers flexibility ground support is usuall

y minimal in shrinkage stopes especially when the method is applied in areas of a vein or ore body that are steeply dipping have strong ore and wall rocks and have few structural features e g faults fractures or joints in the walls or within the vein or ore body however modern mining practice is to err on the side of safety and so in most cases ground control is a primary function rock bolting of stope backs and ribs with mechanical or grouted rock bolts is the preferred method of ground support in shrinkage stopes however in very narrow stopes e g 2 0 m 6 6 ft wide correct installation of bolts in the walls is quite difficult because the bolts must be placed perpendicular to the planes of the hanging wall and footwall more often than not bolts in narrow stopes are installed at acute angles to the walls which diminishes the effectiveness of the installations stulling of the ribs and back is still practiced recent south african innovations in stull installation have made it a much more attractive option stulling is simply the installation of timber props from wall to wall or floor to back typically a board headpiece is inserted between the head end of the stull and the rock hanging wall and the stull is secured by pounding wedges between the rock footwall and the butt end of the stull in the past the butt ends of stulls were almost always installed in a hitch cut from the rock with a hammer and chisel the invention of the jackpot headpiece has resulted in efficiency improvements in the installation of stulls the jackpot is a hollow steel can that fits over the butt end of a timber stull and is pumped up hydraulically to pressurize the stull against the ribs or back of the stope the unit eliminates the need to cut a hitch in rock to accommodate the stull cut the stull to a precise length and then try to tighten the installation with wooden headboards and wedges leaving pillars in areas of low grade ore or waste within the stope boundaries is a very acceptable approach to supplementary wall support the pillar should be shaped in such a way that ore does not remain on top of it on drawdown of the stope sampling of narrow shrinkage stopes is usually done by taking either channel or chip samples at systematic intervals e g 1 5 m 5 ft from the stope back ribs or face samples are usually taken by hand sometimes mechanical means can be applied in the process when high sulfides are present x ray analyzers can be used for sampling in wider stopes drill sampling of the back and ribs can be done the drill sample holes may crisscross the stope back according to a predetermined pattern and may also be drilled into the stope walls drill cutting samples are collected in a canvas sample bag or other container through a hose and funnel or other device stope drawdown shrinkage stope drawdown usually results in a consistent and steady supply of ore to the process plant and or mine stockpiles if a mine has s

everal shrinkage stopes in various phases of the mining cycle there is no reason the ore flow to the plant or stockpiles should ever diminish after completion of the initial stopes in the mine plan maintaining a steady ore flow from shrinkage stoping is largely dependent on engineering the optimum production rate for a given operation completed stopes can be drawn down from strategically placed chutes and from drawpoint crosscuts or they can be remotely mucked with lhd units in some mines slusher trenches developed below the stopes have been used effectively to remove broken ore from shrinkage stopes haulage from the stope extraction points can be done using trackbound or trackless equipment chutes should be robustly designed and well constructed to ensure that they will not be destroyed during the extraction phase shrinkage stopes usually should be drawn down evenly so that if the stope walls do peel or slough the waste material remains atop the pile and does not add dilution to the broken ore after a stope drawdown has started ground control of the walls pillar recovery ore hang ups and so on is minimal reentry of miners into a stope during the drawdown phase is not recommended one of the most dangerous jobs in a mine is the drawdown of shrinkage stopes especially where the ore contains blocky or sticky material that may cause the broken ore to hang up in the stope stope hang ups are often removed by washing them down with water by destroying them with explosives or by blowing them down using high pressure air blowpipes in the past miners entered completed stopes to pick them down by hand but as has been stated previously this is unsafe and not recommended a summary of the advantages and disadvantages of the shrinkage stoping method is shown in table 13 3 1 variations and applications variations of shrinkage stoping include inclined or rill shrinkage and long hole shrinkage large pillars or small isolated ore blocks may sometimes be recovered using shrinkage methods certain underground openings such as short shafts may also be excavated using shrinkage methods a mine that employed shrinkage stoping as a primary mining method was the now closed homestake gold mine in lead south dakota united states the homestake ore bodies were mainly fracture fillings situated in the limbs of folds bullpen stopes i e fixed minimum width shrinkage stopes laid out in a transverse direction to the long axis of a large ore body 15 m 50 ft wide were developed across the great main ledge ore body stopes were mined over timbered sills where strategically located chinaman chutes were constructed for ore extraction stopes were mined overhand for about 21 5 m 70 ft to within 9 1 m 30 ft of the next level ore pillars 7 6 m 25 ft thick were left between the primary stopes and these along with the crown pillars left over in the primary stopes were subsequently extracted with square set stopes homestake abandoned

this type of shrinkage stoping just before world war ii a second example of large scale shrinkage stoping is the shut down idarado mine located between ouray and telluride in colorado united states the idarado ore bodies were mainly steeply dipping veins with excellent ore continuity shrinkage stopes were mined along the veins full width spans that varied between 1 5 and 7 6 m 5 and 25 ft the consistent ore continuity permitted the mine planners to divide the veins into stope panels generally 122 m 400 ft long and about 61 m 200 ft high panels were prepared over a slusher trench situated about 6 1 m 20 ft above the primary on vein drift raises on 7 6 m 25 ft centers developed from the slusher trench were belled out and interconnected above the slusher trench to form the opening for the first stope cut ore extracted from the stope was slushed from the extraction raises to a chute in the exact center of the stope the idarado operations were closed in 1979 variations on the idarado system were practiced at the morococha and casapalca mines of the cerro de pasco corporation in the andes mountains of central peru in south america and at the tayoltita mine in the state of durango mexico stopes in these mines were prepared conventionally over the main on vein development level by driving 7 6 m 25 ft long raises on 7 6 m 25 ft centers the raises were belled out to form the opening for the first stope cut about 5 m 16 ft above the main level on completion of this step each raise was fitted with a timber chute for ore extraction see figures 13 3 1 13 3 2 and 13 3 4 in all of these examples a raise was first developed through each ore block or stope panel for ventilation manway and service way manways from the sill level were carried as timber cribbed raises within the stope as stulled and lagged raises on the ends of a stope or as in the case of the idarado mine as boreholes located 3 m 10 ft inside the footwall of the stope in the vein mines drilling was done with pneumatic handheld stopers or jackleg drills whereas in the homestake mine drilling was done with bar and column leyner type drills the terms inclined shrinkage and rill shrinkage refer to an application in which multiple faces or benches for drilling are carried along the back of the stope as it is mined upward figure 13 3 3 this system was common practice at the pachuca mine in the state of hidalgo mexico stopes were developed conventionally over pillars as described previously and chutes on 7 6 m 25 ft centers were installed on the extraction raises pachuca miners preferred to use jackleg drills rather than stopers and with the rill system multiple faces were available for breast stoping in a given shift see figure 13 3 3 long hole shrinkage figure 13 3 5 is developed in the manner previously described the method was applied with limited success at the stillwater mine near nye montana united states in this method dr

illing of the stope block is done from vertical raises driven through the ore block or panel on 15 to 30 m 50 to 100 ft centers raises in this application are preferably developed with raise climbers or with cage raising techniques the raise climber platform or the cage serves both as the entry exit vehicle for the raise and the platform from which to drill and blast the stope block long holes are drilled from the raise parallel with the strike of the ore zone and these are also loaded and primed from the climber or cage platforms initiation of the stope blast is done from a safe area on the service level below the stope shafts winzes or other large underground openings may be developed through shrinkage methods often these excavations are done as described previously for the long hole shrinkage application in certain situations conventional shrinkage of a shaft winze or other excavation may be necessary safety efficiencies and costs shrinkage stoping has largely been discontinued in most countries because it has numerous safety hazards is very labor intensive and largely inefficient and is generally expensive coupled with these reasons is that more than 60 of the mined ore remains in the stope until stoping is completed which may take many months and sometimes even years during this period amortization of the investment required to develop and mine an ore block by shrinkage stoping may outweigh any advantage in applying the method to a particular operation see figure 13 3 5 typically the efficiencies in a modern shrinkage stoping operation are in the range of about 0 3 to 1 2 t 0 3 to 1 3 st per worker hour in western operations very few miners have experience in the development and operation of shrinkage stopes but the method is still widely practiced in latin america and in developing countries elsewhere in the world costs in a typical north american shrinkage operation are currently on the order of us 25 to us 50 t us 55 st with more than 50 of the costs assigned to labor operations in latin america or asia may enjoy lower costs because labor is less expensive but efficiencies are generally not equal to those of north american operations following are definitions of general terms for this mining method also shown schematically in figure 13 4 1 span the length of the stope along the strike width the perpendicular distance between the footwall and the hanging wall height the distance along the exposed hanging wall and not the vertical height between levels longitudinal pillar a pillar aligned along the strike of the stope rib pillar a pillar aligned transverse of the stope perpendicular to the strike sill pillar horizontal pillars that separate levels or stopes dilution the reduction of ore grade due to mixing of ore with barren rock internal dilution rock that must be mined because of the geometry of the ore body and the requirement to mine rectangular areas the te

rm is synonymous with planned dilution external dilution dilution caused by sloughing or failure of stope walls and back and is outside the blasted stope boundary external dilution is defined as the external waste tonnage divided by the ore tonnage the term is synonymous with unplanned dilution sublevel stoping in the absence of consolidated fill employs pillars to separate the individual stopes to reduce the potential for wall slough sublevel stoping requires a straight linear layout of stope and ore boundaries inside of the stope everything is ore with no chance of recovering small mineralization in the wall rock this method requires knowledge of the ore boundaries as shown in figure 13 4 2 sublevel stoping with no fill is a mining method in which ore is mined and the stope is left empty the result is a large void that requires individual pillars be placed to separate the stopes sublevel stoping is largely restricted to steeply dipping ore bodies 50 90 with a competent hanging wall hw and footwall fw figure 13 4 3 shows the general approach to sublevel stoping whereby ring drilling is used from levels generally spaced 20 m apart in a vertical dimension level spacing is largely limited by the length of the production holes which range in diameter from 50 75 mm and maximum lengths of 25 m this can be modified if in the hole ith drills or top hammer tube drills are used characteristic with the sublevel stoping method are the intermediate levels which largely differ from long hole blasthole stoping as depicted in figure 13 4 4 where the intermediate level has been removed in sublevel stoping the mining is accomplished from individual levels at predetermined vertical intervals these intervals are largely governed by ore geometry in order to minimize internal dilution by enabling the extraction of irregular ore bodies rock mechanics constraints in terms of minimizing the external dilution through wall slough and or operational restrictions such as drilling equipment constraints sublevel and long hole methods require blasting into a vertical slot free face whereas a vcr shown in figure 13 4 5 differs in terms of blasting to a horizontal free face which is largely confined due to the muck remaining within the stope as only the swell is drawn variations of the sublevel method include narrow vein mining alimak avoca longitudinal sublevel retreat and transverse stoping as well as historical methods such as slusher and track mining haycocks and aelick 1998 sublevel stoping requirements and constraints the following variables must be addressed in sublevel stope designs size the minimum width generally ranges from 3 m to 6 m however in isolated cases it reaches 1 5 m clark and pakalnis 1997 and lower 0 8 m the width is governed by the production blast pattern which with the use of 50 mm blastholes is typically 1 2 1 2 m and the stope layout is based on this spacing figure 13 4 6

shape the shape is preferably tabular and regular in shape from level to level dip the dip is preferably greater than the angle of draw which typically is in excess of 50 in practice the concern also is that a shallow hanging wall dip will result in a less stable hw configuration because of gravity influences and increased wall exposures between vertical stope horizons all resulting in increased potential for external dilution geotechnical this requires a moderate to strong ore strength and generally a competent hw fw as these will be exposed and affect the level of external dilution the ore will determine the potential pillar sizes hole squeeze and block size that affect production stope productivity pakalnis 2002 stope spans since this is a nonentry method stope spans can be larger the span should be designed to control external dilution and avoid stope collapse and air blast span length is governed by hw rock mass quality and generally is in the range of 30 m with the stope height inclined in excess of 30 to 60 m pillar size the purpose of the pillars is to support the crosscuts and divide up the stopes the size of the pillars is dependent on induced stresses structure rock mass and operational constraints selectivity selectivity is limited because waste zones can be incorporated as pillars changes in ore body geometry outlines are difficult to address unless the ore body narrows to the next pillar or sublevel where the drill pattern can be modified figure 13 4 7 design considerations for sublevel stoping general design guidelines the design of a sublevel stope starts with an engineered layout that incorporates the geometry of the stope stope span stope height pillar dimensions drill levels and draw levels figure 13 4 8 this layout is then superimposed upon the ore contours plan as defined from the upper drill drive to the lower draw level horizon the example shown in figure 13 4 9 is a schematic of a 60 m long hole stope 151 mm blastholes with geologic contour intervals shown every 10 m in the plan the resultant longitudinal composite is shown in figure 13 4 10 employing a ring burden distance between drill rings of 3 m from pillar stope boundary to pillar stope boundary development considerations sublevel stoping uses long hole drilling employing extension drill steels to achieve the appropriate blasthole depth when ring drilling is used the entire cross section of the stope is drilled with holes that radiate from the drill drive the drilling pattern is matched to the shape of the ore body and location of the drill drift parallel holes are drilled when the drill drive can be silled out from the hw fw but this largely is constrained by the stability of the exposed working back two principal drill systems exist top hammer and in the hole hammer both require long hole rock drills equipped with extension steel in 1 2 1 8 m long sections top hammer drills are more suite

d for narrower ore bodies sublevel stoping while ith hammers are more suited for wider ore bodies long hole stoping these will be discussed in a later section the blast layout for the individual rings will incorporate the ring number hole number on that particular ring the amount of explosive required kilograms delay interval angle of hole to be drilled length of hole to be drilled and the depth of collar stemming to be used a slot raise must be developed in order to accommodate the swell of the blasted muck it generally is developed at the extremity of the stope as shown in figure 13 4 13 and subsequently the slot raise is enlarged fw to hw to open up the area for blasting generally one cannot blast rings into a narrower void so the slot should be located in the largest area of the stope ore handling considerations ore handling in sublevel stoping involves removal of the ore at the bottom of the stope and historically it involved track and or slushers to remove the muck this process is now conducted largely by trackless mining equipment such as scoop trams used for drawpoint loading into mine trucks and or orepasses as shown in figures 13 4 3 and 13 4 4 sublevel stoping sublevel stoping design is schematically shown in figure 13 4 14 and the sequence of development and extraction sequencing is shown in figure 13 4 15 the dimensions noted in the figures are typical of sublevel stoping dimensions and are employed solely to assist in the description of the method and not intended for design as the dimensions of a stope are based on the geometry of the ore body and operational constraints this mining method employs sublevels located approximately 20 30 m apart the distance between sublevels is largely governed by the length of hole that can be drilled with minimal drill hole deviation under 2 the drill hole diameter ranges from 50 75 mm using top hammer drills which restricts the length of the hole to generally under 30 m with blasthole burden and toe spacing between approximately 1 1 m and 2 2 m typically 1 2 1 2 m modern tube drills top hammer at 100 mm in diameter are able to drill 35 40 m long holes generally if the stope width hw fw is greater than 15 m an fw and hw development drive as shown in figure 13 4 14 is used otherwise only a center drive in the middle of the stope is developed the initial development is shown in figure 13 4 15a whereby the drill drives slot x cut and raises are driven the drill drives are comprised of the draw level intermediate level and the upper drill level the undercut figure 13 4 15b is silled out for a vertical height of approximately 12 m above the draw level the height of undercut or void can be minimized through the use of programmable detonators ensuring that sufficient void is created for the subsequent blast the undercut serves the purpose as well to ensure breakthrough of the holes from the upper drill drive a 2 2 m slot is bored bla

sted to the full length of the level above the upper drill level figure 13 4 15b which is subsequently slashed to 3 7 3 7 m for the full stope height and width from fw to hw figure 13 4 15c to provide sufficient void space for the subsequent rings to be mined production blasting figure 13 4 15c is comprised of individual rings blasting into the void for the full stope width on either side of the slot this assumes pillar access exists on either side of the slot normally the production rings blasted from the intermediary level correspond to a similar set of rings on the upper drill level figure 13 4 15d to ensure that a void from draw level to upper drill level exists the geometry shown in figure 13 4 14 employs a ring burden of 1 5 m and toe spacing of 2 1 m the stope is normally drilled off prior to commencement of blasting and only the holes that are scheduled for the blast are loaded the upholes from the intervening levels must ensure interleaf coverage of approximately 1 to 2 m the example shown in figure 13 4 15 uses 15 m long drill holes with uppers and downholes and 1 2 m interleaf coverage long hole stoping long hole blasthole stoping development is shown in figure 13 4 16 with the sequence of development and extraction shown in figure 13 4 17 the subsequent examples given are typical of long hole stoping dimensions and are employed solely to assist in the description of the method and not intended for design as these dimensions change based on the geometry of the ore body and operational constraints long hole stoping largely eliminates the intermediary level with the draw and drill horizon interval governed by the length of hole that can be drilled with minimal drill hole deviation under 2 the drill hole diameter ranges from 75 to 150 mm using ith hammer bits thereby enabling the lengths to approach 30 60 m in length with blasthole burden and toe spacing approximately 3 4 m2 3 3 m the development is as shown in figure 13 4 17 generally if the stope width hw fw is greater than 15 m an fw and hw development drive is used as shown in figure 13 4 16 otherwise a center drive is driven in the middle of the stope the initial development is shown in figure 13 4 17a whereby the drill drives slot crosscut and raises are driven the drill drives are comprised of the draw level and the upper drill level as the intermediary level has been removed the undercut figure 13 4 17b is silled out for a vertical height of 12 m above the draw level a 3 7 3 7 m slot is bored blasted to 12 m above the upper drill level figure 13 4 17b which is subsequently slashed to 6 1 6 1 m for the full stope height and width from fw to hw figure 13 4 17c production blasting figure 13 4 17d is comprised of individual rings blasting into the void for the full stope width on either side of the slot this assumes pillar access exists on either side of the slot the geometry shown in figure 13 4 16 employs a ring bu

rden of 3 m and toe spacing of 4 2 m with 150 mm diameter blastholes the stope is normally drilled off prior to commencement of blasting the example shown in figure 13 4 17 uses 46 m long downholes and 15 m upholes a variation in the above sublevel and long hole mining methods is to use nonconsolidated backfill above the upper drill drive of lift 1 and subsequently drawing out from the level that serves as the draw horizon for the level above lift 2 as shown in figure 13 4 18 this negates the need for cones in ore and consequently maximizes ore recovery as shown in figures 13 4 15 and 13 4 17 the cones can be eliminated with the use of remote mucking equipment however the equipment will be traversing under potentially extended unsupported spans see figure 13 4 5 vertical crater retreat as shown in figure 13 4 18 a vcr is a variation of long hole open stoping where the free face is not a vertical slot but a flat back at the base of the block to be mined spherical charges are used to break the ore into slabs as shown in figure 13 4 5 and have a length diameter l d ratio of 6 1 field testing has shown that a ratio of explosive column length l to hole diameter d of 6 or less will behave similarly to a spherical charge blasting is carried out in horizontal slabs with only the swell being mucked at the drawpoint this is a form of shrinkage stoping where the broken stope muck provides passive support to the stope walls the ore is recovered at the base of the stope through drawpoints similar requirements and constraints to that of sublevel stoping exist except for the need for a competent hw fw due to the option of maintaining the stope full of muck development is similar to that of long hole stoping requiring an upper drill horizon and draw level and it is generally recommended to sill out at the drill horizon to provide drill coverage for the entire block the vertical separation between drill and draw level is largely a function of the ore regularity and drill accuracy as detailed in general for the long hole mining method the dimensions are similar to that of long hole mining where ith drills are employed with heights ranging from 30 60 m and 75 150 mm drill diameters are used a typical loaded blasthole for vcr is shown in figure 13 4 19 employing a single deck charge advantages of the vcr are the high productivity associated with this bulk mining method and the ability to mechanize the ability to only muck the swell enables support to the stope walls an advantage of this method over shrinkage is the nonentry and high mechanization associated with vcr disadvantages of this method are the extensive pre stope planning and development that is required prior to commencement of production mining as the stope must be largely drilled off prior to bench blasting similar disadvantages to that of shrinkage mining exist in having the broken ore within the stope until the end of mining of the block variations on sublevel

stoping the sublevel mining method has variations that have been implemented and will be discussed in the context of its similarity with sublevel stoping vein mining vein mining also termed alimak mining has been employed within narrow vein ore bodies as detailed in the namew lake mine canada case study madsen et al 1991 access to the ore is gained by a bored raise alimak such as that shown in figure 13 4 20 the diameter of the raise is approximately 2 3 m and extends from draw level alimak drive to upper drill drive as shown in figure 13 4 20 item 1 in the figure and spans the length of the ultimate stope span with similar constraints as those detailed for sublevel stoping support may be in the form of cable bolts item 2 in the hw providing the final wall support on stope extraction the ore is drilled laterally by conventional drills long hole jumbos or other methods and ranges from 5 m to 15 m in length to the adjacent stope item 3 with blasting from the draw level vertically to the upper drill level item 4 an intervening pillar may be left between stopes or the stope mined from one alimak raise to the next depending on the geotechnical constraints the major advantage is the ability to mine narrow ore bodies with minimal horizontal development the vertical height of the alimak is largely limited by operational and geotechnical constraints and reaches heights of 30 to 100 m blasthole sizes are generally 50 75 mm with burdens and spacing similar to that of sublevel stoping 1 2 m2 transverse open stoping variations of sublevel stoping with delayed fill are shown in figure 13 4 5 this mining method is largely used for stope widths in excess of 20 to 30 m or as dictated by geotechnical stable back spans otherwise conventional longitudinal or strike mining is used figure 13 4 5 shows the objective is to recover the secondary pillars between the primary stoping blocks which can be excavated by sublevel stoping general and subsequently filled with consolidated fill that can be comprised of hydraulic fill paste or cemented rock fill mining of the secondaries occurs after curing of the primaries to a strength that is able to withstand minimal dilution generally the binder content ratio is 30 1 to 20 1 fill to cement by volume alternatively a permanent pillar is left behind to confine the unconsolidated fill with only primaries excavated along the strike with this variation the secondaries are narrow pillars left behind approximately 3 5 m a disadvantage of this method is its inability to follow the variations of an irregular hanging wall dip longitudinal mining figure 13 4 21 shows sublevel extraction employing mucking along the strike retreat this is a variation of conventional fw drawpoints as shown in figure 13 4 21 the stopes with no fill are as shown in figure 13 4 21 and with delayed fill are as shown in figure 13 4 22 the delayed fill method of longitudinal mining is also referred to as avoca

mining having longitudinal mucking access requires that remote load hauldump lhd equipment be used this method is also referred to as sublevel benching cut and fill is the broad descriptive term applied to mining methods requiring that excavated voids be filled with barren material to facilitate the continuation of ore production this fill is required to provide support for subsequent openings or to provide a working platform for further mining cut andfill methods may be applied in conjunction with other conventional methods such as blasthole stoping or they may be methods that stand alone such as drift and fill classic cut and fill mining involves the successive mining of horizontal or inclined slices upward through a relatively narrow and subvertical tabular ore body this is followed by the placement of uncemented waste rock or hydraulically placed sand fill to create a new higher working level to gain access to ore above and to support the ground below raises maintained up through the fill and occasionally down from upper levels provide for access ventilation ore removal and the drainage of water from the fill material modern variants may use paste fill and typically use ramp access to allow the use of mechanized mining equipment these techniques preclude the need for raises none of the cut and fill methods described herein can match the productivity or the low mining cost of mass mining methods cut and fill methods are chosen because they are more economically attractive than other extraction methods available for the same situation favorable conditions cut and fill methods are applicable to a wide variety of situations including the following ore zones are irregular in shape and orientation ore grade is high and dilution control is critical the precise contacts between ore and waste are structurally critical but not readily visible the waste rock is weak ore zones are large but their rock quality is weak localized underground stability is required surface disturbance must be minimized the value of the ore makes recovery of support pillars economically viable the reduction of surface waste storage is important the need exists to advance a working platform for the upward mining of the ore body cut and fill methods are favored when the ore value is relatively high and ore recovery rate with minimal dilution cannot be satisfactorily accomplished by open stope mining or caving methods if the openings are small enough and the ground conditions are competent irregular ore boundaries can be mined with open stope methods without backfill however when the openings are larger and the ground conditions are less favorable backfilling becomes necessary for safe economic production fill methods are useful when regional stability is needed this is the case where large empty stopes could adversely affect the stability of permanent mine access drifts shafts and raises these stresses could als

o affect new mine production stopes and surface infrastructure there is also the advantage that fill can provide a relatively inexpensive and convenient way to advance a working platform to maintain access to the ore for drilling blasting and mucking another advantage is that ventilation air can be managed more efficiently because vacant and unused mine openings are closed by backfill limitations these factors limit the applicability of cut and fill the availability of a sufficient quantity of a suitable fill material type the cost of binding agents if required production preparation transportation and placement cost of the fill material storage and reclamation facilities to match the mining cycle demand congestion and interruption of production mining activities the principal drawback for the cut and fill methods is the additional cost of producing preparation transporting and placing the backfill not all readily available fill materials are appropriate the backfill material must match its intended use the material that is most readily available and least expensive may not have characteristics that would allow its use as fill material the material most readily available also may not be usable in the stope designs dictated by the size and shape of the ore body further the least expensive delivery system may not be practical assumptions about fill availability should be verified early in mine planning suitable aggregate for cemented rock fill may not be available near the mine site both sand fill and paste fill generally make use of tailings from an on site mill for sand fill however sufficient coarse fraction may not be available within the tailings either because of the fineness of the grind or because of the need to supply coarse fraction for tailings dam construction conversely the grind may be too coarse to provide sufficient fines for paste fill in addition to the added cost backfilling operations can result in congestion maintenance and cleanup problems underground cut and fill imposes limitations on the sequence and timing of production that if not carefully controlled may take precedence over the maximization of production grade usually conventional ground support in the form of rock bolts wire mesh and shotcrete still will be required to temporarily support the stope openings during production mining the cost of producing and placing fill requires consideration of the mining method trade offs unless the deposit simply cannot be mined by any nonfill method the additional cost and the advantages of using fill must result in additional ore production increased recovery reduced dilution and or cash cost reductions sufficient to justify the additional cost of the fill and its preparation and placement types of cut and fill methods there are numerous cut and fill methods and many hybrid combinations the following is a basic set of descriptors vertical progression of the mining und

erhand mining beneath the backfill overhand mining on top of the backfill mining method drift and fill postpillar bench and fill blasthole stoping uphole flat back slicing or back stoping type of fill cemented rock fill uncemented rock fill sand fill hydraulic paste fill timing concurrent placed during the cycle of stope advance delayed placed after the excavation of the stope is complete cut and fill mining is extremely flexible and various combinations of mining methods and fill materials can be used to cope with specific mining situations however some combinations are not feasible such as underhand drift and fill with uncemented rock fill further discussion regarding possible combinations can be found in the cemented fill stopes and uncemented fill stopes sections method selection criteria ore body morphology the size shape orientation and degree of boundary uniformity of the ore body are the first considerations in selecting a cut and fill method the use of fill for support and for provision of the necessary platform to access the ore provides the operator with great flexibility for adjusting the excavation to irregularities along the boundaries of the ore body alternative methods that do not use fill may result in unsafe working conditions excessive dilution or insufficient recovery cutand fill methods can be adapted to flat vertical undulating or even en echelon zones and to situations in which the ore body shape and attitude change quickly along the strike or dip access infrastructure and development cost typically overall project cost for access and infrastructure to support cut and fill mining is less than the cost for bulk mining methods like block caving overall project cost is greater however than the cost for room and pillar or open stoping the use of fill for support rather than leaving unmined pillars of ore results in greater recovery of ore because fewer unrecoverable ore pillars are left behind at the conclusion of mine operations in methods such as shrinkage stoping much of the broken ore is left albeit temporarily in the stope to provide wall rock support and a working platform from which to mine the whole block of ore although the shrinkage ore is withdrawn at the completion of the stope the use of fill as a working platform and wall rock support results in a faster return on the cost of mining amenability to mechanization the mechanization of cut and fill mining has two phases mechanization of the ore extraction i e the cut and mechanization of the fill process the mechanization potential for ore extraction with cut and fill methods varies widely the ground conditions that dictate the use of cut and fill mining may also limit drift dimensions or alternatively may require flexibility to adapt to varying stope conditions this reduces the size of trucks and loaders that can be used options associated with production and delivery of fill

are covered in chapter 13 6 by itself the use of fill seldom enables an increase in the level of mechanization dilution cut and fill methods increase the extractable yield of ore bodies and reduce dilution from boundary overbreak however the effect of dilution from fill must be considered in any economic analysis fill dilution generally carries no recoverable value whereas overbreak into country rock may carry at least some value because of its proximity to the ore dilution from fill can come from stope ribs where the adjacent stopes have been previously backfilled if the stope in production is above a previously filled stope the mucking operations will typically remove a layer of fill material from the floor to clean up as much ore as possible less fill will be taken if the floor surface is cemented fill conversely ore can be lost in the floor of a stope if ore is used to build a roadway on top of the fill fine ore which usually contains a significant portion of value can be lost into the fill from the cut below if the fill is not well consolidated at its placement the dilution potential of various fill materials can differ significantly uncemented rock and sand fill typically dilute at the highest rates dilution is usually less with cemented rock sand and paste fills however dilution from cementing agents may have an adverse effect in the metallurgical plant this difference in behavior must be included when evaluating the relative costs of using these fill materials in the mining operation geotechnical constraints the in situ geotechnical environment including the rock quality of the ore and country rock geologic structures ore body depth water and regional stresses determines the size and shape of the stopes that can be excavated before placing fill the ground support required to provide a safe working environment is also dictated by the geotechnical situation cut and fill methods frequently rely on the initial mining of stopes in virgin ore zones these stopes are called primary stopes they are subsequently backfilled to provide support for the mining of adjacent stopes which are called secondary stopes the size shape and mining sequence of the primary and secondary stopes are dictated by the geotechnical conditions of the ore and country rock and the mechanical properties of the planned fill material sometimes it is advantageous to place higher strength highercost fill in a primary stope and lower strength lower cost fill in a secondary stope the flexibility to control and differentiate fill quality and properties according to the application is crucial to mine operation safety and cost control sill pillars and crown pillars are used to separate stoping zones vertically the dimensions of these pillars are determined by the geotechnical environment and the stresses induced by mining if these pillars contain valuable ore it may be desirable to place highly engineered fill to facilitate the safe r

ecovery of this ore the objectives are to use cutand fill mine planning to maximize the extraction of the ore body and to avoid sterilizing any portion such that it cannot be safely recovered availability of fill material the availability of fill material rock sand or paste affects the selection of cut and fill methods the quantity engineering properties and cost must be matched to the planned use for example if the on site mill tailings are unsuitable a nearby source of alternative material must be found similarly if the price for cement is unreasonably high the cost and availability of the various fill alternatives must be weighed against the engineering requirements of the selected cut and fill method the safest and least expensive combination of fill material and cut and fill method must be found and continually reviewed the use of tailings for fill is often dependent on mill throughput an unfortunate side effect of using mill tails particularly in mines with numerous smaller stopes is that when stopes stop producing to allow for fill placement they must rely on continued production from other stopes to provide fill material the design of the mine must include satisfactory storage capacity for fill because it is usually accumulated at a slow relatively steady rate typically as a by product of the milling operation when fill is required for placement it must be delivered at a much greater rate stope by stope careful planning is therefore critical to ensure that a balanced mining filling cycle is maintained throughout the operation failure to maintain a balanced cycle can result in fill shortfalls leading eventually to loss of production until a proper cycle can be reestablished cemented fill stopes the inclusion of a binding agent such as cement in the fill is required when the fill must have sufficient strength to provide support for the excavation of stopes located above below or adjacent to the stope being mined the term delayed is used to indicate that extraction of all ore in the stope has been completed before the stope is backfilled occasionally when the value of the ore in the pillars of a conventional room and pillar mine is considered high enough cemented fill is placed around the pillars to permit recovery of the ore placement of cemented fill is also sometimes used to facilitate the later extraction of ore from sill pillars where the fill material is normally uncemented the use of hydraulic and cemented fill can cause significant problems with level cleanup sumps settlers and pumps this is usually worse with cut and fill stoping than with sublevel open stoping because there are many more points of drainage to manage blasthole stoping with delayed backfill blasthole stoping with delayed backfill is a method that can be applied to ore bodies that are generally tabular and vertical or near vertical the rock quality of both the ore and the waste needs to be fair to good cemented fill for the

primary stopes is required when a primary secondary plan is pursued cemented fill allows secondary stopes to be mined subsequently the secondary stopes may be backfilled partially or completely with uncemented fill longitudinal blasthole stoping with delayed backfill where the ore zone is narrow and the long axes of the stopes are parallel to the strike the stope orientation is called longitudinal occasionally the ore is contained in en echelon lenses narrow parallel zones of ore separated by narrow parallel zones of waste these conditions require sequences of longitudinal stopes followed by backfilling to prevent stability problems and dilution from the collapse of the waste material between ore lenses see figure 13 5 1 the width of the stopes is generally the distance between hanging wall and footwall the maximum length is generally limited by geotechnical considerations to minimize dilution from overbreak or slough when the hanging wall is weak the installation of mechanical support bolts wire shotcrete or cable bolts may be required transverse blasthole stoping with delayed backfill where the ore zone has sufficient width the long axis orientation of the stopes may be perpendicular to the strike these are transverse stopes geotechnical considerations dictate the transverse orientation as well as the stope dimensions the sequence of lateral and vertical primary and secondary stope mining backfilling and curing requires careful planning see figure 13 5 2 after mining begins the plan seldom permits significant changes it becomes difficult to adjust the production rate and grade drift and fill drift and fill is the term applied to projects in which the ore is extracted by excavating parallel drifts through the ore zone and backfilling them before mining the adjacent drift the drifts are usually alternately filled with cemented and uncemented rock fill the orientation of the drifts mined on a level can be longitudinal transverse or diagonal to the strike of the ore body for additional strength the orientation is sometimes alternated between levels excavation of the ore by drifting is less productive and more costly than excavation by blasthole stoping therefore drift and fill is usually employed where ground conditions are too weak and the ore boundaries are too irregular for blasthole stoping because of its higher associated operating costs drift and fill is best used where the ore grade is high overhand drift and fill overhand drift and fill is a bottom up mining sequence see figure 13 5 3 this method is useful where the ore occurs in relatively thin shallow dipping lenses with weak back or wall rock the bottom up sequence may also be desirable because of production grade considerations ideally this method would be used where the rock quality of the ore in the back is not sufficiently high to permit postpillar stoping this method permits backfilling of secondary stope drifts with uncemented f

ill thus reducing costs the first or lowest cut requires cemented fill in both primary and secondary drifts if mining blocks from below are scheduled to mine up to a previously mined block underhand drift and fill underhand drift and fill is a top down sequence see figure 13 5 4 the use of underhand drift and fill stoping fills the critical need for a method that facilitates the mining of high grade deposits with weak ore and weak country rock this method results in the creation of a strong crown pillar of cemented fill beneath which subsequent mining can safely proceed the initial level may be opened using small drifts that are easily supported after the upper level is completed and backfilled succeeding levels below can be opened to more efficient mining widths orienting the drifts on each level such that they are not parallel to the overhead drifts reduces the risk of ground falls the fill material must be carefully engineered and the samples tested regularly the method may involve the use of mats or tension cables to support and distribute the load above each open drift the ore zone should be vertical or near vertical to maximize the benefit of employing this method uncemented fill stopes the use of uncemented fill has the advantage of eliminating the cost for cement uncemented fill has limitations when used as the primary backfilling material because it lacks the strength of cemented fill therefore it is seldom used where the mining of adjacent stopes or stopes below is planned the use of uncemented fill dictates a bottom up mining sequence or the use of substantial sill pillars between levels usually stope mining with uncemented rock fill has the advantage of being able to use any sort of readily available and inexpensive fill material frequently waste muck from development mining can be used this relieves the expense of transporting thve waste to the surface for disposal alternatively the rock can be quarried on the surface and delivered to the underground stopes via borehole or waste pass because of its coarse texture rock fill is often capped with sand fill or cemented sand fill to provide a mucking surface that is less susceptible to ore loss and dilution postpillar mining in horizontal or near horizontal ore bodies of substantial thickness and width the placement of uncemented fill serves the primary purpose of providing access to excavate successive slices of ore in a bottom up sequence the postpillar method can be described as a room and pillar excavation where backfill usually uncemented rock fill is placed on the floor so that successive layers can be mined upward through the ore zone see figure 13 5 5 the pillars or at least the upper portions of the pillars can be preserved and extended upward to continue supporting the back the pillars can be much taller and more slender when they are surrounded by fill and so recovery is higher than for nonfill room and pillar methods vertical or near v

ertical cut and fill stoping conventional or mechanized the use of uncemented fill in narrow vein stopes that are inclined or vertical provides both stability and a working platform for the upward advance of the excavation of the ore hydraulic sand fill or rock fill with sand is superior to rock fill alone for this application because greater density is achieved frequently a top layer of cemented fill is used to provide a stronger surface for machine operation and to minimize ore loss and dilution using sand usually coarse tailings to backfill stopes was one of the earliest fill applications in mining it was often used in conjunction with timber support but the use of timber is now largely supplanted by bolts and shotcrete the transport distribution and placement of fill are as hydraulic slurry via a pipe system less frequently compressed air is used for the transport and pneumatic stowing of the fill the pipe system prevents congestion of the haulageways and eliminates the need for trucks and loaders with hydraulic fill drainage of the excess water must be accommodated with designed filters drainage raises sumps and pumping systems these features impose substantial operational costs and scheduling limitations pneumatic fill systems have increased ventilation requirements and pipe wear as well as the physical limits associated with distribution and the potential health risks of increased airborne dust in conventional cut and fill stopes access to the level of the stope cut was gained by timbered or steel tubbed raises that were maintained and extended up through the backfill from the haulage level see figure 13 5 6 in the stope drilling was accomplished with hand operated stoper drills for vertical holes or jacklegs for horizontal holes where the cut would be taken by breasting down slushers were the principal means for mucking the blasted ore across the stope to the muck raise from the bottom of the muck raise chutes were used to transfer ore to muck cars for transport to the shaft by rail in addition to ore transport the raises provided manway access hoisting of supplies pipe and power reticulation and drainage of the backfill water in the conventional system it was difficult to bring equipment into or out of the stope after a piece of equipment had been installed in a stope it remained isolated there rising with the level of the cut low unit cost light weight simplicity and durability were important equipment characteristics slushers fit this description well however they were not as productive for mucking as low profile lhds load haul dump units with buckets overshot muckers powered by compressed air were used at some mines to replace slushers overshot muckers are relatively simple to maintain and ventilation requirements are no greater than those required for slushers eventually diesel powered lhds were placed into service figure 13 5 7 with their introduction ventilation requireme

nts increased and maintenance requirements demanded better access than raises could provide the cost of lhds forced the need to achieve higher utilization rates which in turn made it necessary to allow stope to stope movement of lhds this change led to the shift from access by raises to access by ramp systems resuing resuing is the term applied to the mining method whereby high grade ore from very narrow veins is extracted selectively before stopes are widened in waste rock to obtain the additional width required for equipment to operate see figure 13 5 8 resuing requires a working platform of fill bench and fill bench and fill stoping is applied to vertical and near vertical stopes that have sufficient width length and competency of wall rock to permit a system of laterally retreating blastholes followed by advancing waste rock fill see figure 13 5 9 this method allows variations in placement of the fill the amount of waste rock exposed between the blasting and mucking of ore and the advancing face of waste fill can be adjusted to suit local conditions in some cases the waste fill is advanced such that all open space is completely eliminated this lack of open space enhances support for weak hanging wall and footwall conditions thus reducing dilution and ore loss the use of some cemented fill to augment the uncemented fill can also increase recovery stope mechanization the widespread introduction of mobile equipment dieselpowered loaders trucks drills and service vehicles since the 1960s and 1970s has largely replaced the use of jacklegs stopers and slushers the primary underground mining machinery of the prior era the maintenance and maximum usage of this mobile equipment precludes allowing it to become captive in a single stope therefore access to stope levels is now almost exclusively by ramps rather than by raises the demand for ventilation has also been increased because of the increased usage of diesel equipment in recent years production with remote controlled loader operation has been replacing the more costly development of gathering bells and chutes that have been commonly used in blasthole stope mines paste fill mining fill systems based on the use of paste for backfill are supplanting other fill systems for bulk backfill applications the opportunity to transport and place the material by pipe reticulation substantially decreases congestion in drifts and ramps the placement rate of paste fill is much faster than placement using trucks carrying fill from a surface or underground batch plant and then having to rely on lhds and jammers to place the fill into the stope the resulting benefits lower operating costs lower ventilation demands and a faster turnaround time for the filled stope are similar to those achieved by hydraulic systems but without the water drainage issues however paste plants are costly to construct and must be completed during preproduction development cemented rock fill t

he use of cemented rock fill continues to be a common choice where bulk material handling is not advantageous and where flexibility is important to maintaining operations as indicated in the history section the different materials that have been used to backfill mine openings include the following dry sand and rock fill uncemented hydraulic backfill cemented hydraulic backfill cemented rock fill paste fill pneumatic fill flowable fill in subsequent sections of this chapter each of these alternatives will be discussed for each the method of placement drainage requirements if any equipment requirements pipe wear and other relevant design and operating topics are considered dry sand and rock fill dry placement of stope fill materials is often the simplest method of fill placement and has a long history in metal mines in north america and australia that used cut and fill or square set stoping this method dominated until the 1950s and was still taught as a preferred method of backfilling in some mining schools until the late 1960s advantages and disadvantages although it has generally been superseded by other methods there are still situations where dry placement of fill has advantages over other systems for example if a mine does not have a processing plant on site to provide mill tailings other fill options must be considered such fills could include 1 natural surface sands glacial till in north america natural dune sands in australia 2 open cut overburden specifically quarried surface rock or even rock generated in local civil works and 3 underground development waste dry fill is also the preferred option in evaporite and gypsum mines where the ore or host rock is readily dissolved in water or in mines in arid locations where water is scarce dry placement of fill has two main disadvantages first it has a relatively low density as placed which means it can undergo significant compression before it reaches its optimum density to provide stope wall support to prevent spalling of stope walls and minimize stope wall convergence second depending on stope configuration it may be difficult to place dry fill close to the roof of the opening across the full span of the stope roof transport and placement of dry fill the first consideration as with all types of backfill is the means of transport of the dry fill underground the most efficient means for transport of sand or gravel sized particles is a borehole from the surface selection of the size of the borehole depends on two considerations the required throughput in metric tons per hour and the size distribution of the particles for coarse material such as gravel the particle size will govern because the diameter of the borehole should as a general but not necessarily universally applicable rule be a minimum of three to six times the maximum particle size of the gravel in order to avoid hang ups for finer material the governing concern is

the required throughput in general boreholes with diameters of 150 mm or 200 mm 6 or 8 in will be sufficient for most purposes however using this size of pipe one must never let the pipe fill and possibly bridge as it may plug the most common method of transporting dry fill to the workings is by haulage vehicles in conventional older mines this usually meant railcars which would dump the fill into a fill raise more recently haulage trucks have been used and belt conveyors are also being used for sand gravel or rock all of these placement methods require good access to the workings that need to be filled another important consideration with dry placement of granular fill is the method used to place and compact the fill within the stope historically placement was done by hand which is slow and costly if the area to be filled is accessible to mobile equipment spreading and compaction could be accomplished by means of a hydraulic ram and blade mounted on a haulage truck or load haul dump lhd vehicle the blade is pushed laterally against the fill pile by the hydraulic ram to compact it another placement and compaction method that has been tried is a high speed throwing belt the main disadvantage of which is the limited distance that the fill travels beyond the discharge point in the case of rock fill placement is not generally a problem if dumping is from the top of the working because of the mass of the rock pieces gravity flow is usually satisfactory however large blocks contain a significant proportion of voids and such fill may need considerable consolidation before it can provide significant stope wall restraint this disadvantage is often resolved by filling the voids with cemented hydraulic fill as discussed later in this chapter uncemented hydraulic backfill system design hydraulic transport and placement of backfill is the method most commonly used in metal mines but less commonly in coal mines deslimed and partially dewatered mill tailings are the materials most commonly transported underground hydraulically but surface sand e g glacial till in north america natural dune sands in australia is also transported in this manner the fine slimes portion of tailings is undesirable because of its slow drainage but also because of the potentially disastrous consequences of liquefaction resulting from vibrations from for example production blasts transport of slurry solids content mill tailings generally consist of a heterogeneous slurry with approximately 30 40 solids by mass and the rest water the solids and the water represent two separate phases that will readily separate if the flow velocity is insufficient when a heterogeneous slurry is discharged from a horizontal pipe the solids will settle out with the coarsest particles settling first as the solids are discharged they will gradually form a cone around the end of the pipe with the finer materials toward the toe of the slope the

slope of the cone varies with the solids content of the slurry with very dilute slurries the slope may not exceed 2 3 and the sediments will be loosely packed but at solids contents of 60 70 by mass or more the slope will approach the angle of repose of the particles and the solids will consolidate readily as they drain hydraulic backfill systems typically operate at solids contents of 65 70 solids by mass or more because as stated previously the tailings in the mill typically have a solids content of around 30 by mass it is generally necessary to partially dewater the tailings prior to transport underground dewatering is usually accomplished using hydrocyclones or occasionally thickeners or rake classifiers hydrocyclones are vastly preferable because they most effectively deslime as well as dewater one word of caution the cut point commonly used by geotechnical engineers for drainage computations is the d10 size whereas the d20 or d50 size is used by designers of hydrocyclones failure to appreciate this distinction could result in having a larger fines content than anticipated in the backfill it is critical to maintain the design flow rate in a hydrocyclone if the cut point is to be achieved in practice drainage requirements after hydraulic backfill has been placed in a mined opening it is necessary to drain away the water for a number of reasons first water bearing sand slime sized material especially if it contains excessive slimes and or excessive water can liquefy if subject to vibration caused by a seismic disturbance such as an earthquake by a rock burst or by production blasting second drainage of the water is essential to consolidation of the fill because the fill is used as the working floor for the next lift in overhand cutand fill stoping it is generally desirable to have it available for production traffic within 24 hours prevention of the phenomenon of liquefaction requires essential consideration in the preparation and placement of hydraulic fill particularly uncemented hydraulic fill underground a general but not absolute rule is that the gravitational percolation rate of water through the fill should exceed 10 cm h 4 in h because drainage of water from the fill is by gravity with a hydraulic gradient of unity this percolation rate corresponds to a hydraulic conductivity of approximately 3 10 3 cm s 5 9 10 3 fpm which is a typical value for a well graded fine grained sand to ensure this drainage rate is achieved it is customary to remove the fines so that no more than 10 of the solids are finer than 74 m 200 mesh however satisfactory backfilling has been achieved with as much as 10 of the solids finer than 43 m 325 mesh also hydraulic conductivity is temperature dependent so as mine water temperature increases fines content can increase and drainage rates remain satisfactory importantly the converse applies a comprehensive discussion of hydraulic fill drainag

e mechanisms is given by thomas et al 1979 to facilitate fill drainage it is customary to erect fences within the stopes being filled to contain the fill and to drain the water through ladderways or drain towers constructed up through the fill to allow the fill water to drain without loss of fill fines it is customary to place filter fabrics on the inside of fill fences and outside drainage ladderways and towers inside the site to be filled historically burlap in north america or hessian in australia was used as the filter fabric more recently these materials have been largely replaced by monofilament geotextiles though the traditional materials where available are still acceptable to select the correct geotextile the following filter selection criteria developed by the u s army corps of engineers may be used the ratio of the d85 85 finer size of the backfill to the opening size of the geotextile should exceed 1 0 the open area of the geotextile should be less than 40 of its total area for most tailings the d85 size will range from 149 to 297 m 0 0069 to 0 012 in selection of the geotextile should also consider the required water flow rate per unit area of geotextile which depends on the dimensions of the fill fences and drainage towers backfill transport pipe size requirements with twophase solids fluid heterogeneous slurries the transport velocity in a pipeline must be high enough to prevent deposition of coarse particles on the other hand an excessive transport velocity in the pipe will result in excessive friction losses and pipe wear a flow velocity of 1 5 m s 5 fps is often but not always considered the minimum necessary to provide satisfactory results conversely a flow velocity of greater than 2 4 m s 8 fps may result in excess friction losses and pipe wear table 13 6 1 presents hourly placement tonnages that can be achieved at a pipe velocity of approximately 1 8 m s 6 fps and 65 70 solids density by mass for a material with a specific gravity of 2 65 pipe friction losses and pump requirements in mines accessed by a shaft transport of backfill underground will be by means of either a pipe in the shaft or a borehole near the shaft preferable to a pipe in the shaft depending on shaft function that connects the backfill plant to the workings in such cases the elevation difference is usually but not always sufficient to transport the fill to the working place without pumping however where access to the mine is by means of a horizontal adit gravity flow will with the occasional exception not be sufficient to transport the backfill slurry and pumping will be necessary where mine access is by decline pumping may or may not be necessary depending on the general mine layout in all cases it is prudent to calculate the friction losses in the system backfill plant requirements the first requirement in a typical backfill plant is a bank of hydrocyclones to remove excess water an

d fine particles from the tailings the overflow from the cyclones the fines and excess water is typically transported to a surface tailings impoundment for disposal many mines once operated on a 5 day week whereas concentrators typically operate on a 7 day week and in such cases it became necessary to provide a storage silo capable of holding backfill pouring requirements for 1 or 2 days this silo is known as a repulper because the backfill is allowed to drain until needed at which time water is added at high pressure to the base of the tank to refluidize the fill the fill plant will also need transfer pumps and valving and care should be taken to ensure that the capacities of the transfer pumps are sufficient transfer pumps should be able to handle a flow rate approximately 20 greater than the design flow rate for the system solids plus water because if the flow rate decreases below the critical value for sedimentation the inevitable result will be the complete filling of the pipe with solids cleaning out a pipe that has sanded up is time consuming and costly particularly if the fill contains portland cement current practice is more likely to have both mine and mill operating around the clock thus eliminating the need for fill storage tanks and the complexities they involve cemented hydraulic backfill system design correctly prepared uncemented hydraulic fill is predominantly sand sized and must behave as a sand albeit on occasions a slow draining sand consequently it has no true cohesion or until drained no unconfined compressive strength for some applications such as fill placed against unmined rock that will be recovered later it is often desirable that the fill be able to stand unsupported upon exposure for such applications it is customary but on occasions not at all necessary and possibly detrimental to add cement to the backfill because cemented fill is the same fill material as the uncemented fill except for the cement several of the design parameters for a cemented hydraulic backfill system are the same as for an uncemented backfill system however slurry solids content must be more carefully controlled though pipeline velocities remain essentially the same as is the case for uncemented fill the tailings are deslimed with a typical d10 size of 74 m 200 mesh however the cement provides fines to the mix and the drainage rate is typically somewhat slower than that for uncemented fill this potential problem is in part negated for the following reasons 1 portland cement consumes fill water during hydration reducing the volume of water to be removed 2 portland cement produces heat during hydration reducing the viscosity of fill water tending to increase dewatering rate and 3 portland cement lubricates the flow so higher placement densities can be used reducing the amount of water to be removed in pipeline flow portland cemented fill flows more uniformly than uncemented hydraulic fi

ll and this tends to reduce pipeline wear for similar placement rates coarse fill particles have reduced access to pipeline walls optimum solids content for pipeline flow generally ranges from 68 to 72 solids by mass and closer attention is given to control of this solids content costs considered after portland cement is added to hydraulic fill the fill operation moves from underground disposal of mining waste to preparation of a quality controlled engineering material the amount of cement required in a cemented fill depends on the strength requirement and the use for the fill some examples are given in table 13 6 3 these should be used for guidance and not adopted as standard for example use of cement in mucking floors may have advantages for movement of mucking equipment but it prevents almost totally the downward flow of fill drainage water pump requirements where there is extensive horizontal development on upper levels of a mine the elevation head may be insufficient to support flow by gravity alone in such cases pumps will be necessary to transport the fill over all or part of the underground fill distribution system both positive displacement pumps and centrifugal pumps can be used for this application however rubber lined centrifugal slurry pumps are considered a better choice pumps with dry glands which minimize water handling requirements are available with centrifugal pumps the slurry will flow through a pump even if it is not needed to supply head backfill plant requirements backfill plant requirements are similar to those for uncemented fill except for the need for a cement silo plus mixing tanks and feeders cemented rock fill system design cemented rock fill as typical of modern practice was developed at the geco division of noranda mines ltd in the early 1960s to provide support for stope walls that were prone to slough by restraining stope wall rock the fill reduced dilution and improved overall regional stability a cemented rock fill system was developed at mount isa in australia in the late 1960s but the main reason for the rock fill component was to reduce the quantities of cemented hydraulic fill as portland cement was expensive in that part of australia being remote from cement suppliers and prone to disruption of supply during the wet season in a cemented rock fill system the main design function of the cemented hydraulic fill is to fill the voids in the broken rock binder placement the stiffener in a cemented rock fill can consist of deslimed mill tailings fill cemented deslimed mill tailings fill paste fill portland cement fly ash or other pozzolan or a combination of two or more of these all of these binders are transported most easily by pipeline to be added to the rock fill as it enters the top or some more appropriate location in the stope standardization of such practice is difficult and large stopes must be considered on a case by case basis paste fill system design wit

h paste fill the tailings become a high density slurry with a solids content of 75 80 by mass the slurry acts as a viscous bingham solid for which a critical shear stress must be exceeded before the material will flow the critical shear stress for paste fills typically ranges between 250 and 800 pa 0 041 to 0 116 psi boger et al 2006 because of the viscous nature of the paste and the large amounts of energy required to transport paste in the turbulent regime it is customary to transport paste in the laminar flow regime paterson 2006 the drawback to laminar flow however is the potential for sedimentation and blockage in the pipeline though paste fill because of its tightly controlled production and relatively high pulp density is less prone to settling than conventional deslimed mill tailings fill two other important attributes of paste fill are 1 the water is bound into the slurry so that little drainage is required and 2 the fact that fines removal from tailings may not be required pipe wear few data are available regarding the wear of pipes transporting paste fill however because the paste moves as a viscous unit it is expected that the wear rates are less than for conventional hydraulic fills the wear can be checked in vertical sections of the pipelines using a three legged caliper probe it is recommended that readings be taken after 15 000 t 16 535 st have been transported and thereafter at 50 000 t 55 116 st intervals backfill plant requirements components in a paste fill plant typically include tailings storage tanks with a 1 day live storage capacity paste thickeners for reducing water content from 65 by mass to approximately 20 by mass silos for cement and possibly fly ash storage weightometer hoppers for tailings cement and fly ash if used mixers for blending tailings fly ash and cement if used and positive displacement pumps pneumatic stowing pneumatic stowing is often used in situations where hydraulic filling has significant drawbacks examples are situations where it is undesirable to handle the large quantities of water that drain off from the fill or where seams are thin and flatlying so that hydraulic placement of fill tight to the roof is difficult to achieve pneumatic stowing has also been considered as a means of placing broken rock whose ore grade is too low for recovery by conventional means in order to extract a mineral by pressure or flood leaching particle size distribution the particle size distribution will affect the velocity necessary to transport the solids and the pipe size to minimize the possibility of bridging of large chunks the pipe diameter should be at least three times that of the largest chunk carried as a practical matter the maximum lump size should be 100 mm 4 in or less with 75 mm 3 in preferred considered another way typical stowing pipe diameters range from 150 mm 6 in up to 225 mm 9 in consequently the maximum lump size s

hould preferably be less than 75 mm 3 in and ideally 50 mm 2 in or less a smaller maximum particle size will also result in reduced velocity power requirements and pipe wear feeder and blower considerations a pneumatic stowing system has three basic components 1 a hopper and feeder for the solids 2 an air supply consisting of compressed air from either a pipeline or compressor or a rotary blower 3 an injection pipe and nozzle the hopper should be large enough to contain a bucket load of broken material from a typical haulage vehicle the discharge opening in the hopper should have a minimum dimension of at least three to six times the dimension of the largest piece being transported the feeder should be sized to handle the expected hourly throughput of the system three important considerations for the feeder which is generally a rotary airlock type are the number of vanes on the rotor diameter of the airlock and the rotor rpm for a given flow rate in cubic meters per hour better flow characteristics are obtained with a larger diameter airlock operated at a lower rpm six or eight vane models are available in some applications difficulties have been experienced with the eight vane units except for very large tonnage operations low pressure systems using positive displacement blowers provide satisfactory air quantities for stowing these blowers are noisy and should be fitted with silencers injection pipes are typically 150 225 mm 6 9 in in diameter air velocities in the stowing pipe should exceed 20 m s 65 fps and the exit velocity in the nozzle should be about 40 m s 130 fps for optimum results stowing machines are available to place up to 280 m3 h 367 yd3 h solids loading and system pressure drop an important parameter is the solids to air ratio by weight for most stowing applications this ratio will range from 10 to 40 however the actual value of the ratio for a given stowing operation will depend on the quantity and size of the solids to be stowed given that with coarse material the velocity required to keep the solids in the air stream will be higher than for fine material other considerations remote placement of backfills backfilling of former mines as a subsidence prevention measure is becoming increasingly common in general it will be necessary to provide multiple pour points to ensure the void will be as completely filled as possible similar to discharge from a stacker or conveyor the solid discharge from a fill pipe will form a cone that is characterized by finer and finer particles as it moves away from the discharge point the angle of repose of the cone depends on the solids percentage in the slurry the higher the percent solids the steeper the cone because a supply bore will become useless after the cone fills to the top of the opening the spacing between supply borings depends on the height of the opening as well as the percent solids in the fill automated systems because of the

generally batch nature of filling and the need to change discharge locations during the course of backfill placement not all facets of backfilling lend themselves to automatic operation however some portions of the cycle such as starting and stopping of pumps and opening and closing of hopper gates can be automated by means of interlocks level controls and other automatic controls for example the manitoba division of vale inco installed an automated fill pumping system at the thompson mine in the 1960s in theory the valving on the discharge locations could be automated and interlocked with level switches placed below the discharges however the result would be a pouring system in the stope that would be significantly more difficult to install and dismantle effect of cemented fill in mill feed on processing plant recovery when mucking on a fill floor it is inevitable that some fill will be mixed with the broken ore if the fill is uncemented the only negative effect is some dilution however if the fill is cemented there are likely to be unwanted effects on the concentrator circuit especially in the case of a silver lead zinc ore specifically the cement will significantly increase the ph of the process water because the ph of hydrous cement is approximately 10 5 11 in addition the presence of calcium carbonate may chemically affect milling processes occasionally the presence of cement will improve mill recovery percentages heat generation of sulfide backfills tailings containing pyrite and especially some forms of pyrrhotite readily oxidize giving off heat and absorbing oxygen as a consequence it is not uncommon for the temperature in a stope to rise by 3 5 c 5 10 f after the first lift of fill has been placed however the oxidation process may also result in additional cementing of the fill noranda took advantage of this at the horne mine in quebec when designing the hydraulic backfill system there in the 1930s patton 1957 one unfortunate consequence of this was loss of life by removal of oxygen from mine air also it is reported that in the kimberley mine in british columbia iron sulfides in placed rock fill increased fill temperatures to such an extent that fill behind bulkheads became molten the expression cave mining will be used in this chapter to refer to all mining operations in which the ore body caves naturally after undercutting and the caved material is recovered through drawpoints the term encompasses block caving panel caving inclined drawpoint caving and front caving caving is the lowest cost method of underground mining provided that drawpoint size and handling facilities are tailored to suit the caved material and that the extraction horizon can be maintained for the life of the draw factors affecting caving operations the 25 parameters that should be considered before implementing any cave mining operation are set out in table 13 7 3 the parameters in bold are a function of the pa

rameters that follow many parameters are uniquely defined by the ore body and the mining system and are not discussed further the parameters considered later are common to all cave mining systems and need to be addressed if any form of cave mining is contemplated cavability monitoring many caving operations has shown that two types of caving can occur stress and subsidence caving however it is better to use the terms vertical extension to mean upward propagation of the cave and lateral extension to mean the propagation of the cave as a caved block is expanded vertical extension caving occurs in virgin cave blocks when the stresses in the cave back exceed the strength of the rock mass caving can stop when a stable arch develops in the cave back the undercut must be increased in size or the boundaries weakened to induce further caving high horizontal stresses acting on steep dipping joints increase the mrmr this was the situation in block 16 at the shabanie mine in zimbabwe a stable back was formed when the undercut had a hydraulic radius of 28 and the mrmr was 64 when block 7 adjacent to block 16 caved the horizontal confining stress was removed which resulted in a reduction in the mrmr to 56 in block 16 at this point caving occurred lateral extension caving occurs when lateral restraints on the block being caved are removed by mining adjacent to the block this often results in a large stress difference leading to rapid propagation of the cave and limited hulking all rock masses will cave the manner of their caving and the fragmentation need to be predicted if cave mining is to be implemented successfully the rate of caving can be slowed by control of the draw since the cave can propagate only if there is space into which the rock can move the rate of caving can be increased by a more rapid advance of the undercut but problems can arise if an air gap forms over a large area in this situation the intersection of major structures heavy blasting and the influx of water can result in damaging air blasts rapid uncontrolled caving can result in an early influx of waste in conventional layouts the rate of undercutting ru should be controlled so that the rate of caving rc is faster than the rate of damage rd due to abutment stresses thus rc ru rd however in areas of high stress the rate of caving must be controlled to maintain an acceptable level amount of seismic activity in the cave back otherwise rock bursts can occur in suitably stressed areas pillars and rock contacts as advanced undercutting will be used in these situations damage to the undercut and production levels will not be a problem the stresses in the cave back can be modified to some extent by the shape of the cave front numerical modeling can be a useful tool for helping an engineer determine the stress patterns associated with possible mining sequences an undercut face that is concave toward the caved area provides better control o

f major structures in ore bodies having a range of mrmrs the onset of continuous caving will be in the lower rated zones if they are continuous in plan and section this effect is illustrated in figure 13 7 3b where the class 5 and 4b zones are shown to be continuous in figure 13 7 3a the pods of class 2 rock are sufficiently large to influence caving and cavability should be based on the rating of these pods good geotechnical information as well as information from monitoring the rate of caving and rock mass damage is needed to fine tune this relationship particle size distribution in caving operations the degree of fragmentation has a bearing on drawpoint spacing dilution entry into the draw column draw control draw productivity secondary blasting and breaking costs and secondary blasting damage the input data needed to calculate the primary fragmentation and the factors that determine the secondary fragmentation as a function of the caving operation are shown in figure 13 7 4 primary fragmentation can be defined as the size distribution of the particles that separate from the cave back and enter the draw column the primary fragmentation generated by subsidence caving is generally more coarse than that resulting from stress caving the reasons for this are 1 propagation of the cave occurs more rapidly 2 the rock mass disintegrates primarily along favorably oriented joint sets and 3 there is little shearing of intact rock the orientation of the cave front or back with respect to the joint sets and the direction of principal stress can have a significant effect on the primary fragmentation secondary fragmentation is the reduction in size of the original particles as they move through the draw column the processes to which particles are subjected determine the size distribution that reports to the drawpoint a strong well jointed material can result in a stable particle shape at a low draw height draw control draw control requirements and applications are shown in figure 13 7 11 and table 13 7 4 respectively the grade and fragmentation in the dilution zone must be known if sound draw control is to be practiced figure 13 7 12 shows the value distribution for columns a and b both columns have the same average grade of 1 4 but the value distribution is different the high grade at the top of column b means that a larger tonnage of waste can be tolerated before the shutoff grade is reached in lhd layouts a major factor in poor draw control is the drawing of fine material at the expense of coarse material strict draw control discipline is required so that the coarse ore is drilled and blasted at the end of the shift in which it reported in the drawpoint it has been established that the draw will angle toward less dense areas this principle can be used to move the material overlying the major apex by creating zones of varying density through the differential draw of lines of drawbells dilution for

the purposes of block caving the percentage of dilution is defined as the percentage of the ore column that has been drawn before the waste material appears in the drawpoint it is a function of the amount of mixing that occurs in the draw columns mixing is a function of ore draw height range in fragmentation of both ore and waste drawzone spacing and range in tonnages drawn from drawpoints the range in particle size distribution and the minimum drawzone spacing across the major apex will give the height of the interaction zone hiz this is illustrated in figure 13 7 13 there is a volume increase as the cave propagates so that a certain amount of material must be drawn before the cave reaches the dilution zone the volume increase or swell factors is based on the fragmentation and is applied to column height typical swell factors for fine medium and coarse fragmentation are 1 16 1 12 and 1 08 respectively layouts eight different horizontal lhd layouts and two inclined drawpoint lhd layouts are used at various operating mines an example of an inclined drawpoint lhd layout is shown in figure 13 7 17 the el teniente chile and henderson mine layouts are shown in figures 13 7 18 and 13 7 19 respectively numerical modeling of different layouts showed that the el teniente layout was the strongest and had several practical advantages for example the drawpoint and drawbell are on a straight line which results in better drawpoint support and flow of ore if there is brow wear the lhd can back into the opposite drawpoint the only disadvantage is that the layout does not have the flexibility to accommodate the use of electric lhd machines as does the herringbone layout a factor that needs to be resolved is the correct shape of the major apex it is thought that a shaped pillar will assist in the recovery of fine ore also in ore bodies characterized by coarse fragmentation there is less chance that stacking will occur figure 13 7 20 the main area of brow wear is immediately above the drawpoint if the vertical height of the pillar above the brow is small failure of the top section will reduce the strength of the lower section and result in aggravated brow wear more thought must be given to the design of lhd layouts to provide a maximum amount of maneuver room within a minimum size of drift opening thus larger machines can be used within the optimum drawzone spacings another aspect that needs attention is lhd design that is to reduce the length and increase their width although the use of large machines might be an attraction it is recommended that caution be exercised and a decision on machine size be based on a correct assessment of the required drawzone spacing in terms of fragmentation the loss of revenue that can result from dilution far exceeds the lower operating costs associated with larger machines undercutting undercutting is one of the most important aspects of cave mining because not only is a c

omplete undercut necessary to induce caving but the undercut method can reduce the damaging effects of induced stresses the normal undercutting sequence is to develop the drawbell and then to break the undercut into the drawbell as shown in figure 13 7 18 in environments of high stress the pillars and brows are damaged by the advancing abutment stresses the henderson mine technique of developing the drawbell with long holes from the undercut level reduces the time interval and extent of damage associated with post undercutting to preserve stability the henderson mine has also found it necessary to delay the development of the drawbell drift until the bell must be blasted figure 13 7 19 the damage caused to pillars around drifts and drawbells by abutment stresses is significant and is the major factor in brow wear and excavation collapse rock bursts also occur in these areas the solution is to complete the undercut before development of the drawpoints and drawbells the advanced undercut technique is shown in figure 13 7 21 in the past the height of the undercut was considered to have a significant influence on caving and possibly the flow of ore the asbestos mines in zimbabwe had undercuts of 30 m with no resulting improvement in caving or fragmentation the long time involved in completing the undercut often led to ground control problems good results are obtained with undercuts of minimum height provided that complete undercutting is achieved where gravity is needed for the flow of blasted undercut ore the undercut height needs to be only half the width of the major apex this results in an angle of repose of 45 and allows the ore to flow freely support requirements in areas of high stress weak rock will deform plastically and strong rock will exhibit brittle often violent failure if there is a large difference between the rmr and mrmr values yielding support systems are required this is explained in figure 13 7 22 prestressed cables have little application in underground situations unless it is to stabilize fractured rock in a low stress environment the need to constrain the rock laterally and for lining surfaces such as concrete cannot be too highly emphasized support techniques are shown in table 13 7 5 the use of cable bolts in brows is common practice but often these bolts are not installed in a pattern that takes into account joint spacing and orientation cable bolts should not be installed in highly jointed ground as the bolts do not apply lateral restraint to the brow and serve only to hold blocks in place conclusions cavability can be assessed provided accurate geotechnical data are available and geological variations are recognized the mrmr system provides the necessary data for an empirical definition of the undercut dimension in terms of the hydraulic radius numerical modeling can assist an engineer in understanding and defining the stress environment fragmentation is a major factor in asse

ssing the feasibility of cave mining in large competent ore bodies programs are being developed for predicting fragmentation and even the less sophisticated programs provide good design data the economic viability of caving in competent ore bodies is determined by lhd productivity and the cost of breaking large fragments drawpoint and drawzone spacings for coarser material need to be examined in terms of recovery and improved mining environments spacings must not be increased to lower operating costs at the expense of ore recovery the interactive theory of draw and the diameter of an isolated drawzone can be used in the design of drawzone spacings complications occur when drawzone spacings are designed on the basis of primary fragmentation and the secondary fragmentation is significantly different the drawpoint is the only control that the operator has once caving has started the drawpoint should be commissioned as soon as possible after the ore body is undercut it is therefore essential that every effort is put into ensuring the stability of the drawpoint the support system must be designed for the rock mass strength and draw height when it is planned to draw high column heights the extra expense of ensuring drawpoint integrity is essential major drawpoint repairs in a horizontal layout tend to disrupt the draw pattern and can lead to column loading block caving has advantages over large open pits in environmentally sensitive areas the crater is much smaller and if necessary can be filled with waste rock even during drawdown when the operation is completed the rehabilitation costs will be considerably less than with a large open pit longwall mining is a technique that evolved from a largely manual method in which rows of individual roof props supported the roof along a long face the ore was broken from the face first by drilling and blasting and later by mechanical means powered by pneumatic hydraulic or electrical energy ore loading was done by hand shoveling into a conveyor or rail cars at the face this method was labor intensive lacked productivity and exposed workers to significant hazards since then technology evolution has produced spectacular improvements motivated by the need for improved safety and productivity as well as reduced labor and production costs this evolutionary process has resulted in modern longwall systems which are the focus of this chapter the configuration and major components of a modern longwall are shown in figure 13 8 1 the shearer traverses the face excavating the ore within a defined extraction height the mined material is loaded onto the armored face conveyor afc by the shearer and is transported to the main belt conveyor system via the stageloader the stageloader normally has an integral crusher to provide suitably sized material for conveyor belts the shields advance sequentially following the shearer to hold up the roof directly above the face equipment and advance

the afc to repeat the cutting cycle the excavated area behind the shields is allowed to collapse which limits the roof load that the shields must support and starts a process that may create surface subsidence high capacity longwall mining systems depend on the breaking action of a machine to mine the ore as such modern longwalls are used to produce ores that can be mechanically cut or broken without explosives high capacity longwalls mainly produce coal but examples also exist in trona potash and phosphate rock the potential is being investigated to apply longwall technology to platinum and gold reef deposits with equipment adapted to the requirements of stronger rocks longwall applications longwall mining has evolved mostly in response to coalmining applications because many coal mines yield a relatively low value product particularly when coal is used for electric power generation the output of the longwall must be a sustained stream with low cost on a unit of production basis the desirable image often suggested is that of a coal faucet which can be opened or closed as needed embedded in the image is the ability to adjust the flow to desired levels it should be the goal of those involved in the design operation and maintenance of a modern longwall to fulfill the goals of sustainability and productivity suggested by the coal faucet image it is notable that health and safety considerations are intrinsically interwoven into longwall design and operating concepts because they are fundamental elements of sustainability and success it is customarily accepted that longwall mining is the safest and most productive mining method that can be applied to soft rock deposits longwall systems are serially dependent processes this is because all of the associated equipment that supports longwall operation must function simultaneously at the capacity of the longwall to permit its operation this dependence is a particular issue for the outby ore transport method typically conveyor belts frequently shortfalls in outby conveyor belt performance are the most limiting constraint on longwall performance the important point is that the success of the longwall system depends on the success of each of its constituent components not just the machinery composing the longwall this principle extends even to mine development activities that create the panel where the longwall will operate in the future longwalls are typically designed for specific considerations related to the target deposit notable among these considerations are those described in the following sections deposit depth longwall mining has been applied to deposits with very shallow depth of cover perhaps less than 50 m in a few cases and up to 1 600 m at the other extreme applications in the 100 600 m depth are common for longwall coal mines especially in the united states and australia the loads on longwall roof supports shields are largely independent of depth of c

over shield loads are more related to the extraction height and the strata overlying the ore bed seam than to any other factor mining induced stresses on the face front abutment load and gate road pillars side abutment load are depth related not only must gate road pillars demonstrate expected stability but the roof in conjunction with the primary and secondary supports must also be stable floor lift floor heave is frequently encountered at increased depth or with weak materials in the floor i e under clay fire clay or claystone peng 2006 deposit dip most longwall mining systems are applied where the inclinations along the longwall face or the panel axis are low less than 8 5 although longwalls are being used at much steeper inclinations these are not consistent with high productivity steeply inclined longwalls also require special adaptations to stabilize the face against downhill creep shield toppling and product transport issues steep seam gate road development is also problematic steep thick seams offer more opportunity for success but are still difficult as conveyor belt inclination increases so do the issues in its operation up to a practical limit and depending on the material being conveyed for run of mine rom coal this limit is in the 15 range cema 1997 this value can be even less possibly 12 if the material conveyed is wet the wet condition can arise when production volumes diminish but water application is unchanged in another case large volumes of water can be transported by a submerged afc onto the conveyor belt system importantly operation of either conveyor belts or afcs is easier when material flow is uphill numerous difficulties arise when conveying material down an appreciable grade which far overwhelms any energy benefits that might be realized evans 1993 the relative layout of the face and progression of panels in the mining sequence with respect to dip is important because of concerns about water migration methane accumulation and toppling of face and rib materials deposit thickness longwall systems are typically applied to laterally extensive tabular deposits with limited variation in ore thickness after the ore thickness and its variation are characterized within the area to be mined a nominal mining height will be defined practical minimum and maximum mining heights are defined by critical equipment sizes and operating envelopes the longwall has little flexibility in extraction of material within the bounds of a panel and will cut through any area of the panel extracting at least its minimum operating height the only alternative is to bypass undesirable areas within the panel by withdrawing the longwall equipment and reinstalling it outside the undesirable area this is a costly and time consuming process chosen only after considerable planning and reconnaissance where ore thickness is greater the longwall can extract up to the maximum cutting height of

the shearer or near the fully extended height of the shields caution must be used when approaching the fully extended height of the shields to avoid a condition where the shields fail to apply and maintain full loads into the roof without roof failure mining equipment technology has the consequence of defining mining height in the following approximate intervals of deposit thickness t thin t 1 75 m moderate 1 75 m t 3 75 m thick 3 75 m t 7 25 m very thick t 7 25 m thin seam systems have historically applied longwall plow technology but have not achieved the productivity of thickerseam shearer based applications most modern high capacity longwalls operate in the moderate height category from 1 75 to 3 75 m optimum height is generally agreed to be in the 2 5 4 0 m range for single pass longwalls above 3 75 m and up to approximately 7 25 m mining height the thick category is the upper limit of single pass longwall capability in this range normal longwall equipment requires numerous adaptations to perform successfully until it reaches a practical limit at approximately 7 25 m above 7 25 m extraction thickness longwall technology is forced into several alternatives multilift longwall or top coal caving on a worldwide basis as very thick coal seams extracted by surface mining extend to depths beyond economic limits underground mining methods will be required that offer high resource recovery such as longwall top coal caving ideally a tabular deposit well suited to longwall mining would exhibit uniformity in the ore bed coal seam as well as the adjacent roof and floor unfortunately examples of such ideal conditions are few generally discontinuities result from structural or depositional origins but may also be the result of post depositional events structural geologic disturbances normally involve folding and faulting gentle inclinations induced by folding can be considered as previously described in the discussion on dip however if the gentle dips of a broad syncline anticline or monocline go over approximately 8 5 as a trend this may limit applicability of the longwall expectations of high productivity may decline as the inclination steepens from 8 5 up to 15 aside from overall inclination the rate of change of inclination becomes a constraint as the afc shearer and shields all have different concerns in this regard as an example if the radius of curvature of seam topography gets short and the shearer is able to follow it the afc may have less vertical articulation than required to follow the changed topography and bridge forming a span that is not supported by the floor thus the afc would only be supported by the connections that hold adjacent pan sections together dog bones or dumb bells by the afc s own weight or with the additional weight of the shearer the afc connections may fail or the pan structure may fail this can cascade into an afc system failure such as a c

hain wreck where afc chains break flights are damaged and main drive sprockets may fail such incidents may take hours to days to repair and cost in the hundreds of thousands of dollars in direct and lost opportunity costs they also create avoidable safety hazards in an alternate case shields can lift their bases off the floor during cyclic advancement if their canopies become iron bound with adjacent shields because of a short radius of curvature in extraction profile this condition can result in a suspended shield which can fall to the floor unexpectedly and be a safety hazard of course all of these problems are created when the equipment is asked to do something that was never intended by the original equipment designer or conversely was not identified as a performance requirement in the acquisition specifications figure 13 8 2 shows a disturbed area that was successfully mined with a longwall but not without impact to productivity another structural geology concern is fracturing or faulting an example might be fracturing along the axis of local bed flexure which can adversely affect roof control or cause elevated gas or water inflows notably fractures faults parallel to hillsides exist in some areas and may complicate logical panel layouts beneath a ridge line faulting may be present with or without appreciable folding strata displacements may be normal faults and total displacements may be distributed across one or more discrete fault traces within a short distance this is generally a manageable issue so long as the local displacement is not too large because these features are preferably traversed at an acute angle to the face grading roof and floor into and out of the feature makes operating through local offsets feasible up to double the mining height in displacement this will introduce an adverse impact to rom product quality and overall productivity obviously proximity of faults on a single face is problematic as is the case where faults still at acute angle to the face exhibit enduring persistence both of these cases may require revised rom quality and productivity estimates then there are those cases where the fault is at a shallow angle with the face line or possibly even parallel to the face at worst case this scenario can have seriously negative impacts to productivity and may lead to premature face recovery or even threaten the loss of equipment if such a feature is suspected it should be approached with caution after careful risk assessment and with a plan of mitigating actions in place if such a feature is first recognized as it emerges onto the face it is certainly appropriate to make efforts to evaluate it and implement a mitigation strategy before exposing a large proportion of the face to the hazard it has frequently been the case that a longwall could outrun a local adverse roof condition however that is not true for faults parallel or at shallow angle to the face in the case o

f most structural faulting the ground stresses within one block of the fault may not be equivalent to the stresses in others although normal faulting is at some risk in this regard it is particularly problematic in thrust faults the stress differential between blocks in a thrust faulted scenario can result in some blocks having extremely difficult ground control issues compared to others depositional discontinuities develop in response to conditions present at the time that the mineral deposit formed these discontinuities commonly arise in conjunction with sand channels above or below a coal seam or as scours also termed faults or wants in some locales where they erode and replace the coal seam figure 13 8 3 depicts sand channels as they may commonly occur in coal bearing strata in near proximity to these features it is possible to have elevated stresses and differential compaction surfaces these glassy smooth discontinuities slickensides are weak in tension and can constitute a hazard to roof control where large thicknesses of coal occur over a short stratigraphic interval compactional features can exist as faults with displacements mostly confined in near proximity to the seam and having normally short lengths perhaps less than 250 m compactional faults are not normally the source of great concern sand channels above the mining horizon can lead to increased shield set load and yield load requirements and should be evaluated accordingly where these features might be encountered at a shallow angle to the face opportunities to avoid the condition are initially preferred when near seam sand bodies are strong and stiff in conjunction with significant depth the possibility of sudden strain energy release in response to mining can be a concern another variation of depositional discontinuity occurs when soft sediments are injected into or across the coal seam these are not usually large or strong so they represent more of a threat to rom quality and local ground control as opposed to sustainability moore 1940 in a similar fashion it is occasionally the case that igneous materials can be injected into coal bearing strata forming lenticular or vertically crosscutting dikes in various locations around the world including south africa australia and the western united states these igneous dikes can be a serious obstacle to longwall operation where these bodies are narrow perhaps less than 1 0 m wide they are often sufficiently contaminated by entrained debris that they do not achieve strength beyond the limits of mechanical cutting by a shearer or continuous miner however when encountered as thicker bodies typically a stronger core requires other means to penetrate it in south africa this material is identified as dolerite and forms dikes off of sills intruded into the coal bearing strata the dolerite dike material can be more than 140 mpa compressive strength papp et al 1998 it is also the case that post d

epositional events can lead to the formation of natural coke or a natural burn can produce coal loss with only residual coal ash it is safe to assume that all of forms of rock mass nonuniformity noted will have a negative impact on longwall operation as such it is useful to identify any geologic features at the planning stage although reconnaissance drilling and detailed geologic mapping are customary various geophysical methods can be applied to search for the different discontinuity types satellite imagery magnetic or seismic surveys and electromagnetic tomography are a few of the methods used unfortunately these methods are not universally useful and choice of method deserves careful attention it is common to find that none of the methods are particularly informative for any specific site strata gases characterizing the gases and quantities that might be encountered during longwall development or subsequent longwall mining is important when striving to achieve the coal faucet concept if acceptable atmospheric conditions cannot be continuously maintained where people work and travel the longwall is at risk of being idled during periods where excursions beyond acceptable limits occur in the united states concern is emerging for atmospheric safety in the mined out areas gob or goaf behind the longwall or in sealed adjacent areas several gases have prominent impacts on longwall design and operation worldwide these include those described in the following paragraphs methane methane ch4 is common in coal mines to varying degrees and may be encountered in most sedimentary strata coal seams have long been recognized as likely reservoir rocks for methane but it is increasingly recognized that other porous and permeable near seam rocks can be significant sources of methane liberation this is particularly true when the in situ rock mass is fragmented by mining induced caving higher ch4 liberation rates in new longwall gobs result from creation of smaller particles than existed in the undisturbed rock mass release of lithostatic stress also lends to increased production rates of methane in caved material not only does the methane fill the void space of rock porosity at some reservoir pressure it can be chemically adsorbed by coals it is also common for fractures or faults to be charged with fluids occasionally the rock stress augmented with interstitial fluid gas pressure can be greater than the unconfined rock strength which results in one mechanism of outburst when rock failure occurs not only can this create atmospheric inundations from the escaping gas it can result in violent ground control hazards such events are known in a few u s and australian coal mines lastly the most notable hazard of methane in the mining environment is its explosibility when mixed with atmospheric oxygen in certain ratios as one step to avoid this hazard most longwall face equipment is designed tested and certified to be explosi

on proof or intrinsically safe explosion proof equipment will not allow an ignition within the enclosure to propagate outside it intrinsically safe components rely on the principle that methane oxygen mixtures will not ignite below certain initiation energy thresholds and they ensure that such energy is never available even during component failure conditions carbon monoxide carbon monoxide co can be adsorbed on coal and naturally occurring or it can be a product of partial combustion of hydrocarbon fuels or coal this gas primarily represents a health hazard but can be explosive in relatively high levels when mixed with atmospheric oxygen aside from its human toxicity it is often an indicator of combustion spontaneous combustion in some coals or may result from fire or explosion longwall mines with spontaneous combustion propensity may monitor the co production rate from the longwall as an early indicator of an evolving thermal event carbon dioxide carbon dioxide co2 a common strata gas may be present along with methane or may be a dominant constituent in some australian cases high pressure co2 resides in localized areas of the coal seam and requires specialized control measures to manage the hazards of outbursts co2 is also a product of full combustion of coal and hydrocarbons as such it can be one of several gases monitored to detect evolving combustion events spontaneous combustion in particular graham ratio jones trickett ratio and co co2 ratio as well as co production rate may be used to detect the onset or monitor progression of a combustion event simtars 1999 hydrogen sulfide hydrogen sulfide h2s is primarily recognized in the mining environment for its toxic effects on humans it is also a concern in longwall mining for its corrosive effects on metals and its participation in unexpected steel failures h2s can dissolve in still water and be released when the water is disturbed the aqueous mechanism can bring it in direct contact with metals such as bronze or steel in longwall components and result in metal sulfide formation and the liberation of free hydrogen h2 because h2 is not normally a strata gas but is known to be an indicator gas in combustion events this can explain its existence without the presence of a thermal event h2s can be a metabolic product of sulfate reducing bacteria these bacteria are naturally occurring and can form microscopic colonies beneath which high concentrations of h2s can exist perhaps up to 4 000 ppm this can create overwhelming corrosive or metallurgical effects while atmospheric levels remain low other gases and vapors on occasion other gases and vapors may naturally occur low molecular weight hydrocarbons can exist in gas or liquid phases in sedimentary strata their combustibility and quantity may be reason for concern in some instances in most cases the regulatory body that has jurisdiction over mining health and safety for a specific location will have

statutory guidelines for safe and legal levels of the noted gases in order to be successful and sustainable a longwall system must be designed to accommodate safety and statutory compliance in order to avoid production outages associated with gas level excursions above allowable limits water in many areas water is associated with the coal seam or nearseam strata it may also be associated with overlying or updip workings and fracturing or faults in any event its management is reason for careful planning because of its generally negative impacts to longwall development and operation as well as the typically negative impact it can have on ground control where swelling clay minerals are present in the rock the matrix may decompose when the clay minerals hydrate and swell the slake durability test is one measure of a rock s behavior when exposed to moisture when swelling clay minerals are present the rock typically performs poorly in slake durability siltstone mudstone claystone and fireclays are often the operating floors for longwall development sections or the longwalls themselves the often present water coupled with the duty as a roadway for rubber tired equipment turns these floors to mud sometimes hundreds of millimeters deep in the immediate headgate or tailgate entries this can mean that the belt tailpiece and stageloader are resting on mud and may even invite floor heave which requires clearance before equipment advance it also can mean that gate shields may unreliably advance or be unable to create the intended load into the roof leading to a deterioration of roof control in the headgate or tailgate this same principle can leave standing support such as props cans or cribs with an inadequate foundation upon which to create resistance to roof deformation water management or the failure to manage it is often a reason that longwall productivity falls short of its full potential spontaneous combustion spontaneous combustion is a condition where slow oxidation of a combustible substance leads to heat evolution and increasing temperature until the material reaches a point where combustion starts and spreads to adjacent fuel some coals particularly those with lower rank exhibit a propensity toward spontaneous combustion cliff et al 2004 it is a particular concern when spontaneous combustion heatings or thermal events take place in a longwall gob goaf where they can be difficult to locate and identify and even more difficult to access and extinguish such events typically cease production in the affected portion of the mine until in the best case the event ends or in the worst case causes loss of life and assets as serious as this scenario is an even worse one is possible where methane is present in combination with a coal prone to spontaneous combustion in that case atmospheric oxygen in combination with explosive methane levels can be ignited by a small spontaneous combustion ignition source re

sulting in an explosion and fire in such an event multiple fatalities and loss of assets are possible because responsible mine operators are committed to a zero harm philosophy effective controls must be offered to preclude the possibility of a spontaneous combustion event or an explosion and fire the mitigating strategy for all of these concerns is the long known fire triangle concept for fuel oxygen and ignition source considering the fire triangle it becomes immediately evident that in a spontaneous combustion prone coal mine the only component that may be subject to control is oxygen thus the choice of bleederless longwall ventilation explained later can effectively exclude oxygen from the gob goaf and mitigate the hazard as with any underground fire the issue is only partially getting the fire extinguished which usually entails oxygen deprivation but also dissipating the residual heat resulting from combustion so that the event will not simply resume if oxygen becomes available again this is also why early detection and intervention are key components to a spontaneous combustion control plan but the most important element is a strong focus on prevention mitchell 1996 roof and floor strata one of the advantages of a longwall mining system is that it can accommodate a wide range of roof conditions gate road stability issues aside stabilizing the longwall roof is actively done by the shields the shield must immediately stabilize the near seam roof when it is initially set against the roof after it is advanced referred to as set load the set load required to stabilize the roof is a function of site specific conditions and parameters of the face cross section the set load produced by the shield is derived from the emulsion system pressure available to the shield during setting the major stage diameter of the leg and its resulting area set pressure applied to leg area and derated to account for leg inclination angle to vertical functionally defines set load created by the shield to be successful set load required to stabilize the roof must be less than the vertical thrust created by the shield at setting subsequently the shield must stabilize the roof throughout the time when the shearer cuts past it adding a web depth to the span and ultimately as adjacent shields are lowered away from the roof as they are advanced loads also arise from face deterioration and span enlargement to the degree that it occurs in response to development of front abutment loads from web to web if the shield resists this load and the roof remains stable the result is the end cycle load required to support the ground in concept this should be at or below the yield capacity of the shield this yield load is calculated in a manner similar to that of set load except that instead of system setting pressure the opening pressure for the leg yield valve must be considered with set load and yield load independently defined it is

possible to create some of the popularly discussed ratios such as set to yield ratio yield capacity linear length of face and set or yield capacity area of supported roof although these terms are frequently discussed as design factors they are in fact only outcome statistics to the actual equilibrium that must be successfully established they also have limited comparability over a range of possible operating heights because only the vertical component of the leg load primarily contributes to roof stability the horizontal force component is mostly resolved within the shield by the lemniscates links because the interaction of the roof and the roof support shield has been described and the need for stability is clear the factors leading to loads arising from the roof must be characterized roof loads on shields increase with increasing distance along the face from either gate road up to a maximum load specific to any site distances greater than that effectively only expand the number of mid face shields exposed to that maximum load it is not prudent or practical to attempt to select shields in the design process for any load less than the maximum load experienced for the specific installation certainly using outcome statistics or rules of thumb to size shield capacity is poor practice and risks the possibility that an entire longwall will fail to perform as intended should this be the case available solutions may require premature equipment replacement or acceptance of undesirable operating performance evaluating the factors that contribute to roof loading on the shields it is clear that both site specific and face crosssection factors are contributory site specific factors include seam thickness extraction height near seam strata type and thickness seam and nearby strata material properties caving behavior of strata seam inclination natural and mining induced fractures and previous mining above or below the seam in concept the overlying rock collapses into the void created by mining as the intact material fragments and caves with particles rotating and translating into the void its volume bulks to fill the void largely in inverse proportion to particle size this process continues until little enough void space remains between the rubble and the uncaved rock that further displacement of the overlying rock takes place through particle translation but with little rotation eventually even the downward translation reaches equilibrium with the resistance created by the underlying compacted rubble thus the weight of rock overlying the longwall can be supported across three areas 1 unmined face ahead of the longwall front abutment load 2 longwall roof supports 3 rubble material of the gob the area depicted in figure 13 8 4 graphically shows the region of rock that must be stabilized by the roof supports shields in consideration of the loading process described it is evident that the loadi

ng of longwall roof supports has little impact from mining depth it also becomes evident that overlying roof materials that cave as small particles from thin layers can lead to smaller caving heights before particle rotation is diminished and consequently to lower shield loads and thus lower shield capacity requirements conversely thicker and stronger overlying beds can impose much larger loads on roof supports especially if they are close to the mining horizon another consideration is that stiff roof rocks typically strong as well can require high set loads to stabilize them further they can lead to extreme cases of periodic weighting where strong beds can act as cantilevered loads imposing successively greater loads on the roof supports until the cantilever fails such strong beds can be multiple within the caving horizon and cause superimposed effects of periodic loading figure 13 8 5 shows an example of how this variation might look for an example longwall application barczak and garson 1986 wind blasts originating in the gob often coincide with the caving of thick strong beds or the first cave of a new longwall panel in some instances a cantilevered bed failure results in the release of stored energy and in a bump or bounce this mining induced seismicity can be damaging to equipment and hazardous to personnel in general mining induced seismicity is often associated with strong rock or coal and significant depth high overburden stress coupled with mining layouts that may act as stress concentrators can lead to mining induced seismicity originating at elevations ranging from beneath the mining horizon and extending upward to the surface highflow yield valves on shield leg cylinders can afford protection against equipment damage at moderate convergence rates in a periodic weighting event but only purpose designed rockburst protection can prevent damage to shields during a bump where a large convergence can occur suddenly face cross section factors that affect loads imposed on the shields include geometry of the shields shearer and afc depth of cut web and sequence state in cutting cycle because of these factors coupled with geologic variation that takes place on scales smaller than the area mined by a set of roof supports within their service life or even within the course of a panel it is not surprising that the loads imposed on shields vary over a range at any installation figure 13 8 6 shows the cumulative probability distribution of loads required to stabilize the longwall roof these curves can be numerically modeled for a proposed longwall face cross section and strata section and are site specific for every individual application models of this type have shown good correlation with data collected from pressure monitoring on equipment in service as discussed previously the set load curve is derived independently of the yield load curve it can be observed that in the region above approximat

ely 96 cumulative probability cp providing additional support load into the roof offers diminishing returns using this method to select roof support capacity leads to selection of supports that control loads from the roof in the 91 96 cp range this leaves a 4 9 incidence of yielding in most cases yielding events will be of brief duration and equilibrium will be reached by sharing load onto adjacent shields selection of capacity in the higher end of the range 93 96 cp allows for some deterioration in performance as equipment ages without creating performance problems choice of supports larger in capacity than 96 cp of the roof loads that will arise offers higher cost greater weight reduced travelway dimensions and elevated emulsion system demands with little perceptible benefit however because the probability distribution is nonlinear selecting capacity much below 91 cp leads to an excessive occurrence of shield yielding and onset of roof damage companion to the more severe yielding events notably potential shields for any application should be evaluated at different heights leg cylinder inclinations to verify that they will be adequate for any mining height condition they may experience in service excess debris should not be allowed to accumulate on top of the roof canopy or beneath the base as this material can be compacted and may allow excess roof convergence and damage to arise before enough support resistance is generated to create equilibrium this is especially true for strong stiff roof materials such as competent sandstones near the roof line in some coal mines specific energy of cutting the productivity of a mechanized cutting process for rock is largely a function of the material s specific energy of cutting which depends on the type of cutting being applied shearers are different from drum type continuous miners which are yet different from longwall plows while shearers and continuous miners work against forces related to compressive strength of the material being cut plows primarily work against the tensile strength of the material which is typically much weaker than the compressive strength particularly with respect to shearers the cutter motor power used to produce a resultant quantity of rom ore over a period of time defines the specific energy of cutting for the material kilowatt hour per metric ton kw h t this assumes that the cut material can be cleared so that product accumulation does not impede the process this measure is useful for the purpose of production process design and management but is not as pure a parameter as the name might suggest the load applied to the material to be cut can alter this value and is a benefit for the longwall system when the frontabutment load has fully evolved after the start of mining in a panel this load is usually large enough to soften the face and reduce the energy required to produce a quantity of material until the front abutment

fully develops cutting will be more difficult than after it exists the apparent softening can also be noted as abutment load varies in periodic weighting cycles and when production resumes after a stoppage of sufficient duration to allow significant stress redistribution the addition of work done by the abutment stress is also advantageous to creation of a larger product size distribution and lower fugitive dust production on a milligram cubic meter metric ton basis interestingly the cutting bits deeper in the web do more work and produce a smaller size distribution of product than the bits shallower in the web this smaller sized product may be significant in dust production which can become airborne the possibility of improved machine utilization and proportionately reduced dust production exists with adoption of increased web depth in the transition of web depths from 0 76 m to as much as 1 07 m this appears to have been true specific energy of cutting for various materials being cut by shearing machines ranges from 0 05 kw h t for easily cut coals to 1 3 kw h t or more for some very hard cutting sandstone and igneous dike materials where very difficult cutting is expected for an extended duration specially designed rock drums are used in lieu of normal production drums these rock drums are designed with closer bit spacing and more bits per line in the drum lacing to accommodate the heavier cutting demands the style of bits may also change between radial and conical cutting bits to offer improved longevity and performance typically high capacity longwalls operate well when they encounter materials in the 0 05 0 45 kw h t range abrasivity when considering a longwall mining application the abrasivity of the product to be mined and conveyed is of concern most wear on the afc and beam stage loader bsl is related to the tonnage of material that has been conveyed across the systems as an approximation to the wear that might be expected in a particular application standardized laboratory tests have been developed that are also used by the electrical generation industry which has similar concerns as mining about abrasive wear to equipment within its facilities in high level summary sacrificial blades are rotated through a mass of the rock material of interest for a predetermined time at the end of the test the weight of material lost by the sacrificial blades is determined and reported this weight loss ranges from only 8 mg fe in a trona specimen to more than 1 300 mg fe in a severely abrasive coal and sandstone rom product typical values range from 200 to 900 mg fe which roughly correlates with wear of 0 5 to 2 5 mm of deck plate thickness reduction per million metric tons of production unacknowledged in these values are the corrosion processes which can certainly influence the outcome in service improved results can be achieved by use of abrasion resistant steels where the corrosion environment will not provoke pr

emature failures ludema 1996 mine layouts having identified most of the important site specific considerations that might have to be addressed by a successful mine design the general mine layouts can begin to be discussed historically advancing and retreating longwall layouts have been the subject of much discussion worldwide in the advancing system the longwall setup room is close to the main or submain entry set and the longwall mines away from these entries into undeveloped reserves in the process support is constructed to constitute one or more gate roads behind the longwall face mining continues until a designated extent is reached or conditions force face recovery the fact that artificial roadways have to be maintained in the gob adds significant cost and labor to this technique as cost and productivity benchmarks have evolved across the industry the elevated costs and lower productivity of advancing faces have generally seen these operations cease or transition to retreating longwall layouts only in unusual circumstances or where historic precedent dictates are advancing faces operated at present all of the world class longwall installations are retreating faces this trend can be expected to continue into the future retreating longwall layouts may be of two types conventional or punch highwall type figure 13 8 7 shows a conventional retreating longwall layout with a mined out panel adjacent to the active longwall the entries of the previous panel s headgate main gate transition into becoming the tailgate for the active face the active longwall face mines progressively from the setup room at the far inby end of the developed panel and is eventually recovered as it enters its intended recovery position the recovery position of the face is shown by the line that coincides with the near outby extent of the mined out panel optional recovery chutes may be used to aid free equipment recovery the punch highwall longwall is largely similar except that instead of developing a main or submain entry set from which gate roads are driven the gate roads are developed directly off a surface mined box cut or highwall punch longwalling is attractive because it eliminates the need to develop and maintain long term underground workings in the form of mains and submains entries it also lends itself well to the transition from surface mining to underground mining that confronts many mine operators as the depth and production costs of remaining reserves increase gate roads for retreating longwalls are typically formed with two to four entry developments although single entry and dual entry gate roads are common outside the united states dual entry gate roads are an exception in the united states in the interest of worker safety in a fire or explosion emergency u s mining law generally requires a minimum of three entry developments in coal mines to accommodate separated intake fresh air return exhaust and conv

eyor belt entries in cases where it is safer to have a dual entry development than the statutory minimum of three entries a variance may be granted by regulatory enforcement authorities based on a demonstrated diminution of safety argument u s mine operators have succeeded in the diminution of safety argument where adverse geotechnical conditions threaten worse consequences than the reduction of number of entries companion to the grant of such a variance are additional controls to safeguard against the hazards envisioned during statute development often especially in the united states three or even four entry developments are used for gate roads the motivation for creation of more development length in total than is minimally required to sustain longwall operation should be carefully considered it is ordinarily desirable to minimize the ratio of development costs per longwall metric ton produced it is also desirable for time value of money considerations to associate as closely in time as possible the expenditure of costs that provide future benefits with realization of those benefits in the pursuit of these principles it is common for longwalls to be operated in geometric succession with the headgate of the prior panel being used as the tailgate of the current panel and so forth until some predefined extent is reached often this extent is the aggregation of a group of panels to form a district the number of panels in a district is variable typically based on ventilation or geotechnical considerations when managing the constraints of a large district eventually becomes difficult enough mine operators prefer the development of the incrementally additional gate road which is required to develop the first panel in a new retreating longwall district frequently districts are sealed upon completion removing them permanently from the worked and traveled portion of the mine mine operators are then largely relieved of the burden to maintain access for inspection and ventilation of such a sealed district but access must be maintained to inspect the final seals where the district meets with active mine workings and the seals themselves must be maintained in adequate condition accumulations of water explosive atmospheres or spontaneous combustion events can be reasons for concern even in sealed and otherwise abandoned areas relatively recent changes to u s regulatory requirements for seals and monitoring atmospheres in sealed areas address these concerns on occasion it can be necessary to isolate overburden stress effects between the current longwall panel and the prior panel with a barrier pillar where panels are mined at depth with thick stiff beds in the overlying strata failure to adequately support the stiff strata can channel vertical stresses toward the current panel this elevated stress can result in mining induced seismic activity in the strata section with possible events in excess of 3 5 on the richter sc

ale these events can be damaging to equipment and hazardous to personnel particularly when seismic activity originates at or near the face location at a lesser scale stiff coal subjected to large stresses can fail suddenly and result in the expulsion of material from the face toward the longwall travelway creating a hazard to personnel longwall gate road development is characteristically just in time to avoid delaying production activity on the longwall the degree to which the development is ahead of the needs of the longwall in time is often referred to as float time some ventilation plans require the next adjacent longwall to be developed in order to implement the ventilation plan more commonly this is not the case and float times are in the 6 to 50 week range this allows adequate time to withdraw the development equipment and stage longwall move supplies and equipment prior to the onset of a longwall move however it does not impose unnecessary hang time on the gate road as deterioration of conditions can begin soon after development in some settings it also associates the cost of such development in time closer to the future benefits expected to arise from it for many mine operators the longwall move actually starts with emplacement of portions of the new face prior to completion of mining in the current panel such equipment might include tailgate gate shields afc and bsl a new or rebuilt shearer electric power and control equipment and emulsion pumping equipment controls and tank the amount of equipment preinstalled depends on arrangements specific to any mine operator in general shorter production outages occur with increased amounts of equipment preinstalled in the new longwall panel gate road development itself is done with continuous miners and one of the several face haulage mechanisms available figure 13 8 8 shows the general layout of a three entry gateroad development section with a continuous miner and battery coal haulers the section depicted uses a place changing method of operation where a continuous miner mines a cut and then changes places to mine a cut in another entry or crosscut when the continuous miner has withdrawn a separate machine installs roof bolts and possibly wire mesh or roof straps to roof and rib or roof trusses if necessary roof bolting progresses from where support was installed on the prior sequential cut in that working place to the face of the cut after a cut is fully supported the roof bolter is withdrawn and the cut is ready for auxiliary activities that may be necessary before the continuous miner returns to make its next cut in that working place as this process is successively applied to entries and crosscuts the entry set is advanced cuts may range from as little as 3 0 m to more than 12 0 m in good roof conditions the cut depth may be limited to the depth attainable without exposing personnel to unsupported roof or airborne contaminants short cuts have

a decidedly negative impact on the productivity of place changing mining methods as a disproportionately great amount of working time is lost to place changing as depicted ventilation is provided by section fans the coal haulers transport the mined product to a conveyor belt for ultimate clearance from the working section it is common for the conveyor belt to be equipped with a feeder breaker the feeder breaker has a limited amount of surge capacity allowing coal haulers to unload quickly but also meters product through a breaker and onto the conveyor belt at rates that optimize belt system performance this is particularly the case where multiple sections all discharge onto a shared conveyor belt place changing operations are typically applied in relatively good roof conditions and often have better productivity than the alternative in place continuous miners these in place continuous miners also called bolter miners install roof bolts mesh straps and even roof trusses in parallel with the mining cycle thus roof support is installed close to the face and the roof has minimal opportunity to deteriorate before additional support is implemented to augment its natural integrity also this method minimizes the number of times that the bolter miner is moved between working places thereby minimizing damage to potentially weak floors historically seriously flawed weight distribution on bolter miners led to worse floor damage than place changing machines even though machine travel was less recognition of this problem has allowed it to be mitigated on modern machines aside from mining and bolting section operations are much the same between in place and place changing systems some mine operators use loading machines between bolter miners and the coal haulers the loading machine allows the bolter miner to discharge product into a small surge pile on the floor behind the bolter miner the machine loads from this surge pile into coal haulers as they become available and decouples the continuous miner from producing and loading only when a coal hauler is on station and ready to receive product the use of a loading machine also caters to an often noted deficit of bolter miners which is cleanup in cut as a general principle bolter miner productivity is less than place changing operations when a reasonably long cut perhaps greater than 6 0 to 7 5 m is available however if cut depth falls below this range or if ground control is unacceptable in the cut depth available a bolter miner system may become the preferred option current bolter miners have the ability to sump and bolt simultaneously this has a notably favorable impact on productivity when compared to earlier style bolter miners which could only cut as much as was available with the machine frame stationary when drilling and bolting activities were ongoing additional concerns for gate road design are entry and crosscut span crosscut angle and spacing gate road e

ntries must serve several purposes including the following provide adequate cross section for ventilation provide access of personnel and materials transport and installation of longwall equipment installation of the longwall section conveyor belt temporary locations for emulsion pumping equipment and electrical power and controls retreat of the longwall in normal operation standing roof support is restricted to only those areas where it does not obstruct necessary machine travel in the longwall conveyor belt entry supplemental support may be installed as a precaution because there is little opportunity to add remedial support after the belt is commissioned four distinct regimes of increasing loading are applied to gate roads aside from the first and last panels of a district 1 initial development 2 adjacent to a single longwall gob headgate 3 intermediate to adjacent longwall gobs active tailgate 4 longwall gobs on both sides well behind inby the face the primary roof support system controls the ground in the first and second stages but supplemental support is commonly applied prior to use as a tailgate for the active longwall this support is typically installed in the immediate tailgate entry and crosscuts before it reaches third stage loading and where practical it may even be installed at the time of development commonly used materials include wooden cribs 4 point or 9 point cans steel props and pumpable cribs the alternative of right angle intersection of crosscuts to main entries versus inclined is a balance of considerations between the convenience of access afforded by inclined intersections and the propensity of acute angled corners to deteriorate with consequent enlargement of the intersection sum of diagonals sod increasing sod often correlates with decreased stability of intersections floor stability and horizontal stress direction can influence the optimal intersection design the result of entry spacing and crosscut spacing is pillar size the correct selection of pillar size for longwall gate roads is an involved discussion beyond the scope of this chapter in general bigger pillars are more desirable from a stability perspective but the unrecoverable reserves lost to gate road pillars the increased development footage and prolonged development time is not desirable long pillars are favorable to gate road development but wider pillars are not this is because development along panel length is useful to the purpose of completing panel development but distance spent on crosscuts is not access to the longwall face for heavy maintenance is enhanced by shorter crosscut spacing face length and panel length the choice of longwall face length is a balance of the initial capital cost committed to acquire longwall equipment versus the benefits of improved productivity production cost savings and additional reserve recovery for a wider longwall face the decision is also limited by the

capabilities of existing equipment although capabilities have steadily improved over time in subsidence sensitive areas wider faces can cause mine plans to be offset further from sensitive surface structures and sterilize additional reserves in the process a cost weighed against recognized benefits as the industry wide dedication to zero harm redefines acceptable hazards to ever diminishing severity and frequency a conflicting prospect emerges as mining conditions tend to become increasingly adverse and present increased hazards the cumulative result is an increased amount and cost of development footage required to sustain longwall operations consequently increased longwall face length is attractive for most site specific circumstances an optimum longwall face length can be defined often there is a broad optimum in net present value terms between 300 m of installed equipment length and up to perhaps 475 m in other cases present technology allows face length up to approximately 500 m with high productivity common face lengths range from 250 to 425 m achieving maximum lengths implies the use of the largest afc chains and highest rated electric motors and afc gearboxes available longwall panel length is an issue constrained by ventilation system limitations timing considerations to avoid production outages longwall equipment longevity property constraints and reserve extent at some point excessive ventilation pressure losses and insufficient air quantities dictate a practical limit to the length of a gate road development or may constrain longwall operation alternatively the time required to complete a long development may compare unfavorably with the rate at which an active longwall is retreating yet another consideration is whether the longwall equipment can retain the required availability levels as it wears through continued service although durability of the constituent equipment elements of the longwall has been improving through time it is not unreasonable to expect shields afc drives and frames to serve through the longest panels afc sprockets can be reversed or replaced during planned maintenance intervals as can chains flights and chain connectors shearing machines have longevity on par with the longest panels highest tonnages but can also be repaired as subcomponents or a replacement shearing machine can be implemented during mining of a panel more readily than undertaking a complete longwall move panel transfer with this contemporary longwall panels range from 1 850 to 4 600 m in length with contained panel tonnages up to 11 million metric tons of coal thin seams most of the discussion so far has been focused on the most common high productivity longwall variation single lift longwall in the 1 75 to 3 75 m extraction height range however thin ore deposits typically with higher value such as coking coals are of interest thin seams are also a topic where they overlay other coal

seams but must be extracted first to prevent their sterilization by undermining historically thin coal seams were extracted with low productivity shearing machines and longwall plow systems low seam shearing machines of the past are hopelessly obsolete compared to current expectations periodic attempts are made to make low seam shearers practical as a number of mining projects worldwide reside at the edge of viability and would be successful if only better productivity were available shearers in low seams are constrained by factors such as motor size diameter and power vane depth of drums and loading performance clearance shearer tunnel and ranging arm length and height limits traversing undulations and operator ability to travel at tram speed of the shearer while efforts continue to solve the low seam shearer issues progress has been made in the modernization of longwall plow systems figure 13 8 9 shows a longwall plow instead of cutting the coal with a shearer the coal is cut by a plow body captivated on the afc and drawn from gate to gate on a dedicated chain driven by motors on each face end since the plow cuts only a fraction of the web of a shearing machine normally 100 to 140 mm but up to 250 mm on the highest capacity installations multiple transits of the face are required to achieve equivalent production to one pass with a shearer in contrast the plow speed may be more than 10 times that of typical lowseam shearers so plows offer attractive potential in deposits too low for shearers provided that conditions are favorable gluckauf gmbh 1985 favorable conditions for a plow type longwall include strong continuous roof and floor rock ore that plows readily and parts cleanly from roof and floor ore horizon with little thickness variation and limited undulations and ore free of intrusions faults or dikes roof falls in the tailgate or on the face can be serious obstacles to a plow replacement of a significant section of the extracted horizon with sandstone through faulting or washout can be equally problematic the history of plow type longwalls is extensive in europe where almost every conceivable condition has been confronted by plowing but not always to successful outcomes in modern terms the seam conditions that make plow application attractive in a modern context are limited compared to the totality of potential low seam reserves worldwide at present only a few plow type longwalls compete with shearer based longwalls for the same markets thick seams unlike thin seams thick seam longwall has great potential worldwide single pass longwall designs up to 7 25 m high have been proposed with installations more than 6 5 m high in service a variety of issues require adaptation of the designs of these high installations including shearer steering into out of the face structural integrity of afc to bear heavy shearers face operator travelways and safety slab development

or persistence face spalling ahead of the shield tips shield capacity and resulting weight and potential worst case afc starting capability the production history of many of these installations bears testimony to the success of designers and mine operators in overcoming the numerous challenges arising from thick seam operation at any period of time some of these high faces appear in the ranks of the world s most productive longwalls interest is intensifying in longwall extraction of even thicker seams than 7 25 m but there is an accompanying recognition that single pass longwall is not the solution to extract very thick coal seams the longwall top coal caving ltcc technique is being developed two major variations of the ltcc method exist in one the face cross section is much like a conventional shearer based system with a single afc an original pass is made as customary at a convenient height and then a special purpose caving chute integrated into the shield canopy is lowered so that material caving from above the shield is channeled into the afc when product quality or caving conditions dictate the chute is retracted and successive chutes are opened along the face the process is repeated until the majority of the available product has been drawn at this point additional undercutting is conducted with the shearer in preparation for successive drawing of caved material alternatively another technique of ltcc uses an afc both ahead of the shields as expected but also behind them under a powered flipper used to allow or preclude caved coal introduction onto the rear afc figure 13 8 10 shows a typical installation in cross section potential exists for more new ltcc installations worldwide than the total number of modern high productivity longwalls currently in service in north america today ltcc installations exist in china and eastern europe with considerable expansion of the technique possible there as well as significant prospects for introduction to australia and possibly even the americas the ltcc technique could require adaptation to conform to statutory requirements in countries where the technique has not previously been practiced multiseam longwall on occasion thick seams are extracted by single lift longwall operations conducted in a coordinated sequence with the upper longwall progressing ahead of the lower longwall in an interlocked sequence though this method has examples in the field coordinated operation of independent longwalls can be difficult thin interburden is more problematic than thicker interburden as are joints or fractures that transect the interval between the seams thin or low shear strength interburden as well as jointed interburden can transmit stress between mining horizons more directly than horizons separated by even a modest thickness of strong stiff interburden high stresses are often transmitted from one horizon to another where strong structures surrounded by high e

xtraction workings in one horizon are approached by workings in the other horizon in general multiseam operations are more focused on extraction of independent seams at appreciably separated intervals of interburden and time often the earlier workings are not executed with anticipation of modern longwall mining in the future thus making mine planning an exercise in mitigating interseam interactions longwall ventilation ventilation of longwall mining systems seems to exhibit nearly endless variety in the exact details of worldwide plans ausimm 1986 kennedy 1999 taken at a broader level longwall ventilation falls into two general categories with very different mechanisms msha 1996 to introduce these principles some basic attributes of methane and air mixtures should be clearly understood figure 13 8 11 shows coward s triangle coward and jones 1952 other representations of the explosion hazard are popular in various locations around the world but the chemistry and physics of the matter are universal thus other representations describe the same physical behavior two principal regions are represented in figure 13 8 11 the upper portion of the figure describes compositions that cannot be formed from methane and air and is therefore irrelevant below these naturally impossible compositions of methane and air is the region of naturally possible mixtures these possible mixtures include two major subdivisions inert mixtures and explosive compositions the inert mixtures will not explode if exposed to a potential ignition source and occur as fuel lean or fuel rich compositions with respect to the explosible compositions the explosible compositions exist within coward s triangle shown in the approximately 5 15 methane range and approximately 12 21 oxygen range the zone depicted around the triangle is an arbitrary safety margin this physical behavior of methane and air mixtures being an absolute principal two approaches to avoid the explosion hazard exist one is to provide fresh air to dilute methane to a fuel lean state if such a state can be maintained without fail it renders an acceptably safe condition paramount to this strategy is that the fuel lean oxygen rich atmosphere cannot transition from a respirable state required where people work and travel to a fuel rich oxygen deficient state which is likely to occur in the unventilated gob where the respirable mine atmosphere transitions to an unventilated gob atmosphere particularly after mine workings are sealed it becomes impossible to guarantee that the reducing oxygen level and increasing methane level will not cross through the region of explosible compositions without active intervention this case is described by the upper trajectory line in figure 13 8 11 if even a small energy source above the intrinsic safety threshold is available to ignite the explosive composition a violent event will result in proportion to the volume of gas mixture ignited sim

tars 2001 if coal dust becomes airborne secondary to a typical methane explosion a much more energetic and violent event can arise burrows 1989 this mechanism has repeated itself many times worldwide often with disastrous outcomes application of incombustible dust rock dust or stone dust to areas where coal dust can accumulate is effective in mitigating this dust explosibility hazard an alternative mitigation strategy for the explosion hazard involves the deliberate exclusion of oxygen from gob atmospheres followed with forced inertization if necessary to depress oxygen levels below those where an explosion hazard can exist regardless of methane concentration this is represented by the lower trajectory line in figure 13 8 11 which is actively controlled by inert gas injection and never crosses the region of explosion hazard coward s triangle periodic atmospheric monitoring followed by application of the inertization process as necessary can prevent explosible methane oxygen compositions from being created by already inert atmospheres where introduction of new oxygen is possible with this in mind the two major strategies of longwall ventilation are bleeder and bleederless ventilation systems a conceptual flow through bleeder longwall ventilation system is shown in figure 13 8 12 in this system fresh intake air is supplied through the headgate entries and splits at the headgate to ventilate the longwall face an adequate volume is allowed to flow behind inby the face at the headgate to ventilate the perimeter of the gob along the face and at the tailgate air is allowed to bleed into the gob and tailgate to ventilate the methane to fuel lean levels this is the most common ventilation system used in the united states it is notable that because atmospheric compositions are uncertain where they cannot be monitored and controlled in the bleeder ventilated gob inby the longwall and in the adjacent abandoned gobs any potentially persistent ignition source such as spontaneous combustion or intermittent source of ignition energy such as lightning is reason for concern the major alternative to bleeder ventilation of longwalls is bleederless ventilation the bleederless longwall ventilation system is depicted in figure 13 8 13 the notable differences between the bleederless and bleeder ventilation systems are that in the bleederless system seals are built in headgate crosscuts immediately inby the face in order to progressively seal the gob and exclude oxygen as the longwall retreats if concerns exist about atmospheric compositions in the gob or if oxygen invasion could stimulate spontaneous combustion the added feature of forced inertization as discussed previously can be an effective riskmitigation strategy gases such as nitrogen co2 or even oxygen depleted air can be used for safely inertizing mine atmospheres the exhaust gas from combustion processes such as gas turbine engines can potentially be a source of

oxygen depleted gas mixtures for inertization to prevent oxygen inflow into sensitive gobs pressure balancing is often applied to minimize pressure differentials across seal lines and minimize introduction of new oxygen to a gob an alternative technique is the use of balance chambers as depicted in figure 13 8 14 in this method an initial seal is built to separate the atmospheres and then another seal is built outby the initial seal becoming primary inertization gas delivery pipes penetrate both seals to allow inert gas delivery to the sealed area and the space between the seals if the seals have reasonable integrity the space between the seals can be elevated to a low positive pressure and the small leakage volume that occurs is either into the gob the mine airway or both the net effect is to prevent oxygen ingress into the sealed area at low total inert gas expenditure sudden change in inertization gas consumption can be an early indication of deterioration in seal performance a valuable tool to monitor the composition of mine atmospheres in critical areas is a tube bundle system with this equipment atmospheric samples are drawn from critical locations underground through dedicated individual tubes to analysis equipment located at the mine surface this allows sampling of areas where electrical sensors might be prohibited and sampling continuation in the event that electrical power is removed as is the case in many mine emergencies an addition to the controls inherent in either of the ventilation systems described gob ventilation boreholes gvbs can be used to assist the ventilation system by extracting gob atmosphere predominantly methane before it enters the ventilation system gvbs are often centrally located in a longwall panel when the intention is to intercept methane evolved from broken strata before it enters a bleeder ventilation system these holes drilled from the surface in advance of mining are often spaced at larger intervals than gvbs intended to control expansion volumes in a bleederless ventilation system the location of gvbs is typically along the periphery of a panel where void space is connected to workings at seam level when the intention is to control barometric expansion volumes as in bleederless ventilation systems diurnal barometric pressure variation occurs daily with the barometer falling as the atmosphere warms with duration and intensity of solar exposure and rises as night falls when the atmosphere cools and air density increases random barometric pressure variations are related to weather systems the actual barometric pressure is a function of both effects and can be predicted for a surface location with good overall accuracy operation of gvbs in response to barometric pressure variation alternating with inert gas injection can control evolved methane and prevent oxygen invasion into progressively sealed gobs as in bleederless systems or other sealed areas although gvbs are

typically drilled vertically from the surface an alternative method applies medium radius drilling technology to postmining methane drainage the wells start with a vertical collar followed by a build section where the well is steered to a horizontal position the horizontal lateral is located in a stable horizon above the longwall panel holes up to 1 6 km in length producing approximately 1 m3 s of methane have been achieved an alternative to gvbs drilled from the surface is an array of cross measure boreholes drilled at upward inclination from the seam horizon such cross measure holes can drain methane prior to mining and be used later to a similar effect as gvbs when caving behind the retreating longwall opens a communication between the holes and the larger gob gas captured by the holes can be collected by an underground piping system and directed to a borehole to the surface without entering the mine ventilation system horizontal boreholes originating in seam or turning into the seam from vertical wells to the surface can be progressed in seam for more than 1 km to drain methane in advance of mining these holes have proven to be highly effective for methane drainage but must be planned carefully as they can later present hazards that must be mitigated before mining progresses through them contemporary issues improving the health and safety of miners by reducing their exposure to hazards has been a major motivation in the steady progress of mining mechanization longwall mining as it exists today is the product of this ceaseless effort greater productivity and resource recovery are realized while the incidence of injuries and accidents continues to decline as a trend few would dispute that longwall mining is by far the safest and most efficient underground soft rock mining method although many concerns have been addressed with great success others have persisted coal workers pneumoconiosis more commonly called black lung is a chronic disease which investigators have correlated to respirable dust exposure in the coal industry black lung is often debilitating in its advanced phase and may lead to death most coal producing countries worldwide acknowledge this hazard and have occupational health limits for respirable coal dust exposure intended to prevent workers from contracting the disease not all dust is deemed hazardous but only the submicrometer 10 m size particles fraction of dust is believed to contribute to black lung because longwall mining results in large production it is not surprising that increased amounts of dust can also result with respirable dust being a fraction of the total depending on the regulatory strategy applied this can lead to different conclusions one strategy is based on environmental exposure this assumes that a worker will not benefit from any personal protective equipment and that safety is produced only by maintaining environmental levels below disease thresholds this stra

tegy does not give innovation or technology an opportunity to contribute solutions unless it impacts the environmental exposure as opposed to the respiratory exposure however personal protective equipment such as airstream helmets offers demonstrated effectiveness and automation technology has the potential to remove workers from most dust exposure personal dust exposure management is now more feasible than ever with the demonstration and introduction of personal dust monitors seemingly the optimum solution to the problem of dust exposure would apply the best available technologies crediting the benefits in realistic terms to create incentive for continued progress in longwall technology another issue of modern underground mining is related to surface subsidence the most damaging aspect of mininginduced subsidence is the zone that connects undisturbed surface with areas subjected to near maximum vertical displacement this boundary varies from a few meters to a few hundred meters wide and commonly wraps around the perimeter of the subsided area the condition is exaggerated when subsidence takes place years after active mining and may continue for a protracted period of time it is also worsened when small land areas are subsided a few hectares per instance leading to a greater proportion of subsided land area being subjected to the most damaging end outcome kratzsch 1983 saimm 1982 these situations often acting in combination make compensation for damage to surface assets difficult because the mine operators may no longer have a corporate presence or the ability to pay for reparations years or even decades after the fact peng 1992 the development of longwall mining has brought relief to these concerns as modern longwall panels have gotten wider and longer the proportion of subsided land area permanently subjected to the most damaging boundary conditions has decreased also unlike the subsidence of smaller individual areas the subsidence develops to its maximum level in close alignment with predictions and generally does so in a period of weeks to months after active mining by longwall methods thus all stakeholders are in a better position both in terms of damage and timing to make such reparations as may be appropriate longwall mining has been extensively applied to urban and suburban areas in germany poland and the united kingdom to mine under a tremendous variety of structures including public buildings and residences industrial facilities roads bridges railroads pipelines and power lines to successful outcomes longwall mining has also been used worldwide to mine under rivers lakes and oceans successfully with the worldwide knowledge base for mitigation of subsidence damage to surface structures many alternatives can often be applied to minimize damage to surface assets or repair them effectively as longwall mining transitions from rural to suburban areas in the united states and australia it ha

s vocal critics the arguments put forward derive from many origins but generally do not suggest a better method to extract the resource but rather oppose the extraction all together or dispute the value of the damages that might be created or the ability to bring about acceptable reparations conclusion when longwall systems are properly selected and operated the results can be spectacular and sustainable these installations make longwall mining look easy in reality successful selection and operation of longwall systems are based on attention to detail and a commitment to continuous improvement these pursuits require much effort and rely on a capable staff as much as the equipment composing a longwall mining system success also depends on a close collaboration between mine operators and technology suppliers often as a strategic alliance both parties must begin with the end goal in mind which has to be attainment of safe production and profitability for both partners mine operators and technology suppliers must recognize their stake in the shortfalls as much as the successes constantly being prepared to contribute investments in time and money in pursuit of continuous improvement this is especially important in consideration of the long term nature of the relationship created by long equipment service life and large capital investments continuous improvement is necessary to meet the evolving expectations of stakeholders to the overall production process these stakeholders expect the result of zero harm as it applies to health and safety of employees local communities and the environment they also expect improving productivity and diminishing costs even in consideration of future mining conditions that typically are more adverse than past or present conditions for all of these reasons success at longwall mining is exclusively a result of organizations with aligned purpose detailed planning and controlled execution leading to sustainable results on the part of a wide diversity of involved stakeholders sublevel caving is a top down mining method allowing earlier production than sublevel stoping with less upfront development than traditional block caving in large scale operations it has the potential to be a factory method meaning highly repetitive with the cost and capacity benefits associated with high production moderate development requirements and maximum use of automated equipment bull and page 2000 in smaller mines it can be used as a selective method with lower production rates where the capacity benefits are less evident horizontal slices of in situ ore are progressively blasted and extracted with blasted ore and caved rock filling the void created by ore extraction figure 13 9 1 dilution can be managed and minimized by disciplined draw control as this process progresses caving propagates upward to create surface subsidence originally sublevel caving was used in weak ground that would collapse when the

timber supports were removed after a supported drift was pushed across the ore body from an access or perimeter drift the supports would be individually removed in a retreating fashion as each support was removed the ore would cave thus the name sublevel cave and be slushed out to the perimeter drift the next support would not be removed until the amount of diluting material became excessive overall the method was slow and combined poor ore recovery with high dilution later applications of the method saw it applied in relatively competent ore bodies with weak hanging walls in order to facilitate the caving process and avoid the creation of large voids the stronger ground required drill and blast to take the role of the support removal so in this version of the method it is no longer the ore that caves but the host rock above and immediately adjacent to the ore with the introduction of drilland blast sublevel caving could be utilized in ore bodies with footprints too small for natural cave development and or dipping geometry unfavorable for block caving in recent times the method has been applied to ore bodies with strong hanging walls where cave assistance techniques have been used as required current applications of the method are accessed by decline s and have sublevels established on approximately 20 to 35 m vertical intervals far greater than the early applications of the method due largely to advances in drilling and blasting technologies a perimeter drift is driven along transverse layout or perpendicular longitudinal layout to the strike of the ore body and is generally offset approximately 20 m from the ore waste contact some mines develop the perimeter drift in ore and recover it after production is completed from the drifts the perimeter drift is developed to access the production drifts for ore transportation services and ventilation the production drifts are spaced between 12 and 30 m on horizontal centers to enable optimal coverage for drilling and to allow for the downward flow of caved material they are staggered between the levels over the years the ratio of development to production ore has increased from approximately 1 3 to as high as 1 10 the production ore extraction process commences by placing either a slot drift on the far ore waste contact or placing individual slot raises at the end of each production drift the slot drift will have a slot raise at one end and a series of parallel rings generally on 2 to 3 m burden the width of drift reaching up to the level above once the slot has been opened to create adequate expansion room for the first production rings some slashing rings will start the production rings in the drifts until the production rings reach full height the production rings will usually be angled forward by 10 to 20 from vertical to assist the breakage and flow of the ore in the confined blasting environment and to help maintain geotechnical stability at the brow

the production ring burden is generally between 2 and 3 m the production rings drill up through the pillar between the production drifts on the level above as shown in figure 13 9 2 the layout of the method requires secondary ventilation to be forced into the production drifts throughout both the development and production phases of the level load hauldump lhd machines are used to extract the ore from the fired rings and transport it back to either stockpiles or orepasses located on the perimeter drift as the initial level is opened only the swell material from the blast approximately 40 is excavated because caving will likely not be initiated until production on the level is well advanced the extraction ratio of ore increases to a nominal value of approximately 80 to 100 over the subsequent two or three levels the reduction in the early draw rates is designed to minimize the likelihood of air blast until the overlying cap rock and or hanging wall starts to break up and flow in addition the reduced draw builds up a blanket of ore above the active drifts which can assist in minimizing the impacts of dilution in a blind cave as well as protecting the drifts when the sublevel caving is commenced directly below an open pit operational sublevel caving practice varies widely and consequently tonnage and mineral recoveries are also variable from operation to operation overall tonnage extraction rates for large scale sublevel caving mines will be approximately 80 to 120 of the blasted tonnage for approximately 65 to 95 of the total ore body mineral content of the rock drawn from each fired ring approximately 15 to 25 will be dilution more dilution comes at the end of each individual ring s life and varying practices mean that dilution results also vary from site to site power and just 2008 when extracting the ore body strict operational rules are required to ensure the method s success only after the upper level has been retreated a safe distance somewhere between one and two times the sublevel interval can the level below be commenced within the level the production rings need to be brought back in a planned sequence managing the lead lag between adjacent production drifts to minimize the impact of stress and explosive damage on neighboring drifts application of sublevel caving the suitable ore body characteristics that lead to the successful implementation of the sublevel caving method include cavable overlying host rock relatively large ore body footprint or a relatively weak hanging wall to initiate and propagate caving steeply dipping and relatively uniform ore body reasonably uniform grade distribution disseminated ore body with a low grade zone around the sublevel caving production area to minimize the impact of waste dilution competent ore body to minimize the excavation support requirements and improve drift availability to enable continuous production and visually different ore an

d waste properties although sublevel caving can be successfully applied to ore bodies missing some of these characteristics any such omissions generally lead to increased operational complexity power and just 2008 gravity flow the sublevel caving method relies on the gravity flow process to get the ore to the drawpoint as a result gravity flow has been a focus of sublevel caving research for many years gravity flow in the sublevel caving environment is unique in a number of ways the effects of blasting produce differential fragmentation within the ring both vertically and laterally draw occurs from drawpoints on multiple levels laterally offset resulting in multiple stages of flow each individual ring is subject to unique factors including geological structure drill and blast inconsistencies and differing draw patterns as in other fields of research the gathering of information on flow has been a gradual process with successive workers building on the data and theories of past workers early studies of gravity flow focused on small scale glass fronted physical models and over time expanded to larger three dimensional 3 d models such as shown in figure 13 9 3 these models involved the drawing of broken rock from geometrically scaled and well controlled environments so that the resulting movement and extraction zones could be studied in detail and accurately measured research using these physical models resulted in the promulgation of influential theories and relatively precise mathematical formulas to describe concepts such as particle velocity profiles the ellipsoid of loosening and of draw and the relationship between these bodies e g figure 13 9 4 it was later determined from full scale marker trials that the shapes of the flow and movement envelopes were not exactly elliptical janelid 1975 however they are close enough for the ellipse to be used as a basic design concept in sublevel caves power 2004b in practice these theoretical explanations of flow in bins and hoppers serve mainly to give a broad understanding of flow in a highly controlled environment taken in isolation they comprise a highly idealized single level description of granular flow with only a general relationship to granular flow in real sublevel caving environments however results from these models were useful in guiding design of early sublevel caves and provided important insights into the mechanisms controlling flow they also led to the development of some of the fundamental concepts underpinning the method which are well established today e g drawing drawpoints interactively and loading from side to side in drawpoints although model studies provided a basic understanding of gravity flow mechanisms full scale field trials were needed to obtain flow data in operational sublevel caves marker trials involve the placement of markers usually short lengths of steel piping or cable into internal marker rings before the e

xperimental production ring is blasted these markers are collected as the ring is mucked and the resulting data is analyzed to assess the flow patterns and ultimate recovery of the ring the earliest comprehensive trials were those carried out by janelid at gr ngesberg mine in sweden janelid 1972 janelid s marker trials showed gravity flow in the full scale environment but only with reference to recovery from the level on which the rings were blasted they were carried out on relatively small sublevel caving layouts approximately 3 m wide crosscuts on 7 m crosscut centers and 13 m sublevel heights for the relatively small ring tonnages drawn the trials showed high angles of draw and irregular draw patterns not quite elliptical in shape figure 13 9 5 in recent times marker trials have become more popular in response to significant changes in sublevel caving geometry which are partly due to improvements in drilling and blasting technology figure 13 9 6 shows a comparison of changes in sublevel caving geometry at the kiruna mine in sweden between 1983 and 2008 the kiruna 2000 program of marker trials was also single level trials although they did not have the same density of markers as the gr ngesberg trials and relied on visual identification of markers at the drawpoint for recovery of experiment information this resulted in only about a third of the installed markers being recovered gustafsson 1998 a diagram summarizing the results of these trials figure 13 9 7 indicates a total of 272 markers recovered and shows the percentage of markers recovered from various locations in the marker rings this analysis indicates that most markers recovered were almost directly above the drawpoint one important advantage of the kiruna trials was that they gave an indication of recovery from greatly upsized sublevel caving rings with crosscut widths spacings and sublevel heights all approximately double those used at gr ngesberg the kiruna marker trials showed that although mining scale had dramatically increased because of improvements in equipment technology and a focus on cost reduction granular flow properties had not changed leaving large parts of the fired production rings unrecovered if viewed only from a single level perspective marker trials carried out at the ridgeway mine in new south wales australia were the first comprehensive multilevel marker trials concentrating on granular flow through an entire sublevel caving system power 2004a 2004b carried out between 2002 and 2006 they involved more than 20 000 markers in more than 70 trial rings with data on marker recovery collected up to six levels below the level on which the marker rings were fired these trials involved the placement of markers in three planes of markers evenly spaced within the burden of the production rings nominally 0 65 m 1 3 m and 1 95 m from the next production ring to be fired figure 13 9 8 an important component of these trials

was the confidence level gained from the systematic magnet based marker recovery procedure which removed some of the uncertainty associated with the need to visually identify markers at the drawpoint and allowed analysts to more confidently assess areas of recovery and nonrecovery in individual rings analysis of data from these trials first introduced the concepts of primary secondary and tertiary recovery for the three main levels on which sublevel caving ore is recovered primary recovery is defined as ore reporting to the drawpoint on the level from which its sublevel caving ring was fired whereas secondary and tertiary refer to ore that reports on successive levels below this these trials also allowed testing of a number of theories that had gained traction in the absence of solid data and allowed other advances in understanding of granular flow in real sublevel caving environments similar trials were later carried out at perseverance mine in western australia hollins and tucker 2004 szwedzicki and cooper 2007 these trials aimed to determine the precise time of marker arrival at drawpoints and therefore required manual collection of markers in order to determine exactly when in the draw cycle each marker appeared at the drawpoint although this made multilevel analysis and completion of a large statistically significant trial program more difficult important data largely comparable to previous field trial programs was collected more recently research has led to the development of smart markers which can transmit a radio frequency signal to receivers that record their extraction location and time university of queensland 2007 the following sections discuss some of the new understanding that has been gained from full scale marker trial programs carried out since 2000 width of draw width of draw has consistently shown to be on average less than the width of the fired ring the ridgeway layout comprised 5 6 m wide crosscuts on 14 m crosscut centers as this mine was designed using a recovery rather than costreduction focus power and just 2008 here a 6 m wide drawpoint generally produced a primary recovery zone with a maximum width of between 8 m and 12 m therefore despite the relatively close crosscut spacing interaction between primary drawzones rarely occurred figure 13 9 9 illustrates this situation for sections taken through two adjacent crosscuts all the different stages of recovery were recorded and are represented and it is clear that the primary recovery zones do not interact prior to the completion of these trials a theory known as interactive draw as proposed for sublevel caving had been put forward bull and page 2000 and influenced the design of some sublevel caving mines this theory suggests that by drawing from adjacent sublevel caving crosscuts interactively or in close sequence zones of low pressure can be created between the drawpoints allowing the rock to flow at lower angles than in dra

wpoints that are drawn in isolation prior to this physical modeling results had also been used to suggest that drawing from an interactive front rather than in isolation would decrease dilution ingress from the sides of the rings and reduce hang ups janelid 1974 in reality interactive draw did not result in primary drawzones flowing to the full width of the ring and interaction with adjacent crosscuts power 2004a similar trials completed at perseverance mine also indicated no evidence of interactive draw in any of the trial areas hollins and tucker 2004 summary results from these trials clearly show the episodic nature of flow from individual rings those at relative narrow draw widths were similar to ones seen at ridgeway figure 13 9 10 figure 13 9 5 also shows albeit for reduced production tonnages that recovery zones for rings spaced on 7 8 m centers at gr ngesberg are not interacting even though strict interactive draw controls were used in these experiments flow here was only allowed to develop to half the height seen at later mines given the smaller sublevel spacing figure 13 9 7 also suggests no interaction between adjacent crosscuts at kiruna even though these results summarize the marker recovery results of all the trials completed in its program rather than of an individual marker trial this is not surprising given the relatively large crosscut spacing at kiruna mine in practice every mine will have a range of differing input conditions that will affect marker trial outcomes added to which are the significant differences seen between trials at individual sites even so the results on width of draw at different sites are relatively similar depth of draw in approximately 20 of cases draw at ridgeway never extended to the full 2 6 m depth of the ring in more than 50 of cases the primary flow zone only reached full ring depth late in the draw cycle for a small percentage of the ring height figure 13 9 11 shows an example of the primary draw which is both narrow and shallow there is some conjecture about the cause for shallow draw as seen at ridgeway visual observations indicated that unrecovered rock at the back of the ring sometimes retained the original rock mass texture figure 13 9 12 however if a drawpoint opened up the less mobile ore at the back of the ring often collapsed into the drawpoint this indicates that the rock was being fractured but the fragments were not being effectively separated from one another however when hangups occurred and the lhd was used to dig farther into the base of the ring greater depth of draw was sometimes achieved with corresponding higher recoveries the shallow draw at ridgeway shows a variation from the argument that early dilution entry originates behind the ring from the waste side of the ring like other mines gustafsson 1998 hollins and tucker 2004 ridgeway commonly measures early entry of rock from outside the ring at approximately 20 draw

however this rock has been shown to originate from the level above the ring due to flow piping preferentially up the solid face of the ring one of the assumptions underlying the sublevel caving method is that the entire ring is fractured effectively compacting the waste ahead of it and allowing the ore to be drawn relatively undiluted from the drawpoint some however have shown that the space available to allow adequate swell and particle disassociation for all portions of modern sublevel caving rings is insufficient hustrulid 2000 the ridgeway results may support this hypothesis indicating that the rock at the back of the ring is not being mobilized because of the combination of lack of opportunity to swell and cushioning of the ore at the free face of the ring it is also possible that differing modes of dilution entry are occurring at different mines where different blasting and draw control practices are used marker trials allow the opportunity to learn more about the sources of dilution entry and for mine operators to refine drilland blast and draw control practice to postpone and reduce dilution entry ore recovery over multiple levels before completion of the ridgeway marker trials assumptions about levels of primary recovery were significantly higher than those actually measured however although a lower than expected primary recovery is achieved on average approximately 50 for 120 ring tonnage extracted much of the remaining ore is recovered by draw from levels below even with this low primary recovery percentage ridgeway s closely spaced drawpoints favorable ore body geometry and tightly managed draw control enable it to recover a high proportion of the ore fired with relatively low levels of dilution power and just 2008 this indicates the importance of viewing the sublevel caving system as a multiple level system rather than making design decisions based purely on recovery from the primary level alone figure 13 9 13 illustrates how ring recovery is spread over many levels at ridgeway figure 13 9 14 indicates that an equation can be fitted to this data with a good fit it is this general sense of order within an apparently chaotic system that has enabled the development of a number of forecasting tools that can be used with reasonable accuracy in operational mines for estimation of recovery development of draw control strategies and optimization of recovery through the targeting of residual grade back break back break is a phenomenon that is difficult to simulate effectively in physical models but can be seen in many marker trials and generally in sublevel caving production drifts it occurs when a ring blast breaks rock away from the next ring to be fired in the crosscut this back break often occurs at the brow but can also be present along the full height of the production ring depending on ground conditions this was noted in marker trials carried out by janelid at the gr ngesberg mine janelid

1972 and has also been seen at ridgeway power 2004a and perseverance hollins and tucker 2004 back break is still ore recovered on the primary level and thus can be included from a grade forecasting perspective with primary recovery its effect on blasting and flow however can be significant if a blast produces high amounts of back break rock flows through this broken area effectively changing the brow position of the blasted ring and reducing the likelihood of drawing to the full ring depth figure 13 9 15 figure 13 9 16 shows an example of a ring that has suffered significant back break indicating that back break from the previous ring has broken almost to the full burden of the next ring to be fired it is probable that this back break influenced the performance of the blast ring and possibly recovery in this example back break extends to the base of the ring however primary flow still does not reach the central marker plane for the lower half of the ring this indicates that even when ore is likely to have been well fragmented due to a high relative powder factor shallow flow can still occur parts of a ring that have been damaged by the previous blast can be significantly thinner than the rest of the ring they will be blasted with a greater effective power factor causing the blast to behave differently than intended generally the zone of greatest explosive concentration produces most back break because this is also the area of greatest explosive concentration in the next ring this accentuates changes in the actual blasting powder factor distribution for this ring chaotic and variable flow figure 13 9 16 shows how back break in one ring can affect blasting and flow performance of the next ring to be blasted because the blasting and flow behavior of the vast majority of sublevel caving rings cannot be practically measured it is not known what effects are being generated to other rings if it is accepted that performance in one ring will affect performance of rings around it and that these events cannot be predicted it follows that sublevel caving flow should be treated as a chaotic system some of the parameters producing significant variability in sublevel caving systems include rock mass properties blast design ability of charging crew to implement blast design blocked holes etc hole deviation blasting performance of rings around the current ring hang up occurrence impacting draw and grade variations impacting draw although it is impossible to predict the flow outcome from a particular ring the research program carried out at ridgeway identified five common primary flow behaviors by which most of the trial results could be classified these are shown in figures 13 9 17 and 13 9 18 although general primary recovery zones are shaded in these figures it should be assumed that the rippling flow effects discussed later would result in portions of these drawzones being left for recover

y on subsequent levels or not being recovered at all figure 13 9 19 shows the relative frequency of these draw behaviors and that the standard and standard shallow draw classes produce the highest ring recoveries calculated after the secondary level as depth of draw is reduced width of draw increases if apex mobility is low width of draw is also likely to be low and recovery is also likely to decrease although not discussed in detail here coupling full scale flow assessment with analysis of the drill and blast factors producing these flow types is also important marker trials and other drill and blast research programs can be used to develop a greater understanding of drill and blast and its effect on flow can be used to optimize operational drill and blast design this research is also used as the basis for layout blasting and draw control design for new sublevel caving mines many of the figures included here show examples of ripple patterns in ore recovery these patterns produce regions of recovery interspersed with sections of nonrecovery resembling ripple patterns produced in sand due to water flowing normal to the ripples researchers suggest a process whereby rock flowing toward the drawpoint encounters zones of lower mobility and continues around these on its original course these ripple patterns indicate that although a flow zone may progress through a certain region in the ring parts of this region are often left unrecovered ripple patterns are often at slightly different orientations for different draw classes suggesting that flow has progressed at different orientations for different recovery classes the relatively narrow draw pipes progress subvertically from individual drawpoints to heights of at least 55 m as illustrated in figure 13 9 20 which shows the results of a secondary draw trial in this trial no primary draw was taken in the rings on the upper level allowing secondary draw attributable to the rings below to be studied the primary recovery zone in the lower ring is produced by the same flow event as the secondary recovery zone on the level above flow from the ring below appears to have progressed into the center of the ring above left the secondary flow zone in the upper ring seems to have been preferentially broken in the blast possibly due to more concentrated explosive density in the center of the ring it is likely that low mobility in the apex of the ring below has caused this flow outcome also resulting in very little secondary recovery from the upper ring on the right episodic flow patterns from individual rings were also identified in the perseverance marker trials hollins and tucker 2004 similarly field experiments conducted at the kiruna mine showed pulsations of waste entering the drawpoint with increasing levels of waste appearing with tonnage drawn figure 13 9 21 these pulsations were believed to be related to differential mobility in the blasted ore from different pa

rts of the ring and waste rock the results of the early model based research into granular flow in sublevel caving led to advances in design and draw control practice and to a general understanding of granular flow in many cases the data recovered from marker trials now supplements or has replaced information previously gained from physical models such trials allow testing in theoretical thinking on sublevel caving granular flow in full scale operational environments but must be designed and analyzed as part of a systematic scientific program to be completely effective they often produce results that lead researchers in totally new directions of thought other results through full scale research of the more recent understanding of flow particularly in its multilevel nature are improvements in computer based forecasting and optimization capabilities several different methods of numerical flow modeling specifically designed for sublevel caving environments are now available some of these methods not only accurately forecast monthly production grade but also the targeting of high grade zones and regions of the ore body where extraction has been low in some cases they allow shutoff simulation to produce detailed drawpoint by drawpoint production targets in situations where visual or weight based draw control systems are ineffective with the continued development of new technology targeted at sublevel caving e g smart markers and improved computing capacity for research into drill and blast effects and numerical modeling of granular flow it is likely that sublevel caving will continue to be a productive and economically viable mass mining method mine design layout selection is a critical success factor in all aspects of a sublevel caving operation the layout directly relates to the ore recovery drill and blast effectiveness excavation stability and the cost of mine development because it is a fundamental success factor considerable attention needs to be given to selecting the layout prior to commencing mine establishment and determining the critical success factors for the mine a mine focused on low cost production will implement a different layout than a mine focused on maximizing recovery mine layouts will probably change over the operating life of the mine so the original layout is often modified as a result of lessons learned under operational conditions layout orientation layout orientations are generally classed as either transverse or longitudinal the transverse layout is generally applied in wider ore bodies greater than approximately 50 m where the parallel production drifts are driven across the ore body according to figure 13 9 22a for widths below this the production drifts are generally driven along the strike of the ore body in a longitudinal fashion according to figure 13 9 22b the longitudinal layout may contain as few as one drift on a level depending on ore body width the choice of orientation is

predominantly determined by the width of the ore body but can also be influenced by geotechnical and economic factors the transverse layout is generally more productive than the longitudinal because of the number of active production faces that can be made available some mines will operate with a mixture of layouts due to a variable ore body geometry although transition between the two styles is difficult in a vertical sense because of the criss cross nature of the layouts however if transition is required between the layouts it may be prudent to skip a level and recommence sublevel caving operations with the new layout and allow the skipped level to cave in rather than trying to perform complex production ring drilling and blasting this technique was successfully employed at the perseverance mine wood et al 2000 for either layout type the supporting development is similar with a perimeter drift containing either orepass tipping points or truck loading stockpiles the perimeter drifts are also used to provide access to the primary ventilation circuit and enable the installation of secondary fans to allow airflow to be delivered to the production working areas on the level layout dimensions sublevel height the sublevel height is generally measured as the vertical floor to floor distance between the levels it is determined by the dip of the ore body and the physical limitations of the production drill and blast equipment if the ore body is vertical then dip is not a constraint on sublevel spacing however as the ore body flattens the sublevel height needs to be reduced to avoid incurring excessive amounts of hanging wall dilution as well as minimizing the amount of unrecoverable ore on the footwall over the years advances in drill and blast equipment have seen the sublevel height increase from less than 12 m to more than 30 m in some cases the key criterion for drill and blast is to ensure that satisfactory fragmentation is achieved through the entire ring to achieve this the drill holes must be evenly spaced straight and adequately charged with the appropriate type and strength of explosive the proposed drill pattern is an integral part of selecting the overall sublevel layout and impacts not only the sublevel height but also the drift dimensions and the drift and pillar spacing production drift spacing important factors in selection of production drift spacing include potential ring shape produced by the combination of sublevel height and production drift spacing will it produce a ring shape that can be efficiently blasted granular flow properties inherent to the mine based on drill and blast procedures and rock mass characteristics and the likely recoveries to be achieved value of the mineral being mined and whether a costreduction or recovery emphasis strategy will be followed geotechnical factors and production drilling and charging factors for maximum ore recovery the entire level would theoret

ically be excavated similar to a coal longwall operation but this is impractical therefore a trade off is required between the amount of ore potentially recovered on its original level and the amount left in the cave for recovery hopefully on lower levels with more associated risk to its ultimate recovery thus the spacing of the production drifts is strongly influenced by the operating philosophy of the mine the more cost focused a mine is the wider the spacing between the production drifts because this reduces the amount of expensive development required power and just 2008 figure 13 9 23 illustrates this for two different layouts showing theoretical primary recovery zones only recovery focused mines will try to place the production drifts as close together as practical in an attempt to maximize the amount of ore recovered on the primary level to minimize the effects of dilution production drift shape the shape of the production drift impacts the mine s ore recovery and productivity ideally the production drifts are as wide as possible to allow the maximum flow of ore to further assist with flow the drift roof should be as flat as practical while still providing adequate support to the roof and brow a wide and flat roof also assists in improved spacing of the production drill holes which limits the impacts of blasting on the brow as well as reducing the amount of back break if the drift roof is overly arched the flow can be preferentially funneled to the center of the drift leading to early dilution in some cases a trade off will be necessary i e an arch may be required to produce a wider stable drawpoint greater drift widths also improve the productivity of the lhd machines an lhd can tram faster in a drift where it has plenty of clearance at the muck pile the width is also an advantage for the lhd because it is able to load from side to side ensuring a more even flow of material into the production drift as illustrated in figure 13 9 24 in addition the loader is able to extract material from deeper into the drawpoint because it is not extracting the material from the center of the drift which has rilled the farthest out from the brow as illustrated in figure 13 9 25 this further improves ore recovery and influences the burden of the production rings because the cave material loosens in front of the next ring to be fired minimizing the height of the drift roof restricts the rock s ability to rill too far from the brow as previously stated if the depth of material excavation by the lhd can be increased then the overall ore recovery will be increased the drift height however is usually a trade off between length of rill and size of machinery required to achieve the desired production rate although it is possible to obtain custom made machinery to fit the optimal drive height this strategy can have drawbacks if the machinery fails to meet production targets as machinery replacements are g

enerally not immediately available in the authors experience it is generally the production drill that limits the drift height and as the large scale operations push for longer and longer production holes this is likely to remain the case low profile lhds are generally available to achieve the desired production rates but the desire to drill with longer and longer drill rods for greater hole accuracy is forcing average drift heights higher production drill and blast the drill and blast requirements of a sublevel caving mine are unique because the blasting environment is always choked this means that every production ring is blasted against the combination of previously blasted ore and waste sitting in and around the production drift figure 13 9 26 because of the environment in which the blasting occurs i e the only free face is the roof of the production drift the blast s function is different from nearly every other production drilling situation with sublevel caving blasting the intent is to initially compact the caved material in front of the blast to create a void area for the fired ring to access after the fired ring is in the void area the intent is to fragment the ore slice as much as possible so that it will preferentially flow when mucking commences although this is the intent the reality is somewhat different the nature of the fanned ring pattern used in sublevel caving and the reducing ratio of development toproduction work against the intended blasting criteria with the reduction in this ratio the only true swell available to the blasted rock is reduced and in some cases is 10 while the blasted rock itself still wants to swell to 30 of its in situ volume this lack of initial swell results in the holes fired first and those sections of the hole close to the drift have the opportunity to swell and move into the drift as desired while the ends of the holes have limited opportunities to swell and move in addition the fan nature of the drill pattern ensures that explosives are concentrated in the center of the ring relative to the outer holes this creates smaller more uniform fragmentation in the center which is thus more mobile than the material at the edges these fragmentation and flow characteristics affect the gravity flow and ore recovery performance for the mine trout 2002 with the diamond layout style of sublevel caving production drifts a wide variety of production drilling patterns can be used with the wider spaced production drifts a pattern involving larger and longer holes will generally be used although as the production drifts are brought closer together the drill holes will interact when this occurs the production rings often need to be offset by approximately 1 m to reduce the chances of sympathetic detonation between the holes or damage to the neighboring ring when the rings are fired steeper side holes will lead to better funneling of ore to the drawpoint and smaller dea

d zones at the sides of the ring but also to higher apexes at the top of the ring which can be more difficult to blast effectively the production drill rings are generally inclined forward by 10 to 20 this inclination assists in feeding the ore preferentially into the drift by providing a small overhang to the diluting material directly above the inclination also ensures that the holes are all drilled on a similar plane for improved blasting results and additional protection of the ring brow because prevailing structures in the rock mass can sometimes adversely affect rings inclined forward at similar angles the amount of forward inclination may need to be adjusted in different parts of the mine in order to achieve optimal results the sublevel caving production rings can be charged with either ammonium nitrate fuel oil anfo or emulsion products prior to any charging activities the production holes should be dipped by the charging crew to ensure that the holes are clear of obstructions and are drilled to the correct depths often charging personnel must stand on the muck pile to charge the ring as the charging unit frequently cannot get the operators to the production ring because of the rill of broken rock in the drawpoint and the height of the drive figure 13 9 27a this can lead to significant risk of injury and inferior drill and blast results safety hazards include the muck pile rilling unexpectedly or trip hazards resulting from the uneven nature of the muck pile on which the operators work to provide a safer charging environment the preloading or precharging of multiple rings can be implemented figure 13 9 27b although this is subject to the local regulators and may take time to get approval precharging relates to fully loading and priming the production rings whereas preloading refers to the practice of only placing the explosive in the hole but not the detonator which is added immediately prior to blasting when utilizing precharging the tails of the detonators are tucked back up the hole so they do not impede the lhd when it is mucking the rings in front the tails are only pulled out when the ring is ready to be hooked up for blasting the implementation of precharging loading in recent years at a number of large scale operations has had significant benefits for those operations including improved safety of personnel around the drawpoint brows a decoupling of the charging and mucking activities the removal of production disruptions caused by hole blockage and dislocation the elimination or significant reduction in the number of production ring redrills and improvements in draw control as less mucking is needed to prepare the ring for firing drawbacks to precharging loading include the need to have a number of production rings predrilled in order to allow the charging and drilling processes to be independent from a safety perspective when a precharged loaded ring is behind a misfire the scope is lim

ited to get a drill rig back to assist with the misfire unless the drill rig has a remote operation function or the explosives are washed from the precharged loaded ring for this reason precharging is often easier to implement in mines with better geotechnical conditions where charged holes are less likely to be dislocated or brows lost if the ore body contains aquifers or seeping water precharging loading may be also unsuitable when using anfo because the explosive column can be desensitized or washed out by the time the ring is ready to fire emulsion should overcome this issue because of the choked blasting conditions of sublevel caving rings a single production ring per drift is generally only fired at any one time firing of multiple rings may be undertaken in areas of very poor ground conditions but this introduces the risk of misfires and leaves pillars within the cave which can cause problems on the level below it also leads to lower overall recovery because most of the ore in the ring farther from the brow cannot flow to the drawpoint and can be occluded by waste flowing from above depending on infrastructure constraints and required production rate some mines fire rings in multiple production drifts at the same time to ensure broken stocks are sufficient for the lhd fleets hang ups are a normal feature of the sublevel caving mining method with up to 50 of rings hanging up at some stage of the mucking cycle in mines with more closely spaced drawpoints the hang up can sometimes be cleared by removing additional material from the production drifts on either side of the hung up drift this material removal appears to disturb rock in the affected drawpoint by acting to release pressure on either side of the hung up ring other hang up retrieval methods include water cannons and as a last resort explosives because they often damage the drawpoint brow if a hang up cannot be cleared within a specified period the next ring in the production drift must be fired and the mining process continued although production drill and blast design is an input into the overall layout of the sublevel caving mine it can be readily changed as the characteristics of the ore body become more fully understood the shape of the production ring may not vary significantly but the drill and blast parameters may design parameters including hole diameter number of holes in the ring drill hole toe spacing production ring burden and dump angle and explosive column densities and lengths can all be modified with relative ease when adjusting the production drilling parameters it is important to take a holistic view of the sublevel caving operation in order to ensure that the changes made will not dramatically and negatively affect other parts of the process services the active ore production areas are all dead end headings a single fan may be used to supply up to four production drifts as the ducting can be tied off in the headings that

are not in use if multiple secondary fans are required on the level they may need to be set up in a cascading style option a figure 13 9 28 this means that a fan will be placed downflow of another fan and in essence will receive the used air from the upflow fan this cascade style of secondary ventilation can lead to lost production time if there is a significant amount of dust in one of the drawpoints upflow of the secondary fan because all the dust will be blown into the headings serviced by the downflow fan each level is set up as its own primary circuit so the levels do not interact with each other primary intake and return raises with fixed ventilation infrastructure are used to direct the appropriate amount of air throughout the mine figure 13 9 28 shows two styles of level ventilation option a is for mines using the main access decline as their fresh air source whereas option b is for mines that heat or refrigerate their intake air the cave itself may provide a path for airflow after it has broken through to the surface though this will be determined by the porosity of the cave column and the pressure differences between the surface and the cave workings generally movement of air through the caved rock will be in the form of leakage rather than a volume that can support mining activities individual mining circumstances will ultimately determine how much air passes through the cave zone area water and compressed air are reticulated to each level to support the mining activities after development of the level is completed the compressed air line may be recovered unless required by the production drill the water line is used for production drilling and to feed the drawpoint sprays mounted on the drift roof to keep the dust levels down during mucking drainage on the sublevels is important in ensuring that maximum production rates are achieved the production drifts are usually driven on a slight incline of between 1 in 50 2 and 1 in 40 2 5 to remove the pooling of water along the drift or at the face in addition to the introduced water for the mining process water is also associated with aquifers that have been breached during mining sumps are generally located in the perimeter drift with a drain hole drilled to the level below the production water can then be collected at a mine dam located lower in the mine and pumped out via the mine dewatering pump system communications systems are run onto the production levels for radio communications as well as to control any automated or semiautomated mining equipment including production drill rigs and lhds in some of the larger scale sublevel caving operations such systems may also be utilized to transmit real time production data to a central production management area the larger scale sublevel operations using automation or the tele remote operation of equipment are doing so in an attempt to remove workers from hazards associated with drawpoint brows and in t

he general underground environment the lkab mines in sweden are leading the way in this field they use the automation of machinery to provide production increases by allowing the machines to work extra hours while the operators are located on the surface or an area of the mine not impacted by the mining process especially the blasting and clearance cycle geomechanical factors with sublevel caving a number of geomechanical factors must be considered when either laying out the mine or making large scale operational decisions these include quality of the ore body and hosting rock mass orientation and magnitude of the major and minor stress parameters and presence of large scale and microstructures prior to establishing a sublevel caving mine an assessment of the likelihood of natural caving is required the assessment is a specialized geotechnical task and is beyond the scope of this chapter however inputs for these techniques rely on a number of key geotechnical parameters so it is imperative when diamond drilling into a deposit that may be applicable for caving that all of the relevant information is recorded if stability assessments indicate that the ore body will not cave naturally then sublevel caving may not be suitable unless reliable techniques for cave preconditioning and or caving assistance are available the orientation of the stress field will often determine the retreat direction of the sublevel caving production front if there are no other influencing factors it is preferable to retreat perpendicular to the major principal stress direction by retreating in this manner the effects of the stress field can be shadowed out after production has been initiated on a sublevel this creates a distressed environment for the majority of production headings in the mine because the stresses will be forced to wrap around under and over the production areas stress that builds up around the periphery of the working areas is called abutment stresses which can result in rock movement and bursting depending on the rock strength tostress ratio the ground support requirements are therefore greater in the perimeter drives than in the production drifts the bullnose section of the pillars at the turnout points of the production drifts from the perimeter drive are especially susceptible to stress damage as shown in figure 13 9 29 because they are reduced in size with the retreating of the cave front production drive stability can also be impacted by the abutment stresses as production on the level above passes over the lower level production drifts the preferred ground support system in a sublevel caving operation is one that can yield slightly without failing when the abutment loads come on but is still sufficient for working at the brow in the distressed conditions some ground support combination examples include grouted rebar with 50 to 75 mm of fiber reinforced concrete or split sets with wire mesh identifying mi

ne scale and minor structures is important when laying out the mine because they can be used to assist in the propagation of the cave or in controlling the effects of the abutment stresses should the major structures not be favorable to the cave s progression then considerable care needs to be taken to monitor caving progress because it may be necessary to reduce or stop production in certain cave areas to reduce the impacts of these structures the orientation of the minor structures is also important because it can influence the inclination of the production rings such inclination on a plane parallel to the minor structures can reduce the number of production hole dislocations and therefore improve productivity water bearing aquifers or surface water courses that is streams and dams in the caving area also need to be identified prior to the commencement of caving activities to ensure that there are no unexpected water or mud inrushes during the operation of the mine where an aquifer or water course cannot be dewatered prior to caving extreme care should be taken and evacuation procedures implemented as the cave encroaches on the water bearing area generally as the cave propagates it will ensure that a large number of fractures are created in the cap rock which will breach the aquifer and potentially drain it before the main cave back moves through the water bearing zone however in extreme cases uncontrolled mud flows can be generated by excessive and poorly controlled draw in combination with high inflows of water either from aquifers or channeling of precipitation through subsidence zones and suitable clay mineralogy when this is a risk use of a robust inrush hazard management plan with appropriate triggers and controls is beneficial prior to commencing a recent sublevel caving operation at the kiruna mine engineers were forced to drain a small lake while engineers at the ridgeway mine identified an aquifer system below the surface and monitored it as the cave progressed noting only minimal water reporting to the production drawpoints as the majority of the water was absorbed in saturating the cave column production draw control draw strategy and production management are extremely important in sublevel caving operations to enable the accurate prediction of ore and waste flow from the cave after production commences the draw strategies and production management plans used at each mine need to be variable and flexible in order to cope with unforeseen circumstances such as the early inflow of waste hang ups geotechnical instability and arrival of large oversize material at drawpoints draw strategy the objective of a draw strategy is to maximize the overall ore recovery while minimizing the impacts of dilution to achieve this an effective sublevel caving mine relies heavily on a draw control system that is variable and adaptable to the constantly changing operational situations including unscheduled equipment br

eakdowns and production incidents effective draw control practices also rely on an understanding of the cave s flow characteristics and the grade distribution in and above the ore body being mined any adopted draw control system needs to include guidelines to allow the frontline management to make tactical decisions such guidelines will be grouped according to authority levels ranging from the lhd operator to the general manager of the mine in mines with a distinct visual or physically distinguishable difference between the ore and waste rock guidelines will dictate the percentage of ore to waste in the drawpoint when production must stop for a particular drawpoint to allow decisions to be made mines without this distinction must rely solely on the tonnage drawn from the drawpoint followed up with a sampling system regular monitoring of the drawpoint status is imperative to achieve overall efficient draw strategy and to ensure that underperforming or troublesome drawpoints are identified early so their potential impact can be minimized draw strategies can be classed as either independent or interactive and are determined by the mine layout and overall mine objective these terms refer to drawpoints where tactical extraction targets are linked to those of surrounding drawpoints interactive and those where tactical production targets can be generated without reference to surrounding drawpoints independent theories discussed previously concerning the widening of draw columns due to interaction through interactive draw do not have bearing on these draw strategies independent draw is often used by operations with a limited number of production drifts available for mucking under these circumstances each individual drawpoint is often depleted before moving onto the next one the penalty for this type of draw strategy is a greater risk of dilution piping in from above and impacting adjacent production drifts interactive draw style operations rotate the lhd between a number of neighboring drawpoints in a fixed sequence which is repeated until all of the drawpoints have been completed and the next set of rings fired the interactive approach enables the blasted and caved material to be extracted over a larger lateral area thus reducing the opportunity for dilution piping depending on the ore flow characteristics of the mine and the mine layout selected either approach may maximize the overall mine performance as previously identified drawpoints can be operated individually or in groups in mines with substantial numbers of drifts on each level the practice of grouping a number of drifts together and treating them as a single entity can substantially reduce the complexities of managing the operation on a short term basis the orientation and shape of the production front will be influenced by the stress field and major structures no standard exists for the shape of the cave front because all mines will have varying conditions and

operating requirements the three most widely used cave front layouts are the flat en echelon and stepped as shown in figure 13 9 30 when selecting a cave front thorough consideration needs to be given to the lead lag rules that will be used to operate the sublevel caving the term lead lag refers to the maximum and minimum spacing between either individual or groups of drawpoints being retreated across the levels these rules are 3 d because they also include the level above and the level below the specific level when defining the lead lag rules the main influencing factors are stress field constraints which relate to the minimum distance required between production fronts on adjoining levels to preclude stress damage damage can also occur when drawpoints retreat past their neighbors on the same level the impacts of stress can be mitigated by increasing the amount of ground support in the sections of the drift though additional support will be limited by the ability to costeffectively change the lead lag distances another operational limit to lead lag distances is the interaction of drilling and charging activities within adjoining drifts most operations will have an exclusion limit around a charged hole which must be considered when establishing lead lag rules compaction of the cave column is also another potential factor if a level retreats too close to the one above stability of the cave in front of the upper level may be compromised and the ore blasted on that level may be more difficult to recover conversely if the distance between the vertical levels is too great it may take too much time for the lower level to pass under the upper level and the cave may recompact or reconsolidate depending on its characteristics causing considerable production and ore losses production management a number of factors impact the theoretical production rate of a sublevel caving operation including ore body geometry cave fragmentation of both ore and waste mine layout and excavation size equipment employed fleet size and number and materials handling capacity as the sublevel caving method is now often used in more competent ore bodies opportunities for larger stable excavations exist this is transforming production rates for sublevel caving operations as larger equipment is used in most sublevel caving mines the majority of activity occurs within the confines of the level with little activity in the access declines this makes the method efficient with some operations achieving 35 000 t operator per annum as relatively few operators are required underground once the mine is operational overall the aim of sublevel caving mining is to become as factory like as possible this approach is achieved by taking a holistic view of the mining method from the face through to the surface in the understanding that overall mine productivity is governed by the lowest producing component with the increase in equipment siz

e and technology larger fragmentation can be accommodated at the drawpoint but this may cause downstream issues for the shaft conveyor or truck fleet used to haul the material out of the mine as such it is imperative that the mine production system is integrated with the materials handling system as well as with the mill to ensure overall optimum mine performance recovery and extraction recovery factors relate to the percentage of mineral recovered from the ore body relative to the amount blasted mineral recovery from sublevel caving operations can be above 85 of the quantity of in situ mineral blasted however this recovery can be inflated by inflow of low grade mineral contained in diluting rock from outside the mining boundary extraction or tonnage factors relate to the percentage of tons drawn relative to the tons fired this percentage varies from site to site but is generally between 80 and 125 when the extraction rate is more than 100 this is termed overdraw depending on the draw control strategies and grade distributions within the ore body and host rock individual drawpoints may be drawn in excess of 1 000 and still be returning material above the shutoff grade the grade at which drawpoints should theoretically be shut down this generally occurs when previous levels of the mine have been significantly underdrawn because of geotechnical or mine design factors some ore bodies will have physical differences between the ore and waste which are readily identifiable for example color or weight this allows mine personnel to assess in real time whether or not a drawpoint is still running at the required grade in these situations the mines are able to more selectively recover ore left behind on previous levels however they must be careful to ensure that their grade control methods are accurately calibrated in order for their shutoff targets to be met other operations with no readily identifiable physical difference between the ore and waste will often run to a tonnage limit and stop draw at a point independent of the drawpoint grade this tonnage strategy can lead to a considerable amount of mineral being left in the cave which these operations will then try to recover over subsequent levels by increasing their planned draw rates using a broader less selective strategy under these systems total tonnage factors for the mine can be slightly higher than situations where selective grade control can be applied fines migration can also influence extraction strategies especially when sublevel caving is applied beneath open pits with significant exposures of weak weathered materials or in ore bodies with weak hanging walls the cave column applies a significant amount of comminution to the rock as it flows so weak material is often reduced to a small fragmentation size referred to as fines fines can also be created through the overcharging of the blastholes due to the properties of gravity flow fines material will

descend through the cave column faster than the coarser rock fragments sometimes prematurely closing drawpoints affecting ore handling system efficiency or creating the potential for mud rush when combined with the introduction of large amounts of water to the cave column cavability the caving of the host or cap rock above the sublevel caving operation is crucial to the success of the mine mine scale structures and the general geotechnical conditions of the host rock influence the rate of cave propagation and the size of the footprint required to ensure caving there are two distinct phases to the caving of the cap rock 1 initiation of caving and 2 the continuation of caving as the mine continues deeper figure 13 9 31 should either of these stages not occur and the cave stalls or forms a stable arch there are considerable safety risks relating to the creation of voids and the potential for air blasts cap rock failure may occur as either a time or production dependent failure time dependent failure occurs when the cap rock yields and fails only at intermittent periods a mode that can lead to significant cap rock failures depending on how much swell room is available production dependent failure occurs when the cap rock unravels almost immediately after an air gap becomes available largely controlled by extraction of the blasted material time dependent failures are often accompanied by a small number of significant seismic events whereas production dependent failures tend to have a significant number of seismic events of a small magnitude the mode of cap rock failure will determine the production ramp up rate of the sublevel caving mine because a number of safety hazards can be created if the initial production rates are not matched to the cave propagation rate the most significant safety risk is the creation of an excessive air gap which can lead to a dangerous air inrush air blast when the cave back yields in order for rock to cave some form of air gap has to be present the geomechanical properties of the rock mass the footprint of the production level and the draw control strategy of the mine determine the size of the air gap that can form sublevel caving operators must manage their operations to ensure that the size of any air gap that forms is limited and does not constitute a hazard to mining personnel void monitoring is a critical process when establishing a cave under any conditions the results from the cavemonitoring system need to be cross checked with the production figures to determine the type of cap rock failure mode that is in place and to ensure that there is a suitable coverage of blasted and caved material above the active production drawpoints to prevent a dangerous inrush of air into operational areas it may be prudent to reduce or even stop production for a period of time in the sublevel caving to allow the cap rock failure a chance to catch up and return to the safety operating guidelines

continuing to produce while creating a significant air void can have catastrophic results when the cap rock ultimately fails to maintain an understanding of how the cave is progressing cave monitoring throughout the life of the operations is important a cave monitoring system should consist of a number of complementary measurement techniques to allow for the cross referencing of results microseismic systems placed outside the expected cave column give an indication of the seismogenic zone which exists just above the cave back duplancic and brady 1999 and can be used not only until the cave breaks through to the surface but also as it relaxes over the life of the mine open holes into the cave back area can be used until the cave reaches a predefined distance from the surface where it is no longer safe for people to travel above the cave back the open holes can be plumbed for depth using weighted wire rope and in some cases also provide access for borehole cameras as the cave approaches the surface the open holes can be used to place extensometers and timedomain reflectometers which allow for remote monitoring of the failure of the cap rock subsidence as a caving method sublevel caving will create a surface expression and a crater that deepens with time this surface movement zone is referred to as the subsidence area understanding the likely size of this area is imperative when locating major infrastructure both on the surface and underground motion sensors and aerial surveys figure 13 9 32 can be used to monitor the growth of the subsidence area to determine if it correlates closely with the production being drawn from the mine should a discrepancy occur between the surface growth and mine production a review of the potential void creation should be undertaken to ensure there is no risk to the work force as the subsidence zone creates a direct link from the surface to the underground workings it is imperative that water inflow restrictions are placed around the subsidence zone where practical for sublevel caving mines below an open pit or at the bottom of a valley this can be a significant issue as these areas are catchments for rainfall in these situations monitoring of the rainfall and the saturation of the cave column may be utilized to determine the appropriate operational response the flow of water into the subsidence area may not only result in additional pumping from the workings but can create a mud rush issue when combined with the fines located within the cave column in the simplest case of a uniform vertical ore body the initial cave crater will generally have vertical walls until the crater reaches a sufficient depth for slumping failure to occur the surface expression will gradually expand as the crater deepens and an exclusion limit needs to be implemented the size of the subsidence and fractured zone around a cave will vary due to the host rock conditions major structures and so forth but a rule o

f thumb estimate is 65 from the lowest level block caving is a mass mining system that uses the action of gravity to fracture a block of unsupported ore allowing it to be extracted through preconstructed drawpoints by removing a relatively thin horizontal layer at the base of the ore columns using standard mining methods the vertical support of the ore column above is removed and the ore then caves by gravity as broken ore is removed from the ore column the overlying ore continues to break and cave by gravity although some relatively smaller block cave ore bodies are caved and mined as a single production block most existing and planned block cave mines use either of the following an extended block caving system that divides the deposit into discrete production blocks or a single cave front or series of fronts or panels advancing forward through the ore body continually opening up new production areas as the earlier caved areas become exhausted the block caving method typically allows for relatively large volumes of production after the mine has been developed and production ramp up has been achieved the preproduction development period can be significant typically 5 10 years depending on the length of time to achieve the initial access the up front capital required prior to any return on investment is very high because much of the production levels and infrastructure must be in place before caving can begin after the mine has reached its sustained maximum production rate the operating cost tends to be very low with minimal additional infrastructure required to maintain the high production volumes block caving is generally the least expensive of all underground mining methods and can in some cases compete with open pit mining in cost the proceedings from the massmin conferences held every 4 years are outstanding sources of case histories as well as theoretical discussions of block caving many of the discussion points in this chapter are drawn at least in part from the many excellent papers in those proceedings the two latest proceedings are from 2004 karzulovic and alfaro and 2008 schunnesson and nordlund an excellent reference book on the geotechnical design aspects of block caving is block caving geomechanics brown 2003 current and planned caving mines figure 13 10 1 shows the worldwide distribution of active and planned block and panel caving mines a number of the planned mines will be developed under very large open pits that will be exhausted in the next 10 15 years thus many of the planned caving mines will be very large deposits mined at greater depths as technology improves the trend is for higher production rates in these deposits the largest caving mines in the world in 2009 were producing in the range of 75 000 t d metric tons per day from a single footprint at the same time mining complexes such as el teniente in chile were producing 130 000 t d from a number of production blocks

planned production rates from future mines are as high as 160 000 t d indicative of the evolution of caving mines into rock factories capable of these high production rates another trend in the future is for increasing mining depths as the new super caves are developed beneath the open pit mines and other undeveloped deep deposits begin to come on line as technology improves our ability to manage issues in deep mines such as in situ stress temperature gradient and more competent ore bodies this trend is expected to continue currently caving mines are being contemplated in excess of 2 000 m deep ore body characteristics as for any mining method determining the applicability of the caving method must consider many different aspects including the location type size geotechnical characteristics and value of the ore body no single formula can be readily applied to determine the suitability of an ore body to a particular caving method the caving method is typically applied to large fairly flat dipping ore bodies or porphyry type ore bodies with rock mass characteristics that are amenable to sustainable largescale rock mass failure caving the lateral extents of the ore body must be sufficient to both establish and sustain the cave the required dimensions to achieve this will vary and depend on key characteristics such as rock strengths fracture frequency and orientation fragmentation and in situ stress conditions the height of the ore columns must be sufficient to pay for the costs of developing and maintaining the production levels from which the ore is extracted at the same time the ground conditions must allow for development of the mine openings required to exploit the ore body as well as for proximal location of required infrastructure geotechnical characteristics the geotechnical character of the rock mass to be caved is a key factor in the ability to effectively initiate and propagate the cave and successfully and efficiently produce ore for life of mine an understanding of cavability fragmentation and stress are primary elements in designing and operating a caving mine rock mass characterization during the process of mine method selection a thorough understanding of the rock mass behavior is required in order to assess cavability fragmentation mine production level design mine infrastructure design and surface impacts for this purpose topics that require a thorough understanding include geology including geometry lithology and structural features of the ore body as well as the surrounding country rock hydrology including surface features and groundwater flow models geotechnical studies rock mass classifications in situ stresses and discontinuities of the rock mass faults joints bedding etc cavability the cavability of the deposit is a fundamental feature of mine design initiating the cave and then sustaining the cave throughout the mine life is critical for achievi