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Jan 1, 1967

Jan 1, 1969

Jan 1, 1970

By E. W. Cokayne

GENERAL DESCRIPTION OF THE METHOD The original application of sublevel caving was in ground so weak that it would collapse even in small headings when the support was removed. Heavily tim¬bered drifts were driven across the ore body; the timber was removed at the end of the drift, and the ore caved and slushed out. When dilution became excessive the next set was removed and so on. This method gave high dilution and poor recovery and was slow, but it was the only way to mine this type of ore body at that time. In more recent times the method has been adapted to stronger ground which needs to be drilled and blasted, so it is not a caving method as far as the ore is con¬cerned. However, the method does rely on the walls to cave and so the name sublevel caving has been retained. The method as presently generally applied is shown in Figs. 1, 2, and 3. Sublevels are established at approximately 7.6 to 12.2-m (25 to 40-ft) vertical intervals. They are usually accessed from a ramp system. A haulage drift is driven down the strike of the ore body, but in the waste. Production drifts are turned off the haulage drift and driven across the ore body on about 10.7-m (35-ft) horizontal centers. This divides the ore body into a geometrical pattern. Slot raises are driven at the end of the drifts near the ore-waste contact and the slot is expanded to the shape of the fan drilling. Up holes are drilled to break towards the slot at burdens of 1.2 to 1.8 m (4 to 6 ft). They are drilled past the next sublevel to the production drift im¬mediately above. When a sufficient number of fans has been drilled, production can start by blasting one or two fans at a time towards the slot. Load-haul-dump (LHD) units are used to muck the ore, which rills out into the drift. At first, when only a small area is mined, the walls may not cave, and the operation will be similar to open stoping. Later, sloughing from the walls will fill any voids created, and the ore after blasting will be con¬tained by the caved waste and gravitate downwards as mucking proceeds. This process is repeated in each of the production drifts as the face is retreated towards the haulage drift. When the top level has been retreated back a safe dis

Jan 1, 1982

By P A. Jenkins, I W. O Jones

WMC has mined the Perseverance disseminated orebody exclusively by SLC (sublevel cave) methods. Deteriorating ground conditions in the vicinity of an orebody inflection forced a re-evaluation of future mining at Perseverance and a strategy to drop down below the inflection zone and recommence the cave was implemented. This strategy has led to increased production at higher grades, resulting in lower costs and better mill recoveries. This paper describes the evolution of mine planning, development, support and production strategies through the successful process of implementing the ædrop-downÆ and re-establishing the SLC.

Jan 1, 2000

By P Hucker, E Hall

Harmony Gold operates the Mount Marion underground gold mine near Kalgoorlie, Western Australia. The mine consists of two ore zones, East and West, and mining of these zones is currently conducted using the sublevel caving method. When mining commenced in the Western zone, the span was insufficient to induce caving of the country rock above. Dyno Nobel, together with Harmony and 360 Drilling, developed an innovative drill and blast solution in order to promote cave propagation and improve the safety of personnel at the operation. The solution involved a series of shallow angle drill holes up to 90 m into the hanging wall of the West zone. These were charged with Titan« 7000 underground bulk emulsion and the HotShotÖ electronic initiation system. The blast used over seven tonnes of emulsion and 46 detonators and achieved the desired result without damaging nearby mine infrastructure.

Jan 1, 2007

By Lockyer PC, Reed MR Bartlett PM

Difficult ground conditions have plagued the underground operations at Mount Windarra since mining started in 1974.The ore at Mount Windarra consists of seven vertically dipping, sub vertically plunging ore shoots. The ore zone in which the shoots lie is characterised by moderate to heavy cleavage and the presence of serpentanite. The ore is known to extend to 1100 metres below surface.Long hole open stoping under cover of backfill has been used as the primary mining method at the mine, but its poor performance brought the mine near to closure in 1984. The mine had been closed from 1979 to 1980.

Jan 1, 1985

By Ingvar Janelid

During recent years, numerous articles have been published regarding theories and practical applications of sublevel caving. The reason is the great interest this method has created through its possibilities of standardizing and mechanizing operations and equipment. Other advantages are that the method can be adapted to orebodies of different shape, size, strength etc., and especially that working conditions can be made safe and relatively comfortable since all work is carried out in drifts. As is well known, the disadvantages or sublevel caving are that ore recovery hardly exceeds 85-90 % if waste rock dilution is to be kept below 20-30 %. Therefore, the main research and development work has been concentrated on parameters which can affect and reduce these disadvantages. This paper deals only with principles and problems which apply to large, massive orebodies. The problems with narrow, vertical or inclined orebodies are so different that they require a special handling. In my opinion, the intense research and development work during recent years, combined with practical experience, has given enough knowledge of the influence of different parameters, so that normal sublevel caving can be efficiently planned.

Jan 1, 1975

Much of the theory upon which the Sublevel Caving (SLC) method is based was developed in Scandinavia many years ago. It was based mainly on æbinÆ theory and æellipsoid of drawÆ configurations derived from sandbox models. In many cases, classical SLC layouts influenced by these early theories obtained relatively poor results. Early dilution entry, low tonnage factors and low recoveries were commonly experienced. SLC soon gained the reputation of being a high dilution, low recovery and development intensive method. This led to SLC falling into disfavour generally, other than as a pillar reclamation method used at the end of the life of mines. More recently a number of mines, notably those in Australia, have adopted SLC as a primary extraction method and have achieved good recovery results. The reason for the better results is believed to be an outcome of modifications in some of the basic design parameters and operational procedures. This has led to questions being raised regarding the original æbin theoryÆ assumptions and classical SLC layout designs. It is clear that the fragmentation and flow characteristics occurring in a choke blast situation (compaction of waste, variable size and shape of fragmentation throughout the ring) conflicts significantly with the assumptions of regular loose flow of uniform, fine grained, material in a bin. Interactive versus independent draw conditions can also materially effect the outcome of a SLC operation. The good results obtained more recently can only be explained by using ænewÆ models of behaviour, which have to a large extent been borrowed from block caving. If these models are correct, they point to a number of design and operational changes that can be applied to improve the effectiveness of SLC. This paper discusses the factors that may not have been fully considered in early SLC design and modelling. It emphasises the importance of delayed dilution entry as a primary objective of sound SLC design and provides a basis for efficient SLC design and operation.

Jan 1, 2000

By J F. Carswell, G Cheshire

The Pajingo gold operation in Queensland is approximately 55 km south of Charters Towers and 134 km south-west of Townsville. The gold deposit was discovered on the northern end of the Drummond Basin in 1983 and first open pit production began in 1986. Mining progressed underground in the 1990’s and gold production peaked in the early 2000’s as the main Vera/Nancy lodes were extracted, producing 330k oz in 2003 (a record at the time). More recently, as Ore Reserves and grade have diminished, the mine has had to develop innovative mining practices to maximise recovery of the remaining Mineral Resource. This has included a modified sublevel caving method against previously mined/backfilled zones. The method has been coined modified sublevel shrinkage (MSLS). Full scale trials of the new mining method have been successfully completed. This has allowed the site to revaluate the Mineral Resource and include areas into the Ore Reserve that were once deemed sterilised, adding significant value to the operation and extending its life-of-mine. This paper describes the first of two large scale MSLS trials and key operational learnings from it.

Mar 15, 2021

By J. A. Cox, R. S. Mattson

"One of the main problems encountered in employing sublevel caving is the inherent high dilution factor. The operating results at Granduc illustrate that this dilution can be closely controlled and satisfactory recoveries achieved, in base metal ores grading less than 2.0% copper equivalent, by using an assay grade control system based on the fundamental draw characteristics. The history of draw control procedures employed since the start of operations provides an effective comparison between this system and one based mainly on tonnage. The transverse and longitudinal sublevel caving mining methods are described and it is shown how improvements in design layout, operating practice and draw control in narrow complex ore zones (12-foot minimum width) have raised the operating performance of the longitudinal system to a satisfactory level."

Jan 1, 1973

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Ralph S. Mattson, Frederick T. Hancock

GENERAL DESCRIPTION The Granduc mine is situated on the west side of Granduc Mountain in the Leduc River area of north¬west British Columbia about 51 km (32 miles) north¬east of the town of Stewart. Entry is by means of a 16.6-km (10.3-mile) long haulage adit from the concentrator site at Tide Lake. The principal ore mineral is chalcopyrite. Ore mineralization varies from dis¬seminated to irregular semimassive stringers generally associated with highly folded and sheared biotite-rich sections of metasediments. The ore bodies vary in width from 3.6 m (12 ft) to over 18 m (60 ft) in a zone some 800 m (2600 ft) in length dipping at an average of 75° to the west and plunging steeply to the south. Capping varies in thickness from 0 to over 300 m (1000 ft). SELECTION OF MINING METHOD In the early stages of planning the primary require¬ments to be met by the mining method were: (1) suit¬ability to high output and productivity, (2) adaptability to mechanization in order to minimize operating costs, and (3) minimization of the costs and the development period for preproduction. Considering these, the selection of a mining method was narrowed to a choice be¬tween sublevel long-hole stoping or sublevel caving. A method utilizing backfill was ruled out as uneconomical, especially since mill tailings would not be readily available. At the time, the following facts about the ore body were recognized: (1) the average width of the ore body was 12 m (40 ft), (2) 30% of the stoping blocks were in areas of 6 m (20 ft) or less in width; (3) in many cases, ore zones were located in parallel lenses or were irregular in shape; (4) major faults were present in the hanging wall close to the ore, with minor faults within the different ore zones; and (5) the ore itself was lami¬nated with fractures and jointing crossing such bedding planes. Considering these factors, the two stoping meth¬ods were compared. Sublevel Open Stoping The characteristics of sublevel open stoping are: 1) It is generally not used for widths below 6 m (20 ft) since the ratio of waste development increases as level intervals are reduced to insure good control of long-hole drilling. 2) This method is used where ground is relatively competent. Otherwise more pillar support is required, increasing ore losses or mining costs with pillar recovery. 3) The method is not easily adaptable to parallel ore zones separated by relatively narrow waste bands. 4) Close control of pillar and stope design will per¬mit some degree of dilution control. Sublevel Caving Sublevel caving is characterized by the following: 1) The method can be used for a variety of ore widths, recognizing that 3 m (10 ft) in width may be an economical minimum. 2) It is particularly favorable where wall rock is weak. 3) The method has flexibility so that irregular ore bodies or parallel lenses can be mined. 4) The need to leave pillars for support is eliminated 5) Dilution and tonnage recovery are interrelated. With concern about the stability of the hanging wall and the geometry of the ore bodies, there was a growing favor to choose the sublevel caving method. However, this method was relatively new to Canadian mining practice. Teams were sent to Sweden to tour mines using sublevel caving and to study their methods and any extraction and dilution problems. The objective was to develop cost and performance data and ore recovery grades and tonnages for feasibility comparisons. Next, detailed mine plans were developed covering a three-year period using sublevel caving and sublevel open stoping. Extensive analyses of the development and production costs for each method were made. With sublevel caving, ore losses of 10% with a 20% dilution rate were used. The same ore loss was estimated for sublevel open stoping with dilution varying from 11 to 15%. As a result of these studies, the decision was made to use sublevel caving partly because slightly lower op¬erating costs, together with higher productivity, offset the adverse effects of higher dilution; and partly be¬cause, at the start, there would be less risk in being able to handle the ground, and the method offered more flexi¬bility. Open stoping was not ruled out for testing at a later date. It was recognized that some other method of mining would be required for mining the ore extensions below the glacier, where surface subsidence must be avoided. During the operating life of the mine, open stoping was tested in narrow width ore zones and found un¬satisfactory due to wall failure. A cut-and-fill stope, utilizing run-of-mine waste from development, was also operated on a trial basis. A sequence was finally estab¬lished that permitted ground control; however, produc¬tivity needed considerable improvement in order for open stoping to be considered as a sole alternative for mining below the main haulage level. At the termination of operations in June 1978 over 13 000 000 t (14,500,000 st) had been extracted from above the main haulage level at production rates varying from 3270 to 7250 t/d (4000 to 8000 stpd). Sublevel Caving Methods Two basic methods were used, transverse and longi¬tudinal. The transverse method was used mainly in the upper C ore body in widths greater than 18 m (60 ft), and occasionally as an alternative to the longitudinal method in localized areas where strike faulting inter¬fered with normal development. The longitudinal method was used in ore bodies varying in width from 3.6 to 18 m (12 to 60 ft), with a multilongitudinal varia

Jan 1, 1982

By Roger Malmström, H. Heden, Kjell Lidin

INTRODUCTION LKAB was formed at the turn of the century when the railways between the ore fields and harbors were built. The mines are situated north of the Arctic Circle in Sweden. Kiirunavaara is the largest mine, with a pro¬duction capacity of approximately 30 Mt/a. Malm¬berget has a capacity of about 12 Mt/ a. Both are under¬ground mines. Svappavaara, the only opencast mine currently in operation, has a capacity of about 4.5 Mt/a. The ore harbors are at Narvik for the Kiruna and Svappavaara mines and at Lulea for the Malmberget mine (see Fig. 1). Production in the form of finished products is shown in Fig. 2a, b, c, and d. Due to poor ore markets in re¬cent years, production and shipments have been cut back, but there was a positive swing in 1978. LKAB has been 96% state-owned since 1957. In addition to the iron mines in northern Sweden, LKAB has subsidiaries in central Sweden which produce sulfide ores, quartz sand, and uranium (Fig. 3). The mining methods employed are opencast bench stoping at Svappavaara and sublevel caving at Kiirunavaara and Malmberget. GEOLOGY The Kiirunavaara ore body is 4000 m long and av¬erages about 90 m wide, with variations in width from a few meters to around 150 m. The ore has been ex¬plored by deep drilling down to about 1000 m, but mag¬netic measurements indicate a depth of at least 1500 m. The Kiirunavaara ore is oriented in a north-northeasterly direction with a number of ores in the boundaries be¬tween the wall rock types. The ore in Kiirunavaara is mainly a fine-grained magnetite with a varying content of fine-grained apatite. The apatite content of the ore decreases with increasing depth. The other ores are all apatite-bearing and a couple are hematite ores (Fig. 4a, b). The ores are sedimentary, tilted, and tectonically de¬formed. They have a lateral dip of 0.87 to 1.2 rad (50 to 70°). The contacts between ore and wall rock are distinct and sharp, but both the hanging wall and the footwall are brecciated and impregnated with magnetite. The footwall of the Kiirunavaara ore consists of syenite porphyry with a low quartz content and rela¬tively good strength; the hanging wall consists of quartz porphyry. Both types are known as Kiruna porphyries. The ore at the boundary with the hanging wall is often rich in kaolin. There are a few veins of syenite porphyry running through the ore, which has a few major faults (Fig. 4c, d). There are two sharply delineated ore grades in Kiirunavaara: low phosphorus ore with less than 0.1 % phosphorus (about one-third of the ore) and high phos¬phorus ore with greater than 0.5% phosphorus. The iron content is 67% in the low phosphorus ore and 60% in the high phosphorus ore. These ore types are sold as different grades and so are kept separate during mining. MINING AT KIIRUNAVAARA The Kiirunavaara and Luossavaara ores were first mined using opencast bench stoping in 1902. When mining costs rose owing to higher waste removal, under¬ground methods were adopted. In the mid-1950s, the ores were still being mined on the surface at the same

Jan 1, 1982

By Larry J. Tucker

INTRODUCTION The orebody at Pea Ridge strikes NE-SW and dips at about 800 to the SE. The ore and surrounding rocks are Pre-Cambrian in age. These rocks are usually very competent. Locally strong jointing has caused problems and toward the hangingwall dikes have caused particular problems during or after the open stoping phase of production. About 396 m of sediments lie over the ore. The shape and mineral content of the orebody has been fairly consistent so far with production areas lying from the 1375 to the 2475 levels. The deeper limits of the ore- body haven't been found yet. The ore is high grade magnetite with a strike length of over 600 m and a maximum width of over 180 m on some levels at the center of the orebody. Production at Pea Ridge for the first three years of operation came from two types of open stopes. The original stopes were recovered with horizontal production rings being drilled and blasted from vertical raises. Later a more typical sublevel stoping method using vertical production rings from one or more drilling sub- levels was used. There were a number of reasons for the selection of open stoping as the initial production method at Pea Ridge. First, it was judged that the ore was not of a block cave quality. After fifteen years of production, this assumption can be verified. There was some concern during the planning days for the potential for high flow rates of water into the mine from aquifers in the sediments lying over the orebody. Dewatering wells from surface and underground were put in as a precaution and nothing came of this potential problem. Open stoping and its resulting pillar structure put off this problem by supporting the sediments a few years longer. For a moderate investment in level development some production from open stoping was quickly realized. The first open stopes were brought in on what was then the bottom mine production level (2275), near the shaft pillar and situated on the footwall. This production practice of stoping on the heels of development from shaft pillar westward and from footwall to hangingwall continued through the open stope phases of production at Pea Ridge for many years. From the standpoint of mine efficiency and pillar stability, this would not seem to be a recommended mining practice. 74 570 w electric slushers, using 1.37 m dippers, pulled the open stope ore out into 18 ton bottom dump cars. Ordinarily five of these cars made up a train, which was pulled to the dump area by a trolley powered electric locomotive. Open stoping had the disadvantage that 100% of the recoverable portions of the orebody had to be developed in order to recover 33% of the ore. 67% of the ore was left behind in a large pillar system. For many years we had no planned method for recovery of these pillars. Mining methods are of course, limited by the existing technology. Advancing technology, coupled with the limited production capabilities of open stoping, low efficiencies and high relative cost early in the history of the operation recommended that other mining methods be looked at. In Sweden, considerable practical experience, coupled with a large amount of theoretical work had become available on sublevel caving. To Pea Ridge it offered the apparent benefits of: 1) The flexibility of diesel powered trackless haulage. 2) The efficiency of two machines to one operator on longhole drilling. 3) A multilevel production picture with a large number of drawpoints for ore recovery. 4) 100% recovery of the mineable portion of the orebody in one production pass, with no pillar recovery problems left behind. This paper outlines the experiments and experiences which Pea Ridge has gone through with sublevel caving from 1967 to this date.

Jan 1, 1981

By G. W. Johnston

INTRODUCTION For many years mining at the Stobie mine of the Inco Metals Co. consisted mainly of fill-type mining for higher grade ores and block caving or blasthole open stope mining for lower grade ores. Open stope mining, although quite favorable in the initial stoping phase, has always presented problems in the pillar recovery phase, especially at depth. In 1966, sublevel caving was introduced. This method is ideally suited for low-grade ore and can be applied very successfully to intermediate or higher grade ores providing proper controls are enforced. Sublevel caving has three distinct advantages: it is highly adaptable to mechanization; it recovers the ore in one operation; and it provides adequate ground sup¬port and safety regardless of mining depth. This method does, however, make extra demands, particularly for new types of supporting services. The constant presence of rock makes quality control of the grade of the ore being extracted essential. Increased de¬velopment, different types of ventilation, roadways, and maintenance are only a few of the many changes re¬quired. The care and maintenance of the fleets of diesel equipment and the modification of equipment is a con¬stant challenge. CHOICE OF SUBLEVEL CAVING The question arises, why mine by sublevel caving? Fig. 1 shows a typical problem encountered by the open stope or blasthole mining method. Near the surface there is little or no ground support problem. As mining progresses to an intermediate depth, however, ground failure starts to occur in the supporting pillars. If open mining is carried beyond this point, the critical depth is soon reached, accompanied by hanging wall caving, ex¬tensive pillar failure, and the danger of losing control of the mining cycle. The rock caved from the hanging wall mixes with and dilutes the ore being extracted from the blasthole stopes. Ore losses occur when pillar failure pre¬vents drilling and development of part of the ore. The interruption to the mining cycle causes serious curtail¬ment in production until ground failure is controlled and the mining method changed to adapt to the new conditions. Having recognized the limitations of open stope min¬ing, alternatives were sought. The first alternative (Fig. 2) was to fill the open stopes with sand fill be¬fore extensive pillar deterioration occurred, recover the pillar ore by fill mining, and then convert to total fill mining below. If the ore grade is high enough it can support the costly fill method. Recovery will be at a much slower rate and will therefore extend over a longer period of time. On the other hand, sublevel caving gives complete primary mining recovery in a short time and is very adaptable to mechanization, which is now a necessity to combat spiralling labor costs and give greater produc¬tivity per man. Sublevel caving is ideally suited as a mining method for low-grade ore bodies which eventually reach a criti¬cal depth for open stope mining. DEVELOPMENT The development phase comes first and of course its main purpose is to gain access to the ore and provide a route for ore removal from the mine. Initially a 20% inclined service ramp is driven in the footwall rock (Figs. 3, 4, and 5). It is preferable to start at the top of a mining block and drive the ramp down¬ grade so gravity can assist in mucking the blasted de¬velopment round, and also access can be gained more quickly to the initial production area. The ramp nor¬mally turns at 3.1 rad (180°) about every 152.4 m (500 ft) to keep it reasonably close to the rock disposal

Jan 1, 1982

By H. Nan

Using the case of Qianqiu Mine, Yima Coal Mining Administration, theoretical analysis, laboratory physical simulation, and underground observation were performed to study the major parameters for fully mechanized sublevel caving mining method, including drawing spacing, drawing sequence, drawing numbers, and number of shields performing simultaneous drawing. The effects of those parameters on coal recovery and amount of waste rocks were also investigated. The optimal coal drawing technology for this case was obtained.

Jan 1, 2006

By D. T. McMurray

INTRODUCTION Shabanie mine, situated some 180 km east of Bula¬wayo, has been a producer of chrysotile asbestos for more than 50 years. The ore bodies occur in serpentinized dunite, which overlies talc-carbonate schist. A zone of relatively competent rock of varying thickness occurs between the schist and the ore bodies, which are gen¬erally less competent. The hanging wall of the ore bodies is economic, and the hanging-wall serpentine carries a variable subeconomic amount of fiber. It is important to note that, in general, the ore body competence is less than that of the foot and hanging-wall formations. Historical After surface operations ceased, cut-and-fill stoping was used to win ore from underground; this was success¬ful until the increasingly stoped-out area caused insta¬bility in the stope pillars and back. Consequently, dur¬ing the early 1950s, a gradual change to cave-mining methods was made, the ore being won by hand lashing in drawpoints, situated in the basement of the stope blocks, and passed through orepasses under gravity to the haulage level some 13 m below. About this time, interest was focused on the sub¬level caving method in use in Swedish iron ore mines: it was felt that it might be applied economically to the Shabanie ore bodies. Accordingly, in 1958, an experi¬mental stope block was laid out in which sublevel inter¬vals and extraction tunnel spacing were 9 m. The tun¬nels (ring drilling drives) were oriented on strike-in contrast to the Swedish system, in which crosscuts that retreat from hanging to footwall are used. The advantages of the method were quickly appre¬ciated by the operating personnel and, despite the in¬evitable teething troubles pertaining to the introduction of any new mining method, it was not long before sub¬level caving was providing a high proportion of the mill feed. The disadvantages also became apparent at an early stage, however, and, from that time to the present, continuing modifications have been made to mining lay¬outs in an effort to improve ore recovery. GENERAL DESCRIPTION OF METHOD The mine is served by a vertical hoisting shaft, in which two skips, a man cage and a service cage, provide adequate capacity for production requirements. The rock hoist is a Ward Leonard control hoist, in which two electric motors drive a common gearbox. The man winder is driven by an a-c motor. Several auxiliary shafts provide secondary egress and intake and return ventilation. Main haulage levels are above (Fig. I a and b). Blasthole fan patterns are drilled by drifters of 100 mm bore, drilling 41-mm holes; when a sufficient strike length has been drilled, a slot is cut in the upper¬most sublevel and the rings are broken into the slot. Initially, a limited tonnage is drawn, since it is essential to ensure that the hanging wall caves behind the retreat¬ing stope face. Once this has been established, maxi¬mum tonnage can be drawn, as described later in this chapter, under the heading "Draw Control." The broken rock is loaded by 0.14 and 0.20-m3 load¬ers into cocopans (rocker-dumping type of tipping truck), which are hand trammed to orepasses, discharg¬ing on the haulage where 11-t electric trolley locomo¬tives haul 3.95-m3 Granby cars to the main shaft bins. As is evident from Fig. 1, the layout is simple, the block is brought rapidly into production, there is a high degree of selectivity and flexibility, and the result is a low-cost high-productivity mining method. DEVELOPMENT Main haulages are developed at 3.2 x 3.2 m, and once the service winze connections have been completed the development of the sublevels is undertaken. The footwall drives are cut first, to obtain access to the block. These ends are of the standard section, 2.4 x 2.8 m, and from them crosscuts at intervals of 70 m are driven through the ore body to the hanging wall. These crosscuts are used to supplement the geo¬logical information previously obtained from diamond core drilling, and they provide additional and more de¬tailed data on fiber percentages and lengths, structural features, and other relevant criteria which are used to build up the geological assessment of the area and to classify it in terms of the geomechanics rock classification (Laubscher and Taylor, 1977). The crosscuts also allow the necessary orepasses to be sited conveniently so that tramming distances from the loading points are not excessive. Development Drilling Once the skeleton development has been completed, the extraction headings are developed at 2.4 x 2.4 m as shown in Fig. 1. Standard development practice is to use crews of a machine operator and his helper, equipped with air-leg mounted jackhammers, to drill rounds of 1.8 m with integral tungsten carbide tipped drill steel. The round drilled is a normal drag round, as shown in Fig. 2, but considerable attention is paid to the drilling of the perimeter holes to use effectively the

Jan 1, 1982

By R. Malmstrom, H. Heden, K. Likin

Sublevel caving techniques at Luossavaara-Kiirunavaara AB (LKAB) in Sweden have undergone significant modification over the years as management strives to maintain mine viability amidst rising costs and erratic iron ore prices. LKAB, which is 96% state-owned, is Sweden's principal iron ore producer and operates three iron mines north of the Arctic Circle. By far, the largest of LKAB operations is the Kiirunavaara (Kiruna) underground iron mine which extracts an estimated 30 million t/y of ore, Other iron mines include the Malmberget underground nine, rated at 12 million t/y of ore, and the Svappavaara open-pit mine, rated at about 4.5 million t/y. Mined ores from Kiirunavaara and Svappavaara are shipped by rail to the northwestern harbor of Narvik, while Malmberget ores are sent to Lulea, located to the southeast. LKAB iron ores are marketed by Malmexport AB, which is owned jointly by LKAB and Granges, Sweden's base metal producer.

Jan 10, 1979

By R. A. Bowen

THE Montreal mine, operated by the Montreal Mining Co., is four miles west of Ironwood, Mich., in Iron County, Wisconsin. It is the westernmost producing property on the Gogebic Iron Range of the Lake Superior district. Its production is 4000 tons of hematite per day from an average depth of 2300 ft. This tonnage is hoisted through a five-compartment vertical shaft. Several mining methods are in use but in the larger orebodies the method is sublevel caving. ORE OCCURRENCE AND CHARACTERISTICS The main orebodies of the Montreal mine lie in eastward-pitching troughs formed by the intersections of intrusive dikes with impervious members of the iron formation. The formation has an average width of 425 ft., strikes east and west and clips north at approximately 62°. The dikes dip south at 45° and bear north 20° from the strike of the formation, making the intersections with various members pitch east at 16°. The dikes are not continuous across the formation but are interrupted ,by a longitudinal bedding fault, which has a throw of 380 ft. vertically and 885 ft. horizontally. North of the fault zone the formation is called "hanging," south it is called "footwall" formation. The faulting has created two ore-bearing horizons; viz., at the inter-section of hanging dikes with the slate members and the foot dikes with the quartzite footwall. In each formation the main ore-bearing member consists of wavy beds of hard chest interlaid with hematite. In the most productive part of the mine, where the sublevel caving method is employed, a hanging-formation dike lies in juxtaposition with a footwall dike and the iron formation carries ore in certain sections from the quartzite footwall to the hanging (like. Here a horizontal section of the orebody may be 300 to 400 ft. wide for a length of 150 ft. or more and tail out lengthwise in both the foot and hanging formations. The footwall ore section is much longer than the hanging wall because of the orebody's greater height above the "pitching dike."

Jan 1, 1938