the aggregate of operations to open and develop a deposit and remove ores, nonmetallic minerals, and coals. A different technology is applied in underground leaching, underground solution, and other underground mining procedures that use boreholes. The opening operation is accomplished with vertical and inclined shafts or drifts. Development involves dividing a mine area into blocks, panels, columns, and other excavation segments required for extraction operations. Extraction is the essence of underground mining and involves a set of processes to remove the mineral from the ore body, transport the mineral to the loading site, and strengthen and reinforce the area that has been worked.
A plan specifying when and where to carry out development and extraction work is formulated for given mining and geological conditions. This plan basically determines the mining system that is used. A mining system must provide for operational safety, minimal losses of mineral, and consistently high technical and economic indexes. The choice of a mining system is influenced by mining and geological characteristics, such as the ore body’s thickness, angle of dip, value, structure, depth, and gas and water content. It is also influenced by the physicomechanical properties of the mineral and enclosing rock and by mine-engineering factors, such as mechanical equipment and the technological level of the enterprise.
Since mining systems for ore deposits (including mining and chemical raw material) and nonmetallic deposits (mainly gypsum) differ markedly from those for coal deposits, these systems are examined separately below.
Mining of ore and nonmetallic deposits. Ore and nonmetallic deposits are characterized by variation in the form of the ore bodies: strata, tabular deposits, stocks, lenses, and veins. The thickness of ore bodies varies from a few centimeters—the case of rare-metal and gold deposits—to hundreds of meters, as with the iron-ore deposits of the Kursk Magnetic Anomaly and the apatite deposits of the Kola Peninsula. The angle of dip of the beds varies from horizontal and gently sloping (0°-25°) to steeply inclined (45°-90°). Beds may be tens of kilometers long, as with the phosphorite deposits of the Karatau Mountains, and ore bodies may be as much as several kilometers deep. The diverse geological conditions and the physical properties of the rocks dictate the technology used to mine, transport, and ship the mineral, as well as the methods used to strengthen and reinforce the area that has been worked. In particular, geological conditions and the physical properties of the rocks influence the techniques used in the breaking operation, in which the mineral is separated from the ore body and crushed into pieces of the desired size. Explosives are used to break very hard and moderately hard rocks, whereas weak rocks are broken mechanically, by heading and extraction machines. The caving system is used in working thick deposits consisting of weak or fissured minerals; when a sufficient area is exposed, these minerals crumble, under their own weight and because of the pressure of the overlying stratum, into fragments of a size suitable for subsequent extraction operations.
The mineral broken in the excavation area is removed from drifts driven into the bottom of the blocks or from the side of horizontal excavations. The mineral may be moved by gravity, mechanical means, or explosives. Gravity transport, which makes use of the weight of the broken mineral, is carried out right at the excavation site by means of ore chutes and auxiliary installations in the excavation area, such as troughs, flooring, and pipes. Mechanized transport uses scrapers; apron, scraper, and vibrating-type conveyors; and power-driven machines that move the ore in the excavation area—in the case of gently sloping deposits—and along drifts in the foundation (bottom) of the block. Also used are units consisting of loaders and power-driven cars; for thick ore beds, transport is accomplished by short-beam excavators or bucket loaders and underground dump trucks with loading capacities of up to 40 metric tons. There are extremely efficient loading and transport machines that both load and transport ore over short distances.
The excavated area can be supported either naturally, by leaving pillars or strips of untouched ore, or artificially, with flushing materials or mine supports, such as stulls, square sets, chocks, and roof boltings. In some cases, extraction techniques control rock pressure through the caving of the enclosing rock.
There are more than 200 basic systems for the underground mining of ore deposits. Several classification schemes have been proposed by the Soviet scientists N. I. Trushkov, R. P. Ka-plunov, N. A. Starikov and V. R. Imenitov. The most common classification scheme was proposed in 1949 by M. I. Agoshkov and is based on the state of the excavation space during the extraction operation.
Opencast mining is used to work deposits of any form with stable ore and enclosing rock. In this method, the excavations are not filled with flushing materials, broken ore, or crushed rock. The roof and sides of the open excavation space are supported by permanent or temporary blocks of untouched ore.
Sharply inclined veins and tabular deposits of up to 3 m in thickness are worked by overhand and underhand stoping, more frequently the former. Lifting and haulage drifts are cut to prepare the blocks. To preserve the haulage drift when the block is being worked during overhead stoping, temporary drift pillars are left or strong timber flooring is constructed. Broken ore is removed through hatches.
In horizontal and gently sloping beds of average thickness and thicker (to 30 m), the room-and-pillar method is commonly used. This method uses regularly placed supporting ore pillars. The mineral is broken by overhand, underhand, or longwall stoping for the entire height of the room. For beds up to 15 m in thickness, an overhead incision is usually made to permit precise forming of the roof of the excavation space and simplify roof bolting. For larger beds, upper and lower incisions are made. The mineral loss in the mining pillars ranges from 15 to 25 percent, sometimes running as high as 30 or 40 percent. The room-and-pillar method is also used in the mining of potash salt strata, with the rooms sometimes several hundred meters long. Extraction is done by cutter loaders working together with reloading bunkers and power-operated cars that transport the ore to mother conveyors. The rooms are 8–12 m wide, which is equal to two or three passes of the cutter-loaders. Narrow 1- or 2-m pillars are left between passes. The chain pillars between rooms are 8–15 m wide. Up to 60 percent of the reserves remain in the pillars.
Sublevel stoping is used for thick and steeply dipping deposits. In the case of deposits that are 12–15 m thick, the rooms are arranged lengthwise in the ore body, whereas for thicker deposits the rooms are arranged transverse to the length of the ore body. The width of the pillars between rooms varies from 6 to 15 m, depending on the width of the rooms and the stability of the ore. The vertical distance between the sublevel drifts (crosscuts) is usually 10–12 m. A raise is placed in the center or to one side of the block. By extending the raise, a cut is slashed over the entire width and length of the room. The breaking is in sections, with the breaking front usually vertical. Heavy blasting crushes the block of ore and the bottom of the overlying block. Ore losses in room excavation do not exceed 2–3 percent but rise to 30–50 percent when the blocks of untouched ore and the pillars are excavated; losses for the system as a whole range from 8 to 10 percent.
Room-block systems are used to work thick deposits that are steeply dipping or inclined. Removal drifts (funnels and trenches) are placed in the basement rock to reduce mineral loss if the ore beds’ angles of dip are not sufficiently steep. The mineral in the rooms is broken in horizontal, inclined, or vertical levels. Because there are no sublevel drifts, the preparative cutting operations are reduced; breaking losses, however, increase to 10—15 percent, and depletion losses increase to 10–12 percent. The output of a miner is 12–15 m3 per shift.
Shrinkage stoping involves filling the excavated space with broken ore, which is transported after finishing the blocks. Mineral loss ranges from 5 to 15 percent.
The filled-stope system supports unstable surrounding rock with the flushing material that fills the excavated space as the mineral is extracted. In steeply dipping deposits, horizontal or inclined (at an angle of 30°–35°) cut-and-fill systems are used. The transport of ore and flushing material in the excavation space is accomplished with scrapers or power-driven machines in the case of horizontal systems, whereas inclined systems make use of gravity. The flushing material is transported through raises driven along the block edges. Ore chutes are usually installed for transporting the ore to the storage space. To reduce ore losses, the surface of the unconsolidated flushing material is covered with wooden or metal flooring or concrete before the next layer is broken. When thick and gently sloping deposits of valuable ores are being worked, variants of this system are employed that make use of consolidated flushing material and power-driven equipment. Although this type of mining is labor consuming and expensive, it is used for mining valuable ores and ores given to spontaneous combustion for a number of reasons: high extraction of reserves (ore losses less than 3–5 percent), low depletion, the possibility of mining several blocks simultaneously, and operational safety in the excavation.
Mining systems that support the excavation space make use of regularly installed props to reinforce unstable ore and enclosing rock during the excavation. These systems are usually used for mining deposits of medium thickness. The reinforced-stull system is the system most frequently used. Work is generally carried out in horizontal levels or by overhand stoping.
Deposits of weak ores that tend to spontaneously cave after even minor exposure are mined by support and filled-stope systems. These systems are used only for excavating very valuable ores because of the high cost of the mining operation and the low labor productivity. In methods that involve the caving of the surrounding rock, the excavated area is filled with caved enclosing rock immediately after excavation of the ore. Steeply dipping and thick beds with unstable ore and surrounding rock are mined by bottom-slicing methods, in which the ore is extracted in descending horizontal layers 2.3–2.5 m in height. Timber flooring is used to prevent crushed barren rock from penetrating into the useful mineral. Scrapers are used to transport the mineral. Losses with these systems are 2–5 percent. The systems are used for mining valuable ores.
Horizontal and gently sloping tabular veins up to 4–5 m in width are mined by pillar systems. The mine area is divided into pillars 25–80 m wide and from 150 to 500–700 m long. The pillars are constructed along the dip with a longwall or advances. The ore is broken by blasting; cutter-loaders are used in weak ores. The losses of ore are 7–10 percent, with losses reaching 15–20 percent in advance stoping. These systems are used in the USSR for mining deposits of manganese ores in Chiatura and Nikopol’.
Systems with caving of the ore and the enclosing rocks are characterized by mass breaking or spontaneous caving of the ore with its removal from the caved rock. These systems are used in working thick beds in stable and unstable rocks. They are used in mining 90 percent of the iron ore and 100 percent of the phosphate ore in the USSR and are also widespread in nonferrous metallurgy. Sublevel and block-caving methods are distinguished according to the stoping arrangement of the excavation. The sublevels range from 6–8 to 35–40 m in height, depending on the mine-engineering conditions; each sublevel has adits for removal and supply. The parameters chosen for the system determine which methods of explosive breaking are used. Variants of the sublevel-caving system with excavation of the ore below timber flooring are used in mining steeply dipping beds of valuable ores with a tendency to spontaneous caving.
The top-slicing system has many variants, differing in the stoping arrangements, the methods of breaking and removing the mineral, and the transport equipment used. The “closed fan” variant is used for low block heights (10–18 m). If the sublevel is more than 20 m high, the ore is broken into vertical or horizontal compensation compartments. Single-stage caving systems for the ore and enclosing rock without prior excavation of compensation compartments ensure improved technical and economic indexes for the mining operation. Another variation involves sublevel caving with vertical slicing in a compressed medium, that is, into previously broken ore or caved barren rocks. The ore is worked in sections as large as 200 m2. In sublevel-caving systems with inclined slicing, holes are drilled from the drilling or transport horizon to the subbracket space. When holes are drilled from the transport horizon, the holes 10–12 m from the mouth are not worked to prevent destruction of the scraper drifts. Removal is accomplished with two or three rows of funnels below the untouched ore block barrier, which prevents premature entry of barren rock.
There is an efficient variant of the sublevel-caving system with front removal of the ore, in which ore is transported to the ore chutes by power-operated machines. The development of the blocks involves driving sublevel drifts every 7–9 m in contact with the lower wall; crosscuts are driven at staggered intervals to allow for drilling, loading, and transport of the ore. Ore chutes are spaced about 250 m apart. Ore losses are 10–15 percent.
In induced block-caving systems, the ore is broken over the entire height of the block. The ore block undergoing simultaneous breaking may weigh hundreds of thousands of tons. The strength and stability of the ore may vary greatly. Ore is broken into horizontally cut compensation chambers and vertical compensation slots or chambers. The breaking operation in a compressed medium in slices 15–25 m thick is usually carried out on previously exploded ore adjacent to the exploded ore body; the accumulated ore is loosened by partial removal before breaking. Losses are 12–18 percent.
The spontaneous block-caving system is characterized by gradual spontaneous caving of the ore within the working section and its subsequent removal below the caved rocks. The block is from 60 to 120–150 m high and the cutting area varies from 900 to 2,500 m2, depending on the physicomechanical properties of the ore and the rock pressure. To prevent sticking of the ore at the block edges, a side cut is made by sublevel edged drifts, narrow magazines or cutting compartments, and the explosion of a fan-shaped set of holes. When blocks next to excavated segments are mined, the ore is caved in large chunks, thus complicating removal. The advantage of this system is the high output of the mine workers and the low cost of the ore mining. The system has not gained wide acceptance in the USSR, however, because of the large losses and depletion of the ore, which average 20–25 percent. A variation of spontaneous block-caving is used at the Ingulets mine in the Krivoi Rog Basin; the subdrift is 20–40 m high and the cutting area is 400–600 m2. The cut is formed by exploding blastholes drilled into the removal drifts at a depth of 4–5 m.
Thick deposits of minerals are often worked by combined mining systems, in which rooms and pillars of approximately equal size are mined simultaneously or consecutively by different systems. A unified plan is used to develop the blocks.
Outside the USSR, underground mining is widespread in Canada, the United States, Mexico, Chile, Sweden, France, the Federal Republic of Germany, Rhodesia, Zambia, the Republic of South Africa, and Australia. Many underground mines of comparatively low output are located in Italy, Spain, Japan, and the Philippines. The most frequently used systems are spontaneous block-caving, spontaneous sublevel caving, and room-and-pillar and support-and-fill systems. Integrated mechanization is employed in primary and auxiliary processes, and power-operated equipment is widely used. The diameter of the blast-holes usually does not exceed 56 mm, which ensures proper crushing of the ore and high efficiency for loading and transport equipment.
Major trends in improved underground mining include: (1) the opening of thick deposits by inclined shafts and removal of the ore to the surface by conveyors and power-operated equipment; (2) the use of spiral access ramps to transport personnel, equipment, and materials to underground excavations; (3) the use of skips holding 50 tons and more; (4) the construction of concentrated horizons with a higher opening stage; (5) the production of cutter-loaders that are capable of high-speed excavation in hard and medium rock and that use new methods for crushing the rock, and also the production of cutter-loaders and units for excavating ore of medium hardness; (6) the coordinated use of power-operated machines for mechanizing all the primary and auxiliary mining processes; (7) the increasing of the capacity and efficiency of power-operated machines; (8) the reduction of losses and ore depletion in caving systems for ore and enclosing rock; (9) the extensive use of conveyors for underground transport; and (10) the introduction of automated control systems.
REFERENCESTrushkov, N. I. Razrabotka rudnykh mestorozhdenii, vols. 1–2. Moscow, 1946–47.
Starikov, N. A. Sistemy razrabotki mestorozhdenii. Sverdlovsk-Moscow, 1947.
Agoshkov, M. I. Razrabotka rudnykh mestorozhdenii, 3rd ed. Moscow, 1954.
Gorodetskii, P. I. Razrabotka rudnykh mestorozhdenii. Moscow, 1962.
Agoshkov, M. I., and G. M. Malakhov. Podzemnaia razrabotka rudnykh mestorozhdenii. Moscow, 1966.
Kaplunov, R. P., and I. A. Cheremushentsev. Podzemnaia razrabotka rudnykh irossypnykh mestorozhdenii. Moscow, 1966.
Imenitov, V. R. Tekhnologiia, mekhanizatsiia i organizatsiia proizvodstvennykh protsessov pripodzemnoi razrabotke rudnykh mestorozhdenii. Moscow, 1973.
M. D. FUGZAN
Mining of coal deposits. There are various types of coal seams. For this reason and because of economic factors, various mining technologies are used in working coal seams. As a rule, mechanical devices and explosives are used to dislodge the coal; hydraulic and chemical techniques are used less frequently. The excavation technology assumes either the constant presence of workers in the excavation area or automated coal mining. A distinction is made between the extraction of coal by means of cutter-loaders, plows, pick hammers, and explosives. The most promising method uses cutter-loaders and plows together with mechanized supports in mechanized complexes. In 1973 in the USSR, complexes were used to mine 48 percent of the coal in gently sloping and moderately inclined seams in which loading was required. As of 1974, only limited use of complexes was possible in mining steep seams. A distinction is made between longwall and shortwall mining systems.
LONGWALL MINING. A longwall mining system may be continuous, retreating, or combined. Each system has modifications that depend on the direction in which the face advances to the components of the seam, that is, in direction, dip, and rise. There are modifications dependent on the way the level or stage is developed for extraction. When thick seams are worked, there are variations arising from the method of extraction relative to thickness: systems with and without slicing, and slices that may be inclined, horizontal, or transversely inclined.
In the continuous mining system, gateways are constructed at the same time that coal is extracted in a wing of the level or panel. The working face is developed not less than 25–50 m from the inclined drift—gradient, ramp, conduit shaft—or horizontal drift by constructing transport and ventilation drifts and an opening-out raise between them. Mechanical equipment is installed in the opening-out raise and the coal-mining operation begins; the working face advances from an inclined or horizontal drift toward the edge of the level or panel. After the stoping operation, excavation of the area adjacent to the face begins. This arrangement of the faces of the excavation drifts and gateways is retained during the entire working of the level or deck. Other variants of this system depend on the angle of dip of the seams and are classified according to the way the seam is developed and how excavations are made. In the continuous system, the excavations used in developing a new working face are at first small in size. The continuous system’s major disadvantages include the need for complicated systems to maintain the drifts and the considerable air leakage through the excavated space. In addition, unexpected geological irregularities may be encountered and longwalls may become necessary. The continuous system hinders the use of highly efficient complexes and units. The use of this system should therefore be limited to thin seams at great depths and to isolated and unshielded seams that pose a threat of sudden coal and gas blowouts and mine shocks.
In pillar systems, gateways are extended before excavation work is begun. These systems outline the coal reserves within a level, deck, or excavation pillar.
One variation of the pillar system uses the panel method of developing the mine. A receiving and dispatching platform is installed near the main haulageway at inclined drifts so that loads can be received and transported from the shaft area to the working face and back. Inclined drifts—inclines, ramps, and passages—are constructed from the receiving and dispatching platform to the top or bottom of the panel to provide for air supply, auxiliary transport, and the movement of workers. The coal is transported by conveyor belts along the incline or ramp. Deck drifts for transport and ventilation extend in both directions from the inclined drifts; these deck drifts have inlets, connectors, and other auxiliary excavations. As one area is mined out, the next deck is prepared and new drifts are constructed. The pillar system eliminates the disadvantages related to the continuous system, although coal losses in the untouched seams are higher by 5–7 percent and the initial volume of the gateways is greater. The pillar system permits an increased load on the working face and an improvement in the major technical and economic indexes. The system is used widely in mining thin and medium seams and in slicing systems for thick seams.
Increasing use is being made of long-pillar up-face excavation systems and long-pillar down-face excavation systems. The gangway is extended from the opening drift. A level drift and two inclined shafts are driven parallel to the gangway up to the ventilation horizon, where they connect with the opening-out raise. The excavation pillar may be 1,000–1,500 m long or longer, with the width corresponding to the length of the long wall. The next pillar is prepared by cutting new inclined shafts and an opening-out raise. The up-face system permits a reduction in the specific volume of the passage and supporting drifts. It also allows the length of the longwall within the excavation pillar to remain constant, which is especially important in mechanizing the excavation operation. The system ensures simple and reliable underground transport and direct ventilation from the air inlet to the sources of methane release, such as the working face, excavated space, coal on the conveyor, and development excavations. Disadvantages of the system include the large volume of the inclined shafts, which are more expensive to construct and maintain than horizontal excavations. In cases of high water content, a similar long-pillar system is used with a down-face arrangement. Because of their technical and economic advantages, the two variations of this mining system are the most advanced methods for mining thin and medium seams with an angle of dip to 12°–15°.
Shortwall or up-face long-pillar systems are also used in excavating thick and gently sloping seams. The shortwall long-pillar system is most often used in working thin and medium inclined and steep seams. The mining method is a decisive factor in determining the dimensions of the excavation area and the length of the working face. When coal is mined by drilling and blasting, the length of the excavation area does not exceed 300-400 m, but with mechanized excavation the length may exceed 1,000 m. Each excavation area is opened by connecting crosscuts, from which the haulageway (conveyor) and ventilation drifts are cut through the seam. The up-face long- pillar system with shield support is used for mining steep seams with the aid of movable ceiling supports. The system was first proposed in the USSR by N. A. Chinakal and was first used in the Kuznetsk Basin in 1938. The level is 80–100 m high and is divided into excavation areas 250–300 m wide. The excavation areas, in turn, are divided into individual support pillars. The length of the working face and the method of the pillar development depend on the technology used to mine the coal. In the drilling and blasting method, the working face is no longer than 24–30 m; coal chutes are placed every 6 m under protective shields. Among the system’s disadvantages are high operating losses, the large volume of the development work, the low degree of mechanization, the great amount of manual work, and the high accident rate. The system thus does not hold much promise.
In mechanized coal breaking with shield units, coal transport is accomplished through side holes. The working face may be as long as 55 m.
In the shield system, the wall rock caves spontaneously when the up-face support is released. This system is used only for seams with angles of dip greater than 55°.
In pillar flat-back systems for thick seams, the seams are divided into inclined, horizontal, and transversely-inclined levels by arbitrary planes inclined at the dip of the seam, parallel to the ground or roof, horizontally situated between the hanging and lower walls, and with a 30°–40° angle toward the ground. The slice is not more than 3.5 m thick.
The system with upward horizontal strips is used to work seams 3.0–4.5 m thick with angles of dip greater than 60°; the seams are worked with hydraulic flushing of the excavated area and coal getting with cutter-loaders. Thicker seams may be worked in slices not more than 4.5 m thick. A two-winged excavation area extending 300–400 m is opened on the working and ventilation levels by intermediate crosscuts. Ventilation chutes are placed at the margins of the excavation area. The coal chute is cut in the middle of the excavation area in the excavated and filled space as the face advances. In order to combine the operations of extracting coal and raising the flushing material, the left and right excavation strips alternately outdistance one another by a distance equal to half the height of the extracted strip (4.5–5.0 m). The coal is extracted by one or two cutter-loaders in opposite wings of the excavation area. The coal is transported from the cutter-loaders to the coal chute by conveyors. After extraction is completed in a wing, the cutter-loader is moved to the adjacent wing along a transfer truss constructed over the coal chute, and hydraulic flushing is added from a slurrier placed along a side chute.
The rill system for thick and steep folds is used only when roof control is provided by flushing the excavated space. The coal is extracted in the slices by drilling and blasting.
In the rill system for thick and gently sloping seams, the seam is divided into two or more slices. A haulageway is driven from the inclined shafts to the edge of the mine area (panel) for this operation. The haulageway of the excavated upper level is used as the ventilation drift. An opening-out hole and conveyor and ventilation slices are extended to the edge of the mine area (panel) along the highest level. The conveyor slice is connected to the haulageway, and the ventilation slice is connected to the level ventilation drift. The working face is developed similarly along the lower slice. The excavation operations are carried out as the face of the upper layer advances relative to the lower layer. The magnitude of the advance depends on the way the slices are worked. Simultaneous working of slices with small advances between the layers (to 100 m) is practiced, as well as consecutive working with independent development of each slice. Blocks of coal 0.3–0.6 m thick are usually left between the slices; less frequently, flexible metal band or mesh canopies are used. The slices are excavated by the shortwall or up-wall (at angles to 12°–15°) long-pillar system.
The rill system for thick steep and inclined seams is used with roof control accomplished by caving and flushing. When the roof is caved, the layers are worked in descending order by drilling and blasting methods under flexible metal canopies. This system is used for seams over 4.5 m thick.
When roof control is accomplished by flushing the excavated space, the inclined slices are worked in ascending order; there are no more than four slices, and each slice is no more than 3.5 m thick. The inclined slices are extracted by drilling and blasting or by mechanized means. When the drilling and blasting method is used, the width of the excavation area does not exceed 400 m, the slices are worked in strips by shortwall breaking, and the working is no longer than 12 m. When mechanized means are used, the inclined slices may be developed by either the shortwall or down-face long-pillar method. The working face is 30–200 m long, the excavation areas are 400–1,200 m long, and the extracted slice is 2.5–3.5 m thick. The technology of coal mining with complexes involves an increase in the vertical height of the level to 200–250 m and the use in the slices of reinforced flushing materials with higher load-bearing capacity; this flushing material ensures safe operation of the mechanized supports in the subsequent slices without the use of additional canopies.
The mines of the Kuznetsk Coal Basin were the first to use the combined rill system for thick and gently sloping seams with a special coal complex for the transport of strata between slices. The system is designed for seams 7–12 m thick and with low gas content. The seam is divided into two slices, each of which is developed independently. The upper, or installation, slice is 1.5–2.0m thick and is developed by a shortwall long-pillar system. A flexible metal canopy is installed as the coal is mined; the roof rock is caved onto this canopy. The lower slice is developed by the up-face long-pillar system. The pillars are 300–500 m long, and the working face is 40–80 m long. Coal is extracted in the slice at the height of the support (2.8 m) by a cutter-loader and in the strata between slices by drilling and blasting. The broken coal of the statum between slices is released onto a face conveyór through hatches in the support enclosure.
SHORTWALL MINING. Shortwall systems are divided into room and room-and-pillar systems. In room systems, the rooms may be 200–300 m long and 4–15 m wide. The pillars between rooms may be 2–6 m long, with the sectional pillars 5–10 m long. The dimensions of the excavation section are such that the caving of the roof takes place after the section is developed. For gently sloping seams, the section is 50–150 m long.
The room-and-pillar system differs from the room system in that the pillars between rooms are partially removed, thus resulting in a higher degree of coal extraction. One or two rooms approximately 3.5–5 m wide are developed between the conveyor and ventilation drifts, after which the pillar between rooms is removed; the pillar is 15–20 m wide. The pillar between rooms is removed between the passes of 3.5–7.0 m; structural pillars of 0.6–1 m are left. The drifts and rooms are supported by an anchor support; the passes are not supported. Shortwall systems are used when the coal to be mined is of poor quality—usually furnace coal with a high ash content. The system can be used for seams 0.8–3.5 m thick, with an angle of dip of as much as 15°; this permits the use of power-operated equipment. The surrounding rock must have high or medium firmness, and the gas content may be up to 15 m3 per ton of product. The depth of the operation may reach 300 m; increases in depth lead to a rapid rise in coal losses.
The relative importance of the various systems of coal mining in the USSR is shown in Table 1.
|Table 1. Relative importance of mining systems in the USSR, 1973 (in percent)|
|Longwall systems||Shortwall systems|
|Basins||Without slicing||Inclined slicing||Other1|
|1Includes horizontal slicing and combined 212.4 percent with shields 32.2 percent with shields|
|Overall USSR ..||14.5||63.53||7.4||13.8||0.8|
Outside the USSR, underground coal mining is highly developed in the USA, the Polish People’s Republic, Great Britain, the Federal Republic of Germany, and France. Longwall systems are the systems most used in Europe. Shortwall systems are used in the USA, Canada, and Australia because of favorable geological conditions.
The major tasks of the coal industry in the USSR are further concentration and intensification of mining operations. This goal is being achieved by (1) expanded use of long-pillar systems and especially those variations providing constant wall length and separate ventilation of methane sources; (2) efficient integration of development drifts in the coal stratum and the surrounding rock; (3) prediction of geological irregularities in order to ensure stable operation of complexes and units; (4) development of new modifications of mining systems and highly efficient integrated mechanical equipment to eliminate the need for mine workers at the working face; (5) development of new and improved mining systems for thick and steep seams with the use of flushing, especially hydraulic flushing; (6) development of integrated systems for conducting mining operations in deep horizons with prior degassing of the seams; (7) surface control of the rock tract before mining operations are begun so as to eliminate gas and coal blowouts and mine shocks; and (8) development of means to ensure comfortable and safe working conditions.
In 1973, underground mining accounted for 71 percent of all the coal mined in the USSR.
REFERENCESSheviakov, L. D. Razrabotka mestorozhdenii poleznykh iskopaemykh, 4th ed. Moscow, 1963.
Tekhnologiia podzemnoi razrabotki plastovykh mestorozhdenii poleznykh iskopaemykh. Moscow, 1969.
Kiliachkov, A. P. Tekhnologiia gornogo proizvodstva. Moscow, 1971.
Tekhnologicheskie skhemy ochistnykh ipodgotovitel’nykh rabot na ugol’-nykh shakhtakh, parts 1–3. Moscow, 1971–72.
Tekhnologiia podzemnoi razrabotki plastovykh mestorozhdenii. Moscow, 1972.
B.F. BRATCHENKO and A. P. KILIACHKOV