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Underground mining.

Underground Mining

THE 1991 Engineering & Mining Journal annual project survey recorded only a small decline in the total value of new mining and mineral processing projects worldwide compared with the previous year: 296 in January 1991 compared with 318 in January 1990. The projected capital costs were $55,200 million and $59,700 million respectively. The survey was believed to |strongly underscore' the trend towards greater stability, with robust markets returning for certain base metals. Projected investment in North American copper production at $1,260 million almost doubled the previous year's figure, whilst in Asia projected investments soared to $1,680 million, nearly triple the previous year's total. Worldwide productive capacity for copper was predicted to increase by as much as 775,000 t between 1990 and 1992.

In May 1990, Mining Magazine was sounding an optimistic note. Whilst most non-gold metal processes were down on the 1988-89 high, they were nonetheless almost all at levels well ahead of those prevailing a few years ago. This fact, together with the reduced lure of gold, might be expected to increase the funds available for exploration and development in the ferrous and base metals sectors.

In August 1990, International Mining forecast that over the next few years the mining industry will be dominated by three issues: a change in the balance of activity between precious and base metals operations; the impact of newly generated base metal demand from Eastern Europe; and increasingly widespread environmental concerns.

By early 1991, Mining Journal was reporting that there is a general consensus amongst traders and analysts that base metals prices will weaken through 1991 because of economic recession. In January the price of copper, so often a general indicator, was at a 29-month low and silver was at a 15-year low.

Methods and Mechanization

The rapid advance of mechanization in South African gold mines was well illustrated by articles in International Mining (May and September 1990) on experiences at JCI. The company's |Trackless Mechanized Mining Methods' programme began in 1985 and by last year its advance and success had been such that 800,000 t/month, 75% of the group's total gold ore output, came from mechanized mining. Benefits included reduced mining and labour costs, increased efficiency, access to further reserves, reduced dilution and much higher safety standards. Apart from obvious direct financial benefits, there were many less tangible advantages, such as a reduced, more highly skilled and more stable labour force based locally rather than recruited from neighbouring countries.

One of the trackless sections is at Randfontein Estates' Cooke 3 Shaft, 860 m below surface. Here, five reef bands merge, providing a mining horizon 4 to 18 m thick and making room and pillar mining possible. Using 48 trackless machines, this section has achieved a record monthly output of 89,000 t, with an average monthly profit of R4 million. Engineering costs are R14/t and the cost/t milled is R109.

Rooms are mined in up to three slices depending on thickness. The primary production fleet comprises six Atlas Copco Boomer jumbos, three Secoma Pluton 17 roofbolters, a Seco bench rig, three Toro 350D LHDs, three Wagner ST6C LHDs, one Wagner ST2B, one uprated Wagner ST3 1/2 and two Cat 966 wheel loaders. Broken rock is hauled to ore and waste passes by a fleet of articulated dump trucks from various manufacturers (Wagner, Bell, Subroc, Southern Denver), ranging in size from 15 t to 33 t. There are 11 utility vehicles, including four Case 580G backhoes with hydraulic breakers for scaling and two Cat D4H dozers for road maintenance. Great stress is placed on the need for first-class training facilities and on the vital role of engineering.

Engineering and Mining Journal (October 1990) reported on work by the U.S. Bureau of Mines to find alternatives in hard rock mining to drill-blast-muck methods. The Bureau has tested two mechanical systems, the radial-axial splitter and the ripper miner. Both have been shown capable of continuous hardrock fragmentation.

Limitations of drill-blast systems are most apparent where space is at a premium and restricts the movement of men, materials and equipment. Other problems include blasting fumes, overbreak, miner safety and blast vibrations. Continuous miners using drag cutters have found widespread acceptance in coal and other soft rocks but are not suitable for hard rocks. Tunnel and raise boring machines with hardrock disc cutters are now well established and the more versatile Robbins Mobile Miner has been a qualified success in trials at Mount Isa.

The radial-axial splitter is a development of the well-known hydraulic splitter, which consists of inserting a tool into a pre-drilled hole and using hydraulic pressure to a wedge and feather arrangement which breaks the rock in tension. The tool needs a free face to break towards and is thus suitable for breaking boulders but not for a confined site such as a tunnel face. The radial-axial splitter overcomes this problem by adding a hydraulic ram which pushes off the hole bottom. This applies an axial load to the rock and the resulting stress field of radial and axial loads pulls a cone of rock from the face. The Bureau has field tested a version of the tool, mounted on the boom of a drill jumbo alongside a percussive drill. In underground limestone, using a 2 hp air-driven hydraulic pump to power the splitter and drilling a 3 in hole, the system was capable of excavating about 1 1/4 s.ton/break. In a stoping situation in limestone, the Bureau estimated that the system could excavate 100 s.ton/boom/shift. This figure would reduce to about 40 s.ton in a confined tunnel.

A ripper cutter attacks the rock at a shallow angle with a broad, slow-moving bit which causes the rock to spall ahead of the bit, giving a conchoidal fracture surface. The Bureau tested a ripper miner on a simulated face made of concrete and concluded from laboratory tests on blocks of rock that rock with uniaxial compressive strength of 1,860 bar, and possibly more, can be fractured by ripping. In soft to moderately hard rock (up to 1,380 bar) the number of bit changes per 8 h/shift was estimated to range from 0 to 10. For hard rock (more than 2,070 bar) bit wear or instantaneous bit failure could be prohibitive.

Mining Magazine (December 1990) reported on the use of hydro-power to replace compressed air at Kloof Mine in South Africa. This was first considered by the Chamber of Mines Research Organization in the late 1970s. The concept uses the difference in elevation between the shaft collar and the underground working to provide a supply of high-pressure, chilled water which can be connected to the stopes through pipes and valves. A common reticulation system can be used for power supply and for cooling the working places.

Specially designed rock drills powered by pressurized water-emulsion are mounted on thrust-legs and offer significant advantages over pneumatic drills. These include a doubling of drilling rate, significant reduction in working costs, reduced dust levels and the elimination of fogging. At present, one stope at Kloof is operated entirely by hydro-power and it is envisaged that eventually 40% of the mine might be powered hydraulically.

An article in Mining Engineering (October 1990) examined the impact of technology on underground coal productivity in the U.S. 1977 was a recent low point for productivity in the U.S. coal industry (Fig. 1). The underground sector productivity has risen from 7.9 t/man/day in 1977 to 16.5 t/man/day in 1987, equal to a compound annual growth rate of more than 7%. Further analysis indicates that the average growth rate per unit for longwalls has been 16% during this decade, and for continuous miners 7-9%.

Various technological changes are believed to be the main contributors to these productivity increases. Newer and better longwalls now routinely produce at 50,000 t/week. The introduction of shield supports of 550 to 650 t, double the capacity of 1970s chocks, has been followed by more powerful and reliable shearers, wider conveyors and remote pumping stations. Faces are longer and wider. Dust abatement procedures allow more faces to cut in both directions. Higher voltages and electro-mechanical, microprocessor-controlled shields were introduced. Further improvements may be expected from increased web depth and automatic advancing of shields.

The introduction of more reliable solid-state controls and other improvements has reduced average unplanned down-time/shift from 90-120 min to 60-90 min. Some of this change may be attributed to a maturing work force. On continuous miners, the horsepower has increased, bit technology has improved and loading conveyor capacities have been raised. A typical time study in the mid-1970s would seldom show mining and loading rates in excess of 4.5 t/min, whereas today's rates approach 9 t/min.

1990 was the centenary year of LKAB, the company which owns and operates the great Kiruna iron ore mine in Swedish Lapland (Engineering and Mining Journal, August 1990). The principal method in use today is sublevel caving with a 22 m vertical sublevel interval. Current projects include a modified blasthole stoping method using giant stopes and very long blastholes.

Development work has started between the 610 and 740 m levels with a hybrid form of sublevel caving known as blasthole stoping or super stoping, which is effectively sublevel caving on a giant scale. Two sets of three crosscuts, vertically displaced by 130 m, are driven through the ore on the 610 and 740 m levels. Mining will involve drilling downholes up to 120 m in length. A series of stopes with hole length increasing from 40 to 60 to 80 m has tested drilling accuracy and effectiveness of blasting. Stromnes Rabor and Atlas Copco H222 and H450 jumbos are used to drill a fan-shaped series of blastholes upwards from the central of the three lower crosscuts along the full width. These holes are charged, blasted and mucked out to provide a funnel-shaped collection trough. A 2.4 m raise is bored with a Robbins R82 machine between the two central crosscuts to form a starter slot for blasting. Upholes are drilled to slash out the lower part of the raise to ease the start of slot blasting (Fig. 2).

Automated, remotely-controlled Atlas Copco Simba 269 rigs drill a series of 165 mm blastholes down from the upper central crosscut. These are blasted to create a slot running the complete width of the ore and provide a free face for the remaining stope blasting. Holes are then drilled down from the outer two upper crosscuts and the ore is broken off in slices to fall into the collection trough. The Simba 269 uses a COP42 high-pressure DTH hammer with 114 mm drill pipe with guide flanges to aid accuracy. The goal is to achieve an average drilling rate of 65 m/shaft. Blast patterns are calculated to break 45-60 t/m drilled and the maximum size blast is designed to break 200,000 t using ANFO and emulsion explosives. Broken ore is drawn from the trough through an angled drawpoint and trimmed to the ore pass. By drawing off only the swell volume, the stope can be kept full to ensure wall rock does not scale off and dilution does not occur until the final blast, when the bridge above the upper crosscuts is breached. To date, hole deviation has been less than 0.5% of hole length.

Drilling and Blasting

Inability to locate a drill hole accurately is a common constraint in mining. Success achieved at Pasminco Ltd's Elura Mine in New South Wales appears to open up new frontiers (Mining Magazine, August 1990). The challenge was to drill a 93 mm hole and to keep the hole within a 300 mm dia. over 400 m length. This accuracy was necessary to ream the pilot hole to 330 mm and then raisebore a 6 m-dia. shaft. On the first two attempts, the bit wandered out of the 300 mm target cylinder within the first 50 to 60 m and the holes were abandoned. A deviation of 0.5 [degrees], which is not uncommon, would have caused 23 m displacement over the 400 m hole length.

The directional drilling group of Christensen Boyles recommended the use of a Welnav steering tool manufactured by Wellbore Navigation Inc. of California. This provides surface readout information, interpreted through a cluster-software computer package. A |Navi-Drill', which is a specially stabilized directional drilling down-hole motor and deflection shoe, was recommended to correct deviation. With a hole survey every 6.1 m core interval, the Navi-Drill could be employed to correct any deviation.

At a depth of 48 m, displacement of 0.43 [degrees] was measured which, if uncorrected, would have caused 14 m deviation. The directional tool drilled 7.65 m to change the inclination by 0.54 [degrees]. This procedure was repeated each time the hole wandered off course. As directional drilling corrected the angle and direction of the hole, a pilot bit, reaming shell and incline reamers were used to ream the hole to a size to accommodate the stabilized core barrel.

An article in Australian Mining (August 1990) reviewed the history and latest developments in percussive drill bits. Shortly after the Second World War, Sandvik pioneered the development of cemented carbide-tipped steel tools. The new material was a composite of wear-resistant tungsten carbide and cobalt, which acted as a |cement' to bind the carbide grains together. This was brazed to the steel and gave a 100-fold increase in the service life of the tool. A limitation of the new tool was that the greater the wear resistance, the more brittle it would be. Much of drill bit development up to the 1970s concentrated on optimizing the cemented carbide grade and on brazing techniques needed to attach this to the drill bit substrate.

The 1970s saw the advent of the button bit. Dome-shaped carbide buttons made to closely-controlled tolerances were assembled into the bit by press fitting, avoiding the problems associated with brazing and enabling much more wear-resistant carbides to be used. Sandvik has developed Dual Property (DP) carbide. This exploits the role of the cobalt content in varying the hardness/toughness properties of the material.

By selectively modifying the ratio of the two constituents throughout the cross-section of the button it is possible to increase the wear resistance and the toughness of the insert simultaneously, or either property, independently. DP cemented-carbide buttons comprise an outer shell of very low cobalt content, an intermediate high cobalt zone and a core of nominal cobalt content but with a special tungsten-carbide structure. Permutations on these zone structures are possible, providing the means to |tailor' drill bits to specific applications over a broad range.

These new tools are claimed to improve productivity and reduce costs. Examples cited include an increase of average bit life from 1,000 m to 3,121 m in limestone tunnelling work, with penetration up from 1.48 to 1.96 m/min.

Nitroglycerine is increasingly giving way to emulsions and low-cost ANFO compositions. Emulsions are prepared by blending together a hot super-saturated solution of oxidizer salts, mainly ammonium nitrate, into an oil-wax mixture with surfactants. The product can be made primer or detonator sensitive, can be given varying consistencies, has excellent water resistance and a high velocity of detonation, and generates low levels of post-detonation fumes.

The ICI Magnadet electric detonator provides protection against stray currents and static electricity. This is achieved by coupling each detonator via its own transformer to a special exploder. This operates in the range 15-30 KHz and so the detonator is immune to standard 50/60 Hz AC supply and DC currents.

Non-electric initiating systems also protect against stray currents. These are based on a small-diameter plastic tube coated internally with explosive powder. Initiation, by detonator or detonating cord, sends a shock wave down the tube to initiate the delay element of a detonator or primer charge. The system is able to provide accurate timing of the blast.

Prototype testing of an electronic detonator is taking place in several countries. This contains its own miniature electronic timing circuit to replace the chemical delay element in the conventional pyrotechnic detonator. Electronic detonators are programmed with a delay sequence number during manufacture. In the blast they fire in sequence number order with a constant delay time interval between consecutive sequence numbers. This interval is selected by the blasting engineer and set in the exploder, which then communicates the precise delay to all the detonators in the blast. Each detonator has a filter for protection against stray currents, radio waves or static electricity.

Sica in Finland has developed the |Power Cone' which fires a metal slug at hang-ups in ore passes, drawpoints and chutes. This is available in sizes from 3.2 to 21.5 kg with a range of 200 m. A similar South African system fires a 2 kg copper ballistic disc (or |Slugshot'). Using a 7 kg base charge of high explosive, the disc travels at over 2,000 m/s and will shatter quartzite rock up to 1 m thick.

Most safety explosives currently in use in British coal mines are sensitized by mixtures of nitroglycerine and ethylene glycol dinitrate. NG content ranges normally from 10% to 30%. Early trials of emulsion explosives failed because, lacking a recognizable flame retardant such as common salt, there were ignitions of methane/air mixtures under test conditions. Compositions containing varying amounts of flame inhibitor were examined, one of which passed the full range of tests to be classed as a P1 Permitted Explosive (The Mining Engineer, January 1990).

Comparative trials over a period of a year in 1987 were agreed by British Coal and the Health and Safety Executive, involving 250 t of explosive. The emulsion was found to be as effective as a well-established NG type but the relatively soft cartridges made charging more difficult.

An important advantage of emulsion was an average 60% reduction in peak concentration of carbon monoxide, and the volumes of nitrogen dioxide and oxides of nitrogen were on average 2.5% and 20% respectively of the amounts produced by NG explosives. During 1988 some 32 coal mines held trials using amounts varying from 100 to 2,700 kg. Four mines reported misfires but otherwise the emulsions performed well, giving improved fragmentation. The author concludes that in time these newer types of explosive will replace NG types, leading to a worthwhile increase in safety in transport, storage and use.

Mud blasting to break boulders is banned in many coal mines, including those of the U.S., because the danger of an unconfined charge igniting an explosive atmosphere is too great (Engineering and Mining Journal, April 1990). Sheathed explosives have an inert powder which forms a barrier between the flame front and the atmosphere, and have long been used for shothole blasting in coal mines. Since 1981 the U.S. Bureau of Mines has been involved in the development of a sheathed explosive charge that can be safely fired unconfined in gassy mines. The design is a 7 in-dia. disc, 1 in thick, coated with a damp layer of common salt. The salt spreads out as a dust cloud ahead of the explosion shock wave and serves as a barrier between the flame and the atmosphere. One explosives company has submitted such a charge for approval and, following successful tests, this should shortly be available on the market. In tests at one longwall mine which frequently experiences roof falls, it was found that downtime was 15-30 min compared with 1-2 h required when drilling and firing large boulders.

Loading and Haulage

Beginning in late 1989, an unprecedented takeover activity has occurred in the field of trackless equipment. Tamrock, who already owned Toro, announced that it was acquiring various business operations from Baker Hughes, including Eimco Jarvis Clark and Secoma. Some legal hurdles delayed the latter acquisition and Secoma was in discussions with l'Equipement Miniere (EM). The upshot was that Tamrock now effectively controls three of the world's major LHD manufacturers -- Toro, Eimco and EM-France-Loader. Meanwhile, Atlas Copco had acquired Wagner Mining Equipment so that both leading drill manufacturers have become major players in the LHD market. A further move early in 1990 was the acquisition of John Clark Inc. (JCI) by Canada's Continuous Mining Systems. JCI manufacturers a full range of LHDs and end-dump trucks which complement the CMS drilling, crushing and locomotive equipment.

LKAB's Kiruna Mine in Sweden has taken delivery of an unmanned, automated LHD, developed in conjunction with Toro, and trials are taking place at four crosscut drawpoints. The machines are electric Toro 500Es adapted for the installation of the control equipment. Beneath front and rear axles are antenna coils which determine lateral deviation to steer the machine. Analogue sensors (temperature, pressure, steering angle, speed, etc.) and digital sensors (on/off, limit position, alarm level, etc.) are installed. Their signals are sent to a control unit whose functions include steering, speed regulation, gear selection, bucket and boom operation, and the supervision of temperatures and levels.

Information transmission to and from the LHD and a stationary control unit is by radio waves. Four video cameras with video links have been mounted on the LHD to transmit TV pictures to the control room, to enable the operator to load the LHD remotely and to monitor automatic hauling. Conventional cable has been buried in the roadways in loops at a depth of about 150 mm to guide and control the LHD. A supervisory control system holds all information on the production area, such as roadways, speed and loop lengths. This constantly monitors the position of the LHD on the roadway and updates the control unit on the loader with driving instructions. The operator sits in a separate control room while the driverless LHD loads, hauls to the ore pass for automatic discharge and returns to the drawpoint. Oversize boulders can be loaded remotely and transported to a special boulder-handling facility.

Boliden's Kankberg Mine in northeast Sweden is designed for all-electric drilling, loading and haulage (Mining Magazine, April 1990). Initial planning envisaged a ramp-access mine (gradient 1 in 7) with cut-and-fill stoping using conventional diesel LHDs and trucks. Ventilating air at Kankberg has to be heated in winter which makes the use of diesel equipment expensive. Mine labour also has a preference for quieter, fume-free electric machines. The initial cost of installing electrical equipment is higher but this is offset by the long-term reduction in running costs.

Haulage is based on the new Toro 40E electric truck running from an overhead trolley line. The 40 t capacity Toro 40E has a standard AC heavy-duty 370 kW 1,000 V electric motor made by Siemens. It has computer-controlled power shift transmission, a modulated clutch and a fully hydraulic, computer-controlled horizontal cable reeling system. Driving speed at full load on the horizontal ranges from 4.5 km/h in first gear to 24.0 km/h in fourth gear. A computerized monitoring system logs tonnes loaded, distance driven and other parameters to give information on the performance and maintenance requirements of the truck, thus increasing availability. The dumper has an electronic gear selector which automatically selects the correct gear and 80 m of cable which could be extended to 100 m.

The trolley line has a present length of 1.2 km, including 200 m on surface, and extends down to the 170 m level. The mine's loading and transport requirement of 150,000 t/y of ore can be met by just one truck loaded by one Toro electric LHD. Mine management reported few problems once initial operator training and familiarization were completed.

Mount Isa is another mine which is trying out Scandinavian-developed electric truck haulage. (Australian Mining, August 1990). A low-weight, low-profile, rubber-tyred, electric Kiruna truck which can haul a 50 t payload up a 12% incline at speeds of up to 20 km/h is being used to haul up the T62 decline development. The truck takes power from an overhead 1,000 V trolley line fed by two transformers 900 m apart. It has two 239 kW dc traction motors and thyristor convertors. The driver can automatically connect and disconnect the hydraulically-operated trolley boom mounted in the cab in 3-5 seconds, even with the truck not directly below the trolley line. While operating on the trolley line, the batteries are automatically recharged. When disconnected, the truck performs operations, such as dumping and turning, on battery power. Mount Isa reports that the truck generates little heat and is quieter than diesel equivalents.

Continuous mining in coal is not genuinely |continuous' unless it is supported by a reliable continuous haulage system. Mining Engineering (March 1990) provided an overview of the available systems in the U.S., all of which are some form of belt or chain conveyor. Of these, some are in the form of interconnected bridge conveyors and some are self-propelled, flexible conveyors in one length.

Bridge-type systems were first in the field and still predominate. These receive material from the mining or loading machine and transport it in cascading and uninterrupted sequence to the panel belt and thence to surface or to a shaft. The inbye bridge may be attached directly to the tail of a continuous miner or it may be a self-propelled unit, operating detached, which receives the discharge from the miner on a hopper-type receiving arrangement.

Joy Technologies offers a single-length, flexible conveyor system called Flexible Conveyor Trains (FCT). These may be hung from the roof or be crawler-mounted. The FCT uses a patented, controlled pre-stretch belt to keep both edges of the belt in tension around curves. The units have a series of short frame sections connected with ball joints for flexibility. The crawler-mounted model uses these sections to form a flexible crawler frame fitted with a series of crawler pads. In effect, the whole bottom of the machine is one long crawler steered by wheels mounted on the end frames which carry the belt drives. The FCT is capable of negotiating up to four 90 [degrees] curves, provided they have a minimum radius of 8.5 m. It can be built in any length up to 14 t/min.


Shaft drilling was addressed in detail in both World Tunnelling (February 1990) and Mining Magazine (March 1990). In the first of these reports, it was noted that around 400 raise boring machines were in use worldwide, the most popular being those built by Robbins, Tamrock and Wirth. Shaft drilling is not as widespread as raise boring, but machines have been built in China, Germany, U.K., thE U.S. and U.S.R.R The article in Mining Magazine was written by Colin Pigott of Pigott Shaft Drilling Ltd, and concentrated on methods for steering. Machines built by Wirth, Zeni and Pigott were described and Pigott concluded that steered shaft drilling machines will shortly be able to drill to depths of 1,000 m at diameters to 5 m. Such shafts could be designed for specific purposes, for example man winding, mineral hoisting or ventilation, and be located where they were needed and where they were acceptable environmentally.

A paper by S. Keeble in The Mining Engineer (December 1990) reviewed the options for sinking shafts, based on studies commissioned by British Coal. These were intended to compile and evaluate conventional shaft excavation methods, including ground treatment and shaft lining, and to study alternative techniques and their limitations if utilized for U.K. projects. Table I illustrates the conventional and alternative methods.

Shaft drilling excavates a shaft from the collar downwards using a large rig to drill one or more passes: debris removal is normally by flushing with reverse circulation. The shaft is kept full of drilling mud to maintain stability and exclude ground water before lining (Fig. 3).

Shaft boring utilizes a special in-shaft boring machine instead of a drill rig, the debris normally falling down through a pilot hole to an underground muck-handling system. The blind boring system is a variation incorporating a system for collecting and raising debris to the surface. In both methods the shaft can be lined as sinking proceeds.

Raise boring requires a pilot hole to be bored into an underground chamber which is then reamed from the chamber upwards by an in-hole cutter operated by a machine at the top of the shaft. Debris falls into the lower chamber.

Raise drilling excavates from the bottom upwards using a drilling machine located in a chamber at the shaft bottom.

Shaft-drilling rigs are either modified oil-field rigs or specially developed custom-built rigs. All operate with reverse circulation mud-flush systems, where the drill fluid and chippings pass upwards inside the drill pipe. Drilling fluid is a major cost component and open strata may require pre-grouting in order to prevent excessive fluid losses. Deviation from the vertical can generally be maintained within 0.25%. Maximum depth is generally in the range 500-700 m, although some manufacturers claim a 1,000 m capability. The largest diameter drill available is 9 m, but the average diameter of drilled shafts is 2.1 m with the majority in the range 1.6-5.03 m.

In contrast with shaft drilling, which has been practised for well over a century, the first shaft boring machine was produced in 1970. Wirth in Germany manufactures four full-faced boring machines, whilst Robbins in the U.S. manufactures two machines. In addition, various manufacturers produce machines which bore only part of the face at a given time. Shafts of up to 8 m dia. have been constructed to 700 m depth. Shaft boring without pilot holes has suffered from failure to solve the problem of cuttings removal. A major advantage of the method is that lining operations may be carried out as sinking progresses with a wide choice of temporary and permanent lining techniques (Table II).

Raise borers were introduced some 20 years ago and have rapidly gained acceptance for constructing short, small-diameter raises in competent ground. There are six manufacturers producing a total of 33 different models. The largest is now claimed to be capable of excavating a shaft of 6.5 m dia. to a depth of 1,230 m. Limitations are similar in nature to those for shaft drilling, that is to say the strength of the drill string and the directional accuracy of drilling. The technique is mainly appropriate in dry and competent rock, as unstable and wet conditions must be pretreated by either cementing or freezing.

Raise drilling is similar to raise boring, in that relatively dry and stable rock formations are essential. However, the potential of this system is much more limited in terms of hole diameter. The manufacturer's rating for the largest available machine is 2 m dia. and it can bore to 380 m. The technique is therefore a mining aid rather than a shaft construction system.

British Coal shafts are likely to be at depths of 300-1,000 m with finished diameters between 3-8 m. Slim shafts of 2.5-3.5 m finished diameter were examined in relation to ground conditions at two sites. In the first of these the strata consisted of weak pebble and sand beds with high potential water inflow, overlying weak coal measure mudstones, siltstones and seat earths. In the second location there is a generally dry mudstone group with groundwater found in sandstone formations below 350 m in depth. It was concluded that the only practical and secure method of construction would be to drill the shafts. The linings would be either ring-stiffened steel or combined concrete and steel floated in and grouted into place.

When considering larger shafts of 5-7.5 m finished diameter, it was concluded that in U.K. conditions, the realistic technical limitation for drilling is 6 m dia. and 700 m depth. The study concluded that within this size range the only option is shaft boring with a pilot hole and that ground freezing through water-bearing strata would be essential.

The overall conclusion of the study was that although alternative shaft-sinking techniques are available and proven for specific applications, these are as yet limited. Clients are reluctant to finance expensive trials or share construction risks, and manufacturers are reluctant to develop prototypes on a speculative basis for such a restricted market. Shaft drilling stands to gain much greater acceptance when directional control techniques are proven and new and cheaper high strength materials provide better lining systems. Blind-shaft boring requires a reliable muck retrieval system and hydraulic, pneumatic and mechanical systems are currently under detailed assessment. Raise boring is well proven and reliable, but requires competent ground conditions. The use of the raise borer in frozen ground has recently extended the potential of shaft-boring machines using pilot holes.

An article in World Mining Equipment (October 1990) described the sinking and equipping of a 4.9 m dia. shaft to a depth of 1,660 m for the Thayer Lindsey Exploration Project located some 4 km north of Sudbury, Ontario, Canada. The contract for this was awarded in November 1988 by Falconbridge to J.S. Redpath. A shaft with full bottom blasting normally employs a V-cut in the blast design, requiring equipment to be kept 60-80 m from the face during blasting to minimize damage. These distances require additional movement of the sinking stage, disrupting and lengthening the sinking cycle time.

Redpath had experience in 1988 of completing a shaft extension for an Atomic Energy of Canada Ltd (AECL) underground research laboratory in Manitoba. This was 4.6 m in diameter and was deepened by 190 m. A full face blast was designed by utilizing a burn-cut rather than a V-cut. The AECL project demonstrated that it was possible to allow the sinking stage or other equipment to be within 16 m of the shaft bottom.

A two-boom hydraulic jumbo with full shaft coverage was designed and built. Minimum round length was considered to be 3.5 m which might be extended to 4.25 m. Because of the volume of muck to be handled and the time available in the cycle the standard Cryderman was eliminated as an option. A grab system did not look attractive for a 4.9 m dia. shaft and a single large Cryderman type unit was manufactured with a bucket capacity of 0.57 [m.sup.3]. This was named the "Brutus Mucker".

In the sinking sequence after installation of shaft steel setts and guides from the deck of the sinking stage, the stage was lowered to the drilling position. The jumbo travelled the shaft and was installed in its drilling position. For blasting, the stage was raised some 16 m, the jumbo removed to its nest on surface and the Brutus Mucker retracted into the sinking stage. After blasting the stage was lowered to its mucking position and after the muck pile was partially excavated the stage was set for concreting. After completing the pour of a 5 m section of a shaft, mucking was resumed until the rock pile was removed. The cycle was completed by raising the stage to the steel work installation position.

Shaft sinking began on March 13, 1989 and by the end of 1989 845.7 m of shaft had been completed, which included excavation and steel work installation. The project was approximately 60 days ahead of schedule. Average sinking advance was 3.05 m/day. Up to the time of writing the article, there had been no lost time accidents.

A paper in the Mining Engineer (October, 1990) described an interesting operation at Harworth Colliery to increase hoisting capacity. By the 1970s various improvements had been made to the hoisting arrangements with the result that the system was operating at over 100% above its original design limit and there was little possibility of further increasing output. In 1986 the Nottinghamshire area of British Coal targeted Harworth for an increase in output to compensate for that lost from older pits which had closed. The option of sinking an additional shaft with an output of 3.3 Mt/y was costed at 220 million [pounds] and this was not viable. After further discussions it was agreed to install a complete new winding system at Harworth No. 1 shaft in order to increase output from 500 to 800 t/h.

This would be achieved by building a concrete winding tower above the existing head gear, with a four rope, 4 MW friction winding engine with room for two 27.5 t skips to operate. In a three week holiday period the old head gear would be slid out from under the new tower and a new inner tower pulled in and all the new winding, guide and balance ropes installed. The major components of the project were: * Design of the new winding system. * Planning the project. * Engineering work before the change over. * The change-over operation.

A tower-mounted friction winder was selected as the best option. A double conveyance was selected in preference to a conveyance-counter weight system because when mineral is loaded into skips at one level, the double conveyance tends to be more efficient. On the same grounds, a single winder rather than a twin winder was preferred. Skips were constructed from a special steel alloy to reduce tare weight, shot-blasted with zinc and painted with a resin to increase the life of the skip to 10 y. Guillotine doors were preferred to the radial type as there are less moving parts. A direct coupled dc motor with thyristors was selected.

Initial project planning indicated a time of 2 1/2 y for the project, from concept to operation. This was significantly reduced by implementing a |fast-track method'. Prior to the change over, various engineering operations were concluded without interrupting coal production. These included the installation of a new run-of-mine conveyor system, installation of new measuring pockets, erection of concrete winding tower and installation of winding and associated mechanical/electrical equipment.

The change-over operation began on Friday, 18 August, 1989 at the start of the pit's three-week holiday. Skips, winding ropes and guide ropes were removed. Twenty-six 90 t-capacity hydraulic jacks were then positioned either side of the head gear before cutting the load-bearing legs of the head gear. The operation of winching out the old head gear was completed in under 2 h. The new inner tower was then winched into place and the ropes were installed.

Compliance testing was completed on the Saturday prior to the return to work and coal production started on the Sunday, one day earlier than planned. However, on Monday it was discovered that there had been a movement of 6 mm between the stator and rotor on the winding drum motor and repair of this was completed on Wednesday.


In West Germany the average depth of coal mining is 900 m and the deepest faces are almost 1,400 m below surface. The virgin rock temperature at 1,500 m is over 60 [degrees] C and locally, particularly around large faults, even higher temperatures are encountered. Engineering and Mining Journal (February 1990) reported that in the past decade, despite a significant increase in the average mining depth, climatic conditions in mines have not in general worsened, and have often improved because of the systematic installation of air cooling and a tripling of the installed refrigeration capacity.

Shift time, actual working time, break times and employment restrictions on German miners are subject to regulations covering permissable dry-bulb and effective temperatures. The effective temperature is measured by taking dry-bulb and wet-bulb temperatures at prevailing air flow rates and subjecting them to standard calculations.

A normal 8 h shift is only applicable where the effective temperature is below 25 [degrees] C at the working place. For effective temperatures up to 29 [degrees] C the actual working time is limited to 6 h, for 29-30 [degrees] C the working time is 5 h, and for 30-32 [degrees] C the working time is also limited to 5 h. Individual miners may not be continuously employed in such places for more than 4-6 months. Employment is prohibited at effective temperatures exceeding 32 [degrees] C. Despite the progress in climatization, shift time is reduced in more than half of the mine workings in the Ruhr because dry-bulb temperatures exceed 28 [degrees] C.

The first step for reducing high temperatures is to increase air flow along the face, and the average in Germany now exceeds 20 [m.sup.3]/s. However, there are frequent limitations on this solution, because average air flow velocities in open roadways must not exceed 4.5 m/s. Throughout the West German industry, almost all coal mines have refrigeration plants with rated capacity of more than 300 MW. Some 200 water coolers and about 300 air coolers are currently in operation.

In the early days of mine climatization, there was general reliance on local refrigeration plants of up to 1 MW refrigeration capacity. This decentralized form of refrigeration eliminates the need for expensive piping down the shaft and through the mine, but can normally only be applied to capacities up to about 4 MW. For higher capacities a central installation is more cost effective. It has therfore been increasing practice since the mid-1970s to opt for central refrigeration plants.

A paper in the IMM Transactions (May/August 1990) reviewed the ventilation and refrigeration in deep hot and mechanized mines in Australia. The development of refrigeration systems has tended to mirror deep-mining practice in South Africa.

However, many Australian mines use high-productivity, mechanized systems with large diesel-powered fleets. These have a significant effect on total heat load and also influence the ventilation rates which must adequately dilute diesel fumes. Furthermore, many Australian mines are in tropical or sub-tropical areas with high ambient wet- and dry-bulb temperatures on surface. This, together with the mechanization, means that refrigeration is often required at much shallower depths than in South Africa. Table III gives the design wet- and dry-bulb temperatures for three deep Australian mines. The 2.5% design values are used to determine the worst underground conditions and are those temperatures which are only exceeded for 2.5% of the four month summer period. If a refrigeration plant is required, this value will determine the maximum size of plant. The other three temperatures can be used to give an indication of plant operating costs.

Most mines in Australia have "hot working place" criteria, which are either based on statutory regulations for each state or are values negotiated between union and company. The most complete approach to this topic is that of the South African Chamber of Mines. A heat/balance equation is used to determine the cooling power of the environment that is associated with a risk that a given body-core temperature will not be exceeded. This limiting body-core temperature can be set at a value which is based on the probability of the occurrence of heat-related problems such as heat stroke. Most Australian mines operate a "six-hour shift" system whereby the standard shift length is reduced once certain climatic criteria have been exceeded for a given proportion of the shift. The most stringent of the criteria in use are at one Tasmanian mine where when wet-bulb temperature exceeds 24.5 [degrees] C either the six-hour shift was applicable or basic rate of pay was increased by 50%.

Based on design criteria currently in use, a first estimate of the overall ventillation requirements for mines using an open stoping method can be obtained from the expression [Q.sub.T]=160t + 120 where [Q.sub.T] is ventilation requirement in [m.sup.3]/s and t is annual production rate in Mt. For cut-and-fill methods the same expression is used with 320 t taken rather than 160 t. Table IV compares estimated and actual mine ventilation rates.

Developments in the design and operation of refrigeration plant in Australia are effected by two main factors, the high cost of power and the relatively shallow depths at which refrigeration is required. At depths likely to be encountered over the next 15 years, a surface-based plant providing chilled water in open circuit is normally the most economical option.

Computerized monitoring is revolutionizing mine environmental control. Continuous monitoring systems prior to computerization were based on hardware generally with one or two sensors wired to each read-out module. As more sensors were needed in the system, more modules and wiring were required, resulting in a costly hardware-intensive system. Because of the costs involved with hardware-based systems, monitoring can be on a spot check basis rather than mine-wide. The advent of computer-based systems has made it practical and economical to distribute hundreds or even thousands of sensors with monitoring and control functions handled in a central location.

A coal mine in Northern China has 600 sensors continuously despatching information on methane, carbon monoxide, oxygen and air flow to three computer terminals located in a control room on surface. The computer analyses the data, takes corrective action where necessary and pictorially represents the information for officials.

At Amax Coal's Wabash Mine in Illinois more than 2,200 sensors work in concert with the computer to provide information not only on gas levels but also to track production, activate conveyor belt lines and control the gates of a bulk rock-dusting system.

A typical computerized monitoring set-up consists of a central data station with outstations and remote sensors. Electronic signals sent to the outstations are sequentially transmitted to the central computer, but the outstation has the capability to provide reactionary functions without necessarily receiving instruction from the central computer. For example, if methane levels quickly rise to danger point an outstation can trigger an alarm to alert mine workers or shut-down equipment that could be the cause of an explosion. The central data station continuously polls the out station for input on conditions at various sensors.

Its logic function correlates data from the sensors in one or more locations to understand better and to diagnose the overall picture within the mine. It can be programmed automatically to actuate one or more controls in reaction to a mine event or it can signal the operator as to what sequence of actions should be taken on the keyboard to control the situation.

When a mine decides to install a mine-wide monitoring system, typical investment will be in the range of $20,000 to $500,000 with a commitment for at least 5/7 years. Clearly the range of hardware and software options need to be evaluated very carefully when making such a choice. However, the reward is a surveillance system that provides capabilities unparalleled by any other technology since mining began.

Water and Pumping

Despite all the technology now available, water still frequently presents major hazards and severe operational difficulties in underground mines. An interesting example of this was reported in the Mining Engineer (July 1990). Shaft sinking at Monktonhall Colliery in the Scottish coal field began in 1954 and was completed in 1961. The two 7.3 m dia. shafts are of conventional construction with a concrete lining varying in thickness from 300 to 990 mm. During sinking operations, pilot boreholes were put down to prove the nature of the strata lying ahead. Precementation was carried out where known coal wastes were encountered, or where water in quantity under pressure was expected. In total, eight waste areas were dealt with during the course of sinking. Shaft lining was designed in accordance with the criteria accepted at that time to withstand hydrostatic pressure to the base of the series of water bearing strata. The measures taken at the time appeared entirely satisfactory and from completion to 1972 very little water was apparent in either shaft and the small amounts of water make were relieved through the concrete lining in a controlled manner and used for dust suppression in the colliery.

However, a complicated inter-connected network of old workings existed. Water levels within these workings were controlled by pumping from several locations, the last of which ceased pumping in 1979. Adjacent collieries which had been abandoned by this time had pumped approximately 4,000 gal/min to control water entering the workings.

Abandonment of pumping resulted in the flooding of old workings and the progressive rise in ground water levels throughout the whole coalfield basin. From November 1975 Monktonhall experienced a steady increase in water entering both shafts as shown in Figure 4. The reasons for the leakage through the concrete lining to the shafts were a combination of external water pressure exploiting weaknesses in the lining and inadequate precautions taken when sinking through strata, which at the time were dry, but subsequently became flooded.

By late 1982 shaft water make had increased considerably and a decision was taken to undertake backwall injection in both shafts. Following test work a programme was begun and 59 t of cement was injected behind No. 2 shaft wall and 64 t behind No. 1 shaft wall. These operations greatly reduced the quantity of water make, but did not reduce the amount of water entering the shaft via relief holes. Within a year there was a substantial increase in the make of water in No. 2 shaft due to the relief holes discharging considerably more water than before. This now began to cause a problem for the pumping system. By October 1986 a further 61 t of cement was injected behind the No. 2 shaft wall reducing inflow from 550 to 120 gal/min. The problem continued to cause serious disruption to the operation of the colliery and added a heavy burden on finances. Consequently British Coal, in conjunction with consultants, investigated the situation and considered possible solutions which fell into three broad categories: * Reduce the water level in the basin. * Waterproof existing structures. * Reline the relevant sections of the shafts.

The problems of ground water lowering either from disused mine shafts or by deep well dewatering were the considerable capital outlay, ongoing operating costs and a possible delay of some period of time before any benefit was apparent. The problems of cement grouting were that maximum injection pressures were limited by shaft wall strength and previous grouting operations had sealed many of the larger voids but failed to seal the finer fissures. It was proposed that the SCEM grouting system be used to effect a relatively quick sealing of the shaft walls. The SCEM chemical is a patented blend of latex emulsions together with additives that promote flow and adhesion. It contains no toxic or hazardous materials other than ammonia solution which is present as a preservative at below 0.5% w/w.

The SCEM chemical is injected into the water flowing into a crack or fissure via a hole drilled to intersect the fissure for this purpose. The injected emulsion is activated by the turbulent agitation it encounters as it flows in the water and a steady build up of coagulated rubber laths occurs which then forms a seal. If the flow conditions dictate, a chemical activator can be injected either downstream from this chemical injection point or at the same point. The rapid coagulation of the emulsion which then occurs leads to the formation of a jelly like plug of matted rubber laths which seals the fissure. The chemical emulsion also requires a special pump designed to operate with minimal turbulence during fluid transfer.

In No. 1 shaft a total of 15,000 1 of SCEM was injected through 180 holes and succeeded in reducing water make from 634 to 23 gal/min. In No. 2 shaft the injection of 16,000 1 of SCEM reduced water make from 173 to 25 gal/min. The total cost of the remedial works was just over 600,000 [pounds]. This resulted in a direct saving on pumping costs of 172,770 [pounds]/y. Clearly there are many other indirect cost savings as a result of stemming the inflow of water and the operation was regarded as an outstanding financial and technical success.

Goldfields Namibia Ltd operates three base metal mines in Namibia producing in total around 2 Mt/y. There was a major interruption to production at the Kombat Mine in late 1988 when a notorious water-bearing fissure was holed during blasting. This released an estimated 5 Ml/h of water into the underground workings. The accident happened as a result of pre-cementation work underway some 535 m below surface. In little more than four days, the mine was completely flooded. The water level rose to just 50 m below the shaft collar.

Immediately two drill rigs capable of putting down 6.5 in dia. holes attempted to reach the development end and the breached fissure. However, the holes were deflected by adverse ground conditions and it was impractical to continue. Two large oil drilling rigs were at that time available in South Africa. The critical question was whether it would be feasible to drill holes of sufficient diameter and accuracy to intersect the 535 m level with the intention of filling about 50 m of the drive with cement up to the fissure to form a plug. Accuracy would be affected by rock formations underground and known magnetic anomalies which would affect any magnetic instruments. The proposed solution entailed drilling a 312 mm dia. mother hole vertically into the 5 m wide drift. This would be used to supply the cement necessary to fill the complete cross section of an area of the drift some 45 m back from the face forming a primary barrier. This would be followed by more holes deflected from the mother hole to intersect the drift at equal distance from the face and the primary barrier. Thus two more sections of the drift could be filled with cement forming a 45 m thick plug. To achieve these intersections, the target area for each hole was within a radius of 2.5 m.

The mother hole was to be drilled with Deutag's T-12 rig which is a diesel-mechanical Ideco H-525-D Dual Rambler. The mast has a nominal gross capacity of 117 t and a hook load rating at 113 t at six lines. The draw works feature a Micromatic type CB automatic driller and are powered by two Caterpillar D343PC-TA diesel engines.

Major logistics difficulties were overcome in mobilizing the men, drilling equipment and necessary ancillaries a total distance of almost 1,500 km. A special South African transport services train of 24 rail cars and 12 road trucks were used to haul a total of 580 t of equipment. This enabled drilling to start only ten weeks after the water was released.

The mother hole hit the drift 21 days later, a mere 200 mm from the centre of the target. Directional control was achieved by a steering tool and down-hole motor to correct deviations caused by difficult strata coditions. The steering tool sent data from the hole bottom via an umbilical line to the surface. Data includes direction of hole, direction of face and drill bit, inclination of hole, magnetic intensity and other parameters. This information is used to steer the down-hole motor which, by means of rotation at the hole bottom and a bent sub, allows the hole to be directed.

Once the mother hole had been completed, some 140 [m.sup.3] of cement and aggregate were delivered down a 965 mm dia. casing pipe to form the primary barrier. This was achieved in five days. The mother hole was cemented back to 300 m below surface and the first side track was drilled. This holed some 22 m closer to the face along the centre line of the drive. A further 650 [m.sup.3] of cement and aggregate were then placed, filling the drift between the primary barrier and the face. This hole was cemented back to where it broke away from the mother hole and the second side track was drilled. This intersected along the centre line some 40 m from the original mother hole. This second hole gave evidence of concrete and it was assumed that the necessary sealing had been achieved in that area. A third side track holed about 2 m back from the face. Here it was only necessary to inject 18 [m.sup.3] of concrete into a small void that was discovered, thereby clearly demonstrating the effectiveness of the two plugs. Dismantling of the drill rig began on March 29 and shortly thereafter dewatering of the mine commenced. By the end of September water level had dropped to 353 m below surface or 295 m below the level which had been attained after the inrush. No. 3 shaft was dewatered and recommissioning of the shaft and pump station was initiated.

Fosroc supplied a special underwater additive which prevented the cement from being washed out of the concrete mix and a strong super-plasticiser. These additives also retarded the concrete for up to 6 h. The design produced a high slump, very flowable but cohesive creamlike concrete which could be pumped down 530 m vertically as required. The mix allowed for the effects of free fall in the pipe and the placing of concrete under a very high water pressure of 48 bar.

The complete project from spudding-in the drill to final concrete tests, was completed in the very impressive time of 62 days.

An article in Australian Mining (July 1990) summarized the benefits which may derive from using modern submersible pump technology for mine dewatering. There are three different configurations which are principally used for a mine's main pump stations. These are: * Horizontally-mounted multi-stage, centrifugal

clean-water pump, driven by a non-submersible

electric motor. * Horizontally-mounted non-submersible

electrically-driven centrifugal pump of

heavy duty design. * Fully submersible electric pumping unit.

Multi-stage clean-water pump arrangements suffer from a number of weak points. * Inability to pump contaminated water

including abrasive particles. * The motor is not water-proof and hence there

is a risk of complete breakdown if flooding

occurs. * Servicing difficulties. * Limitations of mobility. * A drainage system is required in order to drain

water downwards before pumping it up again.

The author offered large heavy-duty high head submersible main pumps in combination with submersible feeder pumps as a solution to these difficulties. These pumps are not only impervious to damp and water, but are also able to handle contaminated water including abrasive particles. Maintenance is facilitated by the relative ease of transport of such pumps and excavation costs for the pump station are likely to be lower.

Submersible pumps are now available in sizes from the smallest of 1 kW up to the largest typical main pump of 90 kW with a maximum output of about 9,000 1/min. A more powerful 180 kW pump is also now available.

Engineering and Mining Journal (July 1990) reported that three South African Gold Mines have ordered tube presses to filter mine drainage water underground before pumping it to surface. Installation of the three tube press plants in the existing underground pump chambers will allow replacement of positive displacement pumps with more economical multi-impeller pumps and thus reduce energy and maintenance costs and eliminate thickening and treatment of solids at the concentrator.

The problem with underground filtration has always been that the large size of most filters would require expensive additional excavation and the filter cake might liquify on contact with wet ore and cause mud rushes. Tube presses overcome these problems. They occupy a very small floor area and can normally be accommodated in existing underground pump stations with an additional 2 m of head room. They have also demonstrated an ability to produce a hard cake that maintains its integrity throughout the skip loading, hoisting and discharge cycle.

The ECC tube presses ordered were manufactured by Charleston Engineering and are automatic filter presses operating at pressures of up to 140 bar. They consist of two concentric cylindrical tubes, the inner filter candle and outer hydraulic casing. Between candle and casing is a flexible membrane. Hydraulic pressure exerted on the flexible membrane applies filtration pressure to the slurry contained between the membrane and the candle. In operation, slurry is fed into the volume between the flexible membrane, filter media and end pieces. When full the feed is stopped and low pressure hydraulic fluid pumped into the space between the flexible membrane and casing. This reduces the volume and expels entrained air. High pressure is then applied and filtration commences. Once complete a vacuum retracts the membrane, the candle is lowered to open the machine and a pulse of air at the centre of the candle dislodges the cake. One of the mine's major concerns was the stability of the filter cake. This was tested by tipping a 55 gallon drum of filter cake into the bottom of a skip, filling it with wet run-of-mine ore and hoisting to surface. The procedure was repeated about 50 times and it was found that the cake retained its integrity and was still recognizable on the concentrator feed belt.

Ground Control

Measurement of in situ rock stress is normally undertaken by the slow, expensive method of over-coring. This requires drilling a pilot hole followed by a second surrounding hole, measuring the change in shape in the first hole as pressure is relieved. Australian Mining (October 1990) reported a quick and cheap way developed by CSIRO, using the |Minifrac' device marketed in Australia and overseas by Mindapa Pty Ltd in Melbourne. This system involves drilling a hole, normally 38 mm in diameter, and inserting two cylindrical packs, connected together, which are inflated by pumping in water or oil (Fig. 5). After inflation, the packs seal off a section of hole which is then pressurized by pumping in oil until fracture cracking occurs and the pressure drops. The crack is allowed to close and is then reopened. The pressure changes are monitored and enable the rock stress to be calculated. A longer cylindrical pack, covered with soft replaceable rubber, is then inserted and inflated. The rubber film is extruded into the crack and an image of the crack is imprinted on the cylinder surface. This image can be recorded and the direction of the crack indicates the direction of the stress field.

All equipment required to operate the |Minifrac' system fits into two portable containers. The basic unit with manual gauges is expected to be priced around $A20,000. More expensive models may include electronic pressure measurement. |Minifrac' can take stress measurements at depths to 20-30 m from surface or in tunnels. Greater depths will require more sophisticated equipment.

World Mining Equipment (April 1990) reviewed the current status of rock bolting. The expansion-shell bolt, in spite of its disadvantages of bleed-off of tension and unreliability in poor rock, still accounts for about 70% of all rock bolts installed today. This is probably because of its relative cheapness and immediate holding power. Wedge bolts are also still in use, although are far less popular. Both of these types are much more effective when they are cement-grouted for full-column support.

Cement-grouted rebar bolts are more reliable and can take high loads with good corrosion resistance, but have a relatively long setting time. The rebar/resin capsule types, as supplied by many manufacturers, can be loaded almost immediately after installation when bottom anchored and give fast full column support when they are backfilled with resin grout. Their long-term durability and effectiveness is adequate for most purposes.

Disadvantages of this type of bolt include high cost, limited shelf life and handling hazards associated with toxicity and fire risk. They can also be rapidly degraded by heat, age and in many cases moisture. However, continuing improvements have given this type of bolt an increasing share of the market.

Competition has been strong between resin and cement capsules. In the early 1980s low-hazard resin capsules appeared and were countered by self-contained cementitious capsules that were pre-packed with wax-incapsulated water. Resins won the day because of their fast strength development. Resin-held bolts need to be spun in. Steels in the UTS range 540-695 MPa are generally used in nominal diameters of 15-40 mm. These give breaking strengths of 10-60 t. A minimum pull out per 10 mm bond length of 4 kN is typically used in determining length of resin anchorage. The ratio of hole diameter to bar diameter is crucial, with hole diameter needing to be 5-12 mm larger than the bolt.

Corrosion is mostly controlled by grouting or by using galvanized steel. Glass fibre or wooden dowels are used to provide temporary reinforcement in locations such as coal faces where steel bar would damage cutting machinery.

Variations of the basic rock-bolt type include Atlas Copco's Swellex, Ingersoll-Rand's Split-Set and Du Pont's Fastorque. Swellex consists of an axially-folded steel tube which is expanded tightly into the hole after insertion by water pressure at 300 bar. Split-Sets are made from high tensile steel tube slotted along the length and pressed into the hole. Swellex and Split-Set support the rock along their full length and are quick and easy to install. Du Pont's Fastorque is a pre-tension bolt system that uses a patented, low-friction thrust-washer assembly and a fast-setting resin capsule. Its advantage is said to be that it gives greater anchorage strength for a given bolt diameter and grouting length than other systems, especially in weak strata.

In recent years there has been an increasing use of cable bolting as a passive support method. Conventional rock bolts seldom exceed 4 m in length, but cable bolting is in lengths of up to 40 m, often in patterns with single- or double-length cable typically 15 mm diameter and with 25 t breaking strain. A heavy grout, usually water and cement at 0.3:1, is pumped into the hole and the cable pushed through it. This is a difficult, slow job manually, but mechanized systems are now available for drilling, grouting and installing long cable bolts using a single machine. Cable bolting is mainly used for reinforcement of stope walls and roofs and for pillar reinforcement between stopes.

Approximately 30% of all rock bolting is done with lightweight, hand-held drilling equipment and manual setting of the bolts. Semi-mechanized bolting, for instance using mechanized drilling with manual bolt insertion, represents about 65% with fully mechanized bolting accounting for the remaining 5%. However, fully mechanized systems are increasing in number and may be expected to capture a larger share of the market.

A topic of importance since the earliest days of rock bolting has been the durability of the installation. World Mining Equipment (April 1990) also reported on the results of Swedish research on this topic. This was conducted by the Swedish Rock Engineering Research Foundation at the Centralgruvan Mine in North Kvarntorp. This mine has applied systematic roof bolting using various systems for over 20 years. Results have been monitored periodically by means of convergence measurement. The mine uses regular room-and-pillar systems in horizontally layered sandstone, which makes it easy to compare different reinforcement systems. The mine began in 1967 with the installation of cement-grouted rebar. In 1969 standard rock bolts with expanding shells were implemented, followed in 1972 by the use of resin-grouted rebar. In 1988 Swellex bolts were introduced.

Two methods were employed to test and assess bolt integrity. Boltometer measurements of 52 operational bolts selected at random from different reinforcement systems was the first method selected. The Boltometer is a battery-operated instrument with a sensor head that is placed against the protruding end of the installed rock bolt. A transducer transmits compression and flexural waves to the bolt, which propagate down the bolt at varying velocities depending on factors such as the quality and volume of grout. The waves are reflected back to the sensor head, which acts as a receiver and analyses the characteristics of the waves electronically. By processing the signal time history, the Boltometer determines the bolt length, overall bolt condition and the quality of grouting.

The second method of assessment consisted of over-coring selected bolts using a diamond drilling machine and removing the entire bolt together with the grout and surrounding rock. This was then cut to expose a cross-section at various points and also bisected. The condition of the bolt and surrounding grout was analysed and recorded.

The results of the Boltometer measurements were that only 20% of cement-grouted bolts showed optimal grouting quality, whilst a further 30% showed somewhat reduced grouting quality. The remaining 50% were unsatisfactory. Corresponding results for resin-grouted bolts were 37.5% showed optimal grouting with the same percentage showing reduced grouting quality and 25% were unsatisfactory. Results from the overcoring programme were not entirely satisfactory because some bolts were accidentally cut due to the core barrel being off-centre.

The principal conclusions of this study were that approximately 50% of cement-grouted bolts have reduced or insufficient grouting quality. Severe corrosion in the form of pitting was noted in cement-grouted bolts. Ungrouted bolts generally showed a lower degree of corrosion. Resin-grouted bolts indicated increasing rust formation with age. However, roof areas reinforced with cement-grouted bolts over 20 years ago showed no convergence in spite of reduced bolt effectiveness.

An article in Engineering and Mining Journal (May 1990) discussed the selection of shield supports for longwall mining. Longwall is becoming increasingly important in the U.S. Advantages include significantly lower accident rates than room-and-pillar mining, and a doubling of productivity over the past five years with an expected re-doubling over the next decade. Ground control is a fundamental requirement if the potential of the longwall method is to be realized. Face support now represents approximately 70% of the total cost of a longwall system with $6-8 million required to equip a longwall face with the latest powered supports.

Since the mid-1970s, shield supports have dominated and two basic types are available, two-leg and four-leg. There is a clear trend towards the use of high capacity supports and also towards two-leg rather than four-leg shields.

A census of longwalls examined relationships between support capacity and seam height, panel width and depth of overburden. No strong connections of these parameters were discovered, which suggests that support capacity selection has been independent of these factors.

The principal performance difference between two- and four-leg shield designs is the ability of the two-leg shields to maintain an active horizontal force. The inclination of the leg towards the face produces a horizontal leg component which is constantly trying to push the canopy towards the coal face. This force is transmitted to the strata through friction at the canopy-roof interface as an active horizontal force. This is most beneficial in maintaining the stability of highly fractured or friable immediate roof, by exerting a compressive force towards the face that tends to arrest fracture development. In contrast, because of the opposing leg inclinations of the four-leg shield design, horizontal components of the front- and rear-leg forces cancel out. Likewise, the chock shield does not produce horizontal leg-force components. Because of their larger capacity and resultant supporting force further from the face, four-leg and chock shields offer some advantage in heavy roof structures that tend to cantilever beyond the rear of the supports. These characteristics help to maintain moment equilibrium of the cantilevered strata. Another advantage of the four-leg and chock shields is a more uniform pressure distribution on the canopy and base. This can be beneficial in soft floor conditions.

The article concludes that the trend to select higher capacity supports appears to be levelling off and obviously cannot continue indefinitely. Shields purchased in the U.S. in 1989 were typically in the 600-700 s.ton range. The author foresees a continuing decline in the use of four-leg shields in favour of two-leg shields and the use of electro-hydraulic systems will continue to grow in the next decade. A major change in shield design is likely to be the application of wider supports. Historically, shield widths have been 1.5 m, but recently some 1.6 m and 1.7 m supports have been installed. Wider designs will reduce the number of hydraulic leg cylinders required per unit length of face and may help reduce support costs.

Health and Safety

The U.K. Inspectorate of Mines and Quarries annual report for 1988/1989 was published in February 1990 (The Mining Engineer, March 1990). During that year the continuing rationalization of British Coal's deep mine operations further reduced the number of mines, the number of working faces per mine and the numbers employed. Output was maintained by increasing productivity, which involved more intensive mechanization, the introduction of new techniques and procedures, and increased use of contractors.

Falls of ground account for 20% of all fatal and major injury accidents underground. Historically, the most dangerous place in a coal mine was always on the face. It is an interesting commentary on the way in which technology has developed that the focus of attention is now on haulage and transport operations as the principal source of fatalities and major injuries.

A total of 43 underground fires were reported in 1989 compared with 57 in the previous year, but the Inspectorate continues to express concern that such outbreaks are too frequent and there is still scope for further improvement. More than half of the fires involved belt conveyors. More comprehensive environmental monitoring is being introduced to detect underground fires at the earliest possible stage, but the Inspectorate's report emphasizes that fires do not develop without a source of fuel and that more attention needs to be paid to cleanliness.

For his presidential address to the Institution of Mining Engineers, Mr A. W. Davies chose the topic of |Influences on Health and Safety in Mining in the Next Decade'. Mr Davies emphasized the enormous steps forward which have been made in health and safety since the days when it was commonplace for more than 1,000 persons/y to be killed in the British coal mining industry. In contrast to that, it is now 11 years since an explosion claimed loss of life in a British coal mine. In addition to this, many health problems, such as pneumoconiosis, have been brought under control. The relationship between increasing productivity and safety is well illustrated in Fig. 6. Although there have been arguments suggesting that the opposite trend applies, statistics always support the conclusion that improved productivity produces improved levels of safety in the workforce. The most obvious reason for this is that, in an industry that faces many natural hazards, the number of persons at risk should be minimized.

Transport outbye of the face is still labour-intensive in mines in Britain, and this is an area which currently contributes the majority of underground accidents. The use of free-steered vehicles (FSV), whilst introducing new immediate hazards, should produce safety and health advantages in the long term. The prospect of one person and one FSV being able to provide servicing and supplies from the surface to the coal face must be preferred to a series of rope haulages which require large teams of men.

Winding systems in shafts still employ many people performing operations that have to be coordinated and therefore carry a risk of failure through human error. There is significant opportunity to introduce more sophisticated systems with less human involvement by the use of radio signals and other means and this undoubtedly will reduce accidents associated with winding in the near future.

The author commented on the present U.K. Government's intention to privatize the coal industry if it returns to power after the next election. The safety record of the small private licensed mining sector currently operating in the U.K. has been the subject of criticism. In 1988, which was the year that produced the least number of fatalities ever recorded in British coal mines, 12 people died in total. Nine of these were in British Coal mines employing over 100,000 and the remaining three in the licensed mines employing less than 2,000. However, the current licensed sector operates under severe constraints that frequently lead to very small mines working difficult pieces of remnant coal with inadequate resources and it would not be appropriate to extrapolate this experience to a major privatized coal mining sector.

A 90% improvement in productivity has been achieved by the U.K. coal mining industry since 1985. It was prophesized at that time by some that there would be a decline in health and safety performance associated with this push for higher outputs.

These prophesies have been proved completely false by the overall downward trends in accidents, and the improvement in general health standards of the workforce.

Engineering & Mining Journal (October 1990) reported that higher fines for many mine safety and health violations will result from the recent adoption of a new programme to identify mines with an "excessive history of violations" in the U.S. The Mine Safety and Health Administration (MSHA) indicated that such mines could also face stepped-up inspections and enforcement actions such as closure orders. According to William J. Tattersall, Assistant Secretary of Labour for Mine Safety and Health "increased penalty assessments and enforcement scrutiny at mines with excessive violations, should serve as a more effective deterrent and help reduce the number of violations at these mines". The programme will affect approximately 18% of all mine safety and health violations and about 7% of all mines.

Under this new policy a rate of more than 1.7 violations per inspection day at a mine during a two-year period, or 11 violations of the same safety or health standard in a one-year period, constitutes an excessive history of violations. MSHA penalty assessments for most of the serious violations at mines found to have an excessive violations history, will increase by 20-40% depending on the degree of excess. These mines could also face a five-fold increase in penalties for non-serious violations.

The Journal of the Mine Ventilation Society of South Africa (February 1990) reported that the Government Mining Engineer has now granted approval for three models of |ResQpacs' to be used in all South African Mines. This follows close collaboration with the industry and stringent testing carried out by the South African Bureau of Standards and the Chamber of Mines Research Organisation (CMRO). |ResQpacs' are self-contained self-rescuers, which protect mine workers from respiratory hazards that could result from fires or explosions by preventing the inhalation of noxious fumes and providing oxygen so that the wearer may reach a place of safety.

Impetus for the investigation into the benefit of self-rescuers was prompted in the early 1980s by the disturbing number of fatalities resulting from ignition of flammable gases and the concomitant over-exposure to the toxic products of combustion. In 1982, the CMRO initiated studies into self-rescuers. At that time, a number of countries had introduced legislation or terms of employment which required underground personnel to be provided with protective equipment to help safeguard them against noxious gases. The most popular device in use was the filter-type self-rescuer. After a detailed study of six coal mine accidents in South African collieries, it was clear that the most effective device would be a self-contained, self-rescuer which must be small and light enough to be carried on each person. In September 1984, the Government Mining Engineer issued a draft proposal for the introduction of legislation which would make it compulsory for all persons going underground to be provided with a body-worn, self-contained self-rescuer which would have to be worn at all times and would provide oxygen nominally for 30 min at a ventilation rate of 30 l/min. Prototypes were produced in 1985 by MSA and Draeger and in 1986 by Fenzy. Extensive field trials of these self-rescuers were conducted from the end of 1985. After a preliminary laboratory assessment, units were tested in the field in varying conditions over a six-month period. By the end of 1986 the |ResQpacs' were approved for use in coal mines, and in 1989 a more thorough test protocol was instituted to determine the suitability of the sets for all South African mines. Following test work, all three models were approved for general use in October 1989.

Mining Engineering (January 1990) reported on the development of a novel fire-warning system for underground mines. Present mine fire-warning systems such as stench, audible or visual alarms, telephones and messengers are often slow, unreliable and limited in coverage. Recent research by the U.S. Bureau of Mines has established the feasibility of ultra-low frequency electro-magnetic signalling for underground fire warnings.

In field testing completed to date, at five mines, signals from 630 Hz to 2,000 Hz were transmitted through rock up to distances of 1,645 m to a pencil-sized ferrite core receiving antenna and intrinsically safe receiving circuit. The prototype system uses off-the-shelf components and state of the art technology to ensure high reliability and low cost. This technology should enable simultaneous and instantaneous warning to all underground personnel, thereby increasing the likelihood of successfully evacuating a mine during a disaster.


Environmental awareness increases with each year that passes and the mining industry worldwide faces an ever-growing volume of legislation and regulatory requirements. This phenomenon is no longer largely confined to the affluent developed world. Ghana is a typical example of a developing country, heavily dependent economically on its mining industry, which is in the process of introducing comprehensive environmental regulations and which requires new entrants into the sector to undergo Environmental Impact Assessment (EIA). Supra-national funding agencies, such as the IFC, as well as conventional sources of finance, are increasingly sensitive to accusations of environmental unconcern and commonly require EIA even when this is not a stipulation required by the host government.

Australian Mining (November 1990) reported on the decision by the Australian Government to establish a national Environment Protection Agency (EPA). This was first mooted by Mr Hawke in a pre-election announcement and the Minister for the Environment has now confirmed that the Government is to go ahead with the plan. The aim of the EPA is to ensure uniformity of guidelines and standards across the country on environmental matters. The Government is expected to use the U.S. EPA, which has extensive regulatory powers, as a model for the Australian authority. At the same time, steps are being taken to implement a report proposing the merger of all Federal conservation agencies into a body called the Australian Nature Conservation Authority.

These developments pose new uncertainties for an industry facing what the Australian Mining Industry Council has called a |Shrinking Australia' and burdened by a range of restrictions and constraints on the exploration for and access to mineral resources. [Tabular Data 1 to 4 Omitted]

PHOTO : ARA TORO 300D diesel-driven LHD.

PHOTO : Left: Tamrock |Datamatic' drill rig. Above: Atlas Copco Boomer.

PHOTO : ABB-Kiruna electric truck in Mt. Isa mine, Australia.

PHOTO : M.A.N. GHH LF-9.2 Super Low Profile LHD with capacity of 9 t.

PHOTO : Atlas Copco Boltec 335 H-32 rock-bolting rig.

PHOTO : Fig. 1: Impact of technology on productivity.

PHOTO : Fig. 2: Mining method at Kiruna.

PHOTO : Fig. 3: Methods of Shaft Sinking.

PHOTO : Fig. 4: Water flows into shafts -- 1976-89.

PHOTO : Fig. 5: 'Minifrac' rock stress measurement system

PHOTO : Fig. 6: Fatal accidents in U.K. coal mines.

Dr John Stocks, Ph.D., A.R.S.M., C.Eng., Jay Mineral Services Ltd, Bissoe, Truro, Cornwall TR4 80Z.
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Title Annotation:1990 update of underground mining techniques and instrumentation
Author:Stocks, John
Publication:Mining Magazine
Date:Jan 1, 1991
Previous Article:Surface mining.
Next Article:Mineral and coal processing.

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