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By L. A. Morely

Volume II of the final report concerns the application of sensitive ground-fault interrupters (GFIs) to ac utilization circuits in U.S. mines The main concepts examined are shock prevention and methods for reducing nuisance tripping. The research involved a literature search, input from previous Bureau of Mines contracts, and contacts with electrical manufac-turers and mining personnel. Following this background work, several devices that showed promise were obtained and tested. Results of testing indicate the existence of GFIs that may be modified to perform adequately on U.S. mining systems. A recommended set of guidelines is included in the report, covering design, construction specifications, and performance tests.

Jan 1, 1982

By Thomas A. Lang, John A. Bischoff

The use of rock bolts for the support and stabilization of coal mining excavations was introduced over 30 years ago and, since that time, has progressively increased until today it is the primary means of support and stabilization of large underground openings. Early analytical studies and work with photoelastic and physical models established that even highly fragmented rock could be stabilized by rock bolting and form a load-carrying structural member that would span an opening. For stability, interaction of the bolts is essential and it depends on the bolt tension, the length and spacing of the bolts, the characteristics of the rock, and the applied load. However, the development of an explicit functional relationship embracing all these factors proved to be elusive. A review of the early rock bolting investigations led to the concept that the basic element of a rock bolted roof is the reinforced rock unit (RRU), consisting of an individual bolt and the rock immediately surrounding and adjacent to it, and to the development of mathematical equations which give the minimum bolt tension required to ensure that the RRU's are stable relative to one another and act together as a structural member. Analytical procedures have been developed for integrating RRU's into a reinforced rock structural (beam-arch) member which spans an underground opening and providing a rational basis for the design and installation of rock reinforcement systems.

Jan 1, 1982

By Willard A. Murray

Investigations of underground coal gasification (UCG) have been actively pursued by the U.S. Department of Energy since the early 1970s; the Lawrence Livermore National Laboratory (LLNL) conducted its first field experiment in 1976. In conjunction with field and laboratory studies of the process of recovering energy in the form of combustible gases from the burning of coal deposits in situ, an extensive environmental program has been conducted at LLNL under the auspices of both the U.S Department of Energy and the U.S. Environmental Protection Agency. This program has focused on groundwater contamination and ground surface subsidence caused by UCG. The material presented in this paper is a summary of results from the environmental studies of UCG activities conducted by LLNL as described by Mead, et al. (1981) and supplemented by the author's experience with this program in 1980 and 1981. The following sections describe specific groundwater contamination and surf ace subsidence problems encountered during field experiments in northeastern Wyoming, and will discuss implications of preliminary findings from geohydrological investigations of a site in southwestern Washington. Finally, possible environmental control technologies that are most likely to be effective in mitigating environmental impacts are discussed.

Jan 1, 1982

By Robert L. Bonomi

One of the major operating costs for any labor-intensive operation is that of its employees' wage and benefits package. Today's packages are extremely complex, with multiple pay rates for different grades and shifts, intricate bonus criteria, and expansive benefit packages that include all sorts of medical, dental, pension and other costs that can be "hidden" from direct examination and consideration. The problems of estimating the total cost of such a package to a company becomes even more acute at negotiation time, when pressure from both sides to settle intensify and changes to package offers come rapidly. A comprehensive system is needed that can evaluate these costs and perform sensitivity analyses, on the various components of an agreement so as to aid in the solving of major labor deadlocks, productivity problems, and cost impacts created by those agreements.

Jan 1, 1982

By J. C. Motz

Introduction The variations in crushing plant design seen during the past 30 years provide a valuable source of information and experience from which one can select the optimum arrangement for a specific requirement. The purpose of this paper is to examine some specific equipment configurations as they relate to operating efficiency, plant availability and final product size in the form of mill feed. Due to the varied ore properties, physical space limitations, size of ore body and financial resources, there is no single crushing or grinding flowsheet that is best. Since energy costs have risen dramatically during the recent past, it became obvious that significant savings are possible by selecting equipment and a circuit which offers low power requirements balanced with low wear metal and installed equipment costs. Relatively clear and measurable are the metal and maintenance costs and the ever increasing cost of energy. Metal costs in the form of crusher liners have not quite doubled in the last 20 years, while power costs have increased by a factor of four or five, depending on the region considered. Estimates made only a few years ago on the increased cost of energy are obsolete in view of the recent unstable conditions. Since power is therefore becoming the largest cost area, it behooves us to investigate potential cost saving methods in the form of lowering the KWH/Ton of crude. Crushing To reduce run-of-mine to mill feed requires reduction in a series of stages. The physical size and power requirements of a crusher capable of reducing hard rock vary depending on the application. Extended experience in most crushing plants show that a minus 19MM (3/4 in.) rod mill feed can be and is being produced in three stages of crushing. In the majority of cases, the last crushing stage is closed circuited using anywhere from 19MM (3/4 in.) to 14MM (9/16 in.) slotted openings on the sizing deck. In some copper concentrators, the mill feed is all passing 13MM (1/2 in.) using three stages of crushing and 14MM (9/16 in.) square openings on the closed circuit screens. It was felt by some that crushing to this fine size is only possible if the ore is friable and is relatively soft in comparison to the harder ores. Our results show that some of the copper ores are as hard as taconite and are crush-

Jan 1, 1982

By L. R. Hossner

Chemical and physical properties of surface soils and overburden materials must be evaluated prior to strip mining activities. This information can be used to assess whether subsurface overburden can be equal in quality to the existing surface soils, especially when these soils are classed as prime farm lands. In Texas, materials from the Carrizo, Calvert Bluff, Simsboro, and Hooper Formations overlie the lignite. The Carrizo Formation consists of thick clean sands with poor water and nutrient retention capabilities. These materials are least suited for placement of the surface of reclaimed land. The oxidized, medium textured overburden from the Calvert Bluff and Simsboro Formations possess adequate water and nutrient properties and are better suited for sustained plant growth. Non-oxidized Calvert Bluff and Hooper sediments generally have acceptable properties for reclamation. However, caution must be exercised in vegetating these materials since they may contain acid forming sulfide minerals.

Jan 1, 1982

By Frederick K. Allgaier

As part of an ongoing research program, the Bureau of Mines, U.S. Department of the Interior, is monitoring surface subsidence over six longwall panels at three Utah coal mines. These sites are characterized by mining depths of 244 to 610 meters (800 to 2,000 ft .), thick and massive sandstone layers in the overburden strata, multiple seam mining, extraction heights of approximately 1.8 to 3 meters (6 to 10 ft.), and rugged surface topography. Maximum subsidence, angle of draw, timing relative to mining, horizontal strain, and rate of subsidence development are being determined and correlated with mining and geologic variables. Results to date indicate the maximum vertical subsidence over the monitored panels ranges from 0.6 to 0.9 meters (2 to 3 ft.). The major factors controlling the amount of subsidence are the panel dimensions and the status of adjacent panel mining. Only minor surface cracking at the panel ends was evident at any of the sites. In all cases, there was a time lag of several months between the start of mining and subsidence at the surface. As subsequent panels are monitored and additional data become available, better estimates of the effect of various site characteristics on the subsidence will be possible. Ultimately, improved prediction capabilities will allow mine operators in the West to evaluate the potential for subsidence-related damage which could affect the current or future land use above the mine.

Jan 1, 1982

By D. A. Riva

"The Line Creek Project is located in southeastern British Columbia, approximately 25 kilometres north of Sparwood. The design is to produce a total of2.7 million tonnes of coal annually. This is made up of1.4million tonnes of thermal coal and 1.3 million tonnes of metallurgical coal.Construction commenced in the spring of 1980 and mining operations began in July 1981. The thermal processing plant, maintenance and dry facilities, as well as extensive road and bridge construction, were completed in early 1982. A collective agreement was signed with the International Union of Operating Engineers, Local 115, in June 1981.The first unit train left Line Creek on February 18, 1982 with a load of thermal coal bound for the Ssangyong Corporation in Southern Korea. It was the start of a ten-year contract to supply 350,000 tonnes annually to fire their cement kilns. Additional contracts for 750,000 tonnes will fuel the boilers of Korea Electric Company's San Chon Po and Go-Jeong power stations. Metallurgical coal shipments are scheduled to leave the Roberts Bank port in Vancouver in April 1983.Tracking an introductory presentation given at the 1981 CIM Annual General Meeting in Calgary, this paper highlights the salient developments by which ""The Line Creek Project"" has progressed during the past year to the"" Line Creek Operation"". Line Creek has com e on stream as scheduled and will meet all contractual commitments pertaining to timing, volumes and quality for thermal coal. The first train load left the plant in mid-February, and the first boat load to Korea departed Roberts Bank in mid-April. Metallurgical coal deliveries will commence in April 1983."

Jan 1, 1982

By Bob J. Rawicki

TYPES OF WET DUST COLLECTORS Basically, there are two types of wettype dust collectors. One is mechanical, incorporating pumps, motors, fans, sprays, filters, or flooded beds. These come in many forms, but their operating principle is basically the same. The screen or flooded bed is wetted by a series of sprays. Polluted air is drawn in through a fan; here the dust particles impinge on the screen and are flushed off into a settlement tank. The air then passes through spray eliminators to the atmosphere. A great deal of time and money has been spent developing this type of collector. In laboratory tests where clean water at constant pressure, air at constant volume, and dust at low but constant mass to volume were applied, these collectors gave high percentage dust collection efficiencies. Theoretically and technically, they are very good. In practice, however, under rugged mining conditions where there are no constants, and dust load, particle size, water pressure, and volume of air vary with time, their efficiency varies dramatically. Some of the most common problems are: 1. clogging of screens 2. blocking of nozzles 3. replacement. of damaged pump stators Any of the above result in loss of air flow and dust collection efficiency. Generally, there are so many mechanical moving parts, something is always breaking down. This type of collector is very expensive in parts and labor to maintain. The second type of dust collector, which I personally developed, with the help of the British National Coal Board of Great Britain, has none of the above mentioned problems and it works on a completely different principle. My collector, the Mark III Precipitaire Wet-Type Dust Collector, (an improved version of previously successful models) has no internal moving parts, no flooded beds, screens, spray nozzles, etc. It is highly efficient - 99.8% efficiency on total collected dust and 97% on respirable dust. The pressure drop through the collector is constant, hence, air flow remains constant. There is no maintenance expenditure, be it labor or parts. The only maintenance that there is, is desludging. This, of course, varies with the dust load. With a low dust load, desludqinq may only be required every 4 weeks, but this is a very simple operation that can be handled by unskilled men and it takes only a very short time. My collector is powered by a ventilation fan that is placed on the clean air side for a longer fan and motor life. Dust is extracted solely by water action. The collector's operating principle is: Dust laden air is drawn in along ducting into a tapered scrubber section, shaped to create a self-induced curtain of water. This action washes the dust particles from the air and the collected dust settles into the bottom of the tank. Cleaned air is then exhausted via a series of spray eliminators where it is completely dried and on into the atmosphere. These collectors have been sold in Great Britain, to the National Coal Board, and throughout Europe. They have been working very well for many years now. GENERAL APPLICATIONS Dust is very dangerous; silicosis and black lung kill and disable many workers each year. Many of the mine explosions can be attributed to coal dust. By the use and application of dust collectors, these deaths and disasters can be avoided. There really is no excuse. However good and efficient a dust collector may be, it's positioning and installation arrangement is critical. The collector may remove 99% of dust supplied to it, but it takes the skill and experience of a dust control engineer to design and fit ducting, hoods, etc. to remove polluted air at source and deliver it to the collector. In some cases this is very simple; in the case of longwall machines, it is very difficult. In general, dust collectors should be used to collect dust at every point of dust generation. I have noted a great reliance on dilution as a means of dust control. I personally think this is a dangerous and unreliable method. As mines become deeper and the use of larger and more productive equipment becomes more common, reliance on dilution will be impossible and dust collectors will be essential.

Jan 1, 1982

By K. J. Henderson

INTRODUCTION Shrinkage mining at Inco Ltd.'s Crean Hill mine can best be defined as a horizontal breasting method utilizing the broken ore for the mining floor and to provide wall support, with draw from a series of boxholes or draw¬points at the base of the stoping block. In choosing to shrink a given block of ore, the initial investigation centers on the ore configuration and grade, the expected ground conditions in the stoping area, the expected profit compared to alternative mining methods, the future production requirements, development lead time, the availability and type of equipment in the area, and the availability of experienced miners. These factors are weighed against other mining methods and shrinkage mining is chosen when the advantages outweigh the dis¬advantages. These will be discussed in more detail later. At Inco Ltd.'s Crean Hill mine, shrinkage mining makes up 17% of total mine production and is decreasing due to the advent of advanced blasthole techniques, notably in-the-hole drilling. Shrinkage is an adaptable mining method that is currently employed in narrow irregular ore zones, as undercuts for in-the-hole blast¬hole stopes, in ore bodies that do not extend from level to level, and in areas where quick production is required with a minimum of capital outlay or minimum develop¬ment. Two major types are currently in use which are classified on the basis of type of draw system employed. These are the utilization of load-haul-dump (LHD) equipment drawing from drawpoints, and the slusher and train method drawing from boxholes into a slusher trench and loading into a train. A third method being phased out is drawing from chutes directly into a train. In all three methods the actual stoping operation remains the same. In the initial planning of a mining block, the ore body is first located using an exploration diamond drill¬ing program. This is essentially done at a 130-m (400¬ft) spacing and it gives an indication of the amount of ore, grade of ore, the configuration of the ore body, as well as other geological features such as ground condi¬tions which will affect future mining. This drilling is done from an exploration drift 165 m (500 ft) long in the hanging wall and covers a vertical extent of 330 to 670 m (1000 to 2000 ft). However, these data are not accurate enough for detailed planning. When the de¬cision has been made to mine, a preliminary develop¬ment drill drift is driven in such a position that it can serve the dual purposes of allowing the ore body to be drilled off in detail, or of later serving as a haulage or access drift. Diamond drilling from this drift is usually on a 15-m (50-ft) grid spacing and gives adequate information for detailed planning. Utilizing the drill data, a decision is made on the mining type. For a shrinkage operation, the stoping block is ideally 49 m (150 ft) in length and a maximum of 13 m (40 ft) in width. Lengths of 65 m (200 ft) are considered a practical maximum because of the length of time the back of the stope remains open during the min¬ing of a cut. A minimum length of 13 m (40 ft) is used because the amount of time spent in opening a new cut increases in comparison with the total stoping time. The maximum width of 13 m (40 ft) is used because ex¬perience with the ground indicates that generally stope backs tend to be unstable over greater widths. However, widths of 33 m (100 ft) have been mined. A minimum width of 5 m (15 ft) is used to allow an effective draw, but in unusual cases, a width of 3 m (10 ft) has been mined successfully. The dip of the ore should not be less than 1.04 rad (60°). Angles flatter than this will not pull properly resulting in inefficient mining because of ore hang-ups on the footwall side of the stope and the need for large stagings on the hanging wall. In considering shrinkage, mining ground conditions play an important role. Strong shearing, cross faulting, or dikes are avoided. These usually cause dangerous mining conditions and an alternative safer mining method which will give adequate support to these adverse ground conditions is employed. Production requirements are taken into considera¬tion when planning a shrinkage mining program. In it¬self, shrinkage mining is capable of producing ore in a relatively short time with little capital development. If quick production is required, this method can produce ore rapidly as all that is required is that the draw system be completed and the miners with jacklegs and stopers produce a limited amount of ore. The obvious disadvan¬tage is that not all the ore broken can be pulled from the stope immediately. However, a mining horizon with sev¬eral shrinkage stopes in different stages of production can adequately maintain scheduled tonnage require¬ments, the cycle being stopes under development, stopes actively mining, and stopes completed and pulling. When the decision has been made to use shrinkage mining in a given block, the haulage and draw system is driven. The proximity of the orepass system dictates whether the system is to be a LHD operation or a slusher trench operation. If the orepass system is in close prox¬imity to the stoping block, LHD equipment is employed because of its ability to operate efficiently over short haulage distances. A long tram to the orepass usually dictates a slusher trench operation with haulage by train. Internal orepass systems are used effectively with LHD equipment with a load-out and retramming taking place on a lower horizon. In addition, stoping blocks in older areas of the mine are developed track because of the capital cost involved in converting an existing level to trackless mining. DRIFT DEVELOPMENT USING LHD EQUIPMENT The haulage drift for a draw system using LHD equipment is driven parallel to the ore body in the foot¬wall at a sufficient distance to allow an 11.5-m (35-ft) drawpoint to be driven into the ore (Fig. 1). The actual position of the drift depends entirely upon the ore con¬figuration and the drift size depends on the equipment to be used. Drift sizes of 4.6 x 4.1 m (14 x 121/2 ft) for an St-4 and 3.8 x 3.6 m (111/2 x 11 ft) for an St-2 meet company and government clearance standards and allow sufficient space to accommodate a ventilation pipe. The

Jan 1, 1982

The accurate determination of trace level gold in ores, metallurgical samples and process solutions is required because of the value of gold, the efficiency of the cyanide leach process and because of the high gold recovery from these cyanide solutions. It is thus necessary to routinely analyse geological and metallurgical samples with gold concentrations of 3 g tonne (ug g- 1ppm) and process solutions with a gold concentration of 0.01 mgL 1 (ppm) preferably with sufficient accuracy to quote at least two significant figures. Several techniques have been reviewed for their suitability for the routine gold analysis of these samples which also contain signif- icant concentrations of other ions. The classic fire assay technique has achieved the respect and confidence of mining personnel because of the results achieved by very experienced assayers. Other extraction and concentration techniques have now been shown to be equally effective. The use of instrumental analytical techniques particularly atomic absorption spectrometry have shown that cheaper procedures, which are as reliable, are now available to gold producers for the routine analysis required to maintain an efficient gold recovery process.

Jan 1, 1982

By Carl Erickson, James Hislop

INTRODUCTION As practiced today, longwall mining is a mechanical version of an established mining system. The equipment for longwall mining consists of four basic units: roof supports, face conveyor, winning or mineral-getting ma¬chine, and stage-loader conveyor. The system mines relatively thin slices of mineral across a predetermined face length in a virtually continuous sequence. As the winning machine cuts, the mineral is loaded automatically onto the face conveyor. In turn, the face conveyor is pushed over into the newly cut track by hydraulic rams attached to the conveyor and the roof supports. The roof supports then are lowered individu¬ally from the roof, advanced to the repositioned con¬veyor, and reset to the roof. At the end of the pass, the longwall cutter is pushed over, and a slice is taken in the opposite direction, so the system repeats itself. Fig. I illustrates a typical longwall mining system. The principle behind the longwall mining system is that the roof is supported only immediately adjacent to the worked seam, keeping the blocks of immediate roof clamped securely from the face to the rear of the sup¬ports. At the rear of the supports, the roof is allowed to fall and break up, with the swell of the broken material eventually establishing an equilibrium in the strata be¬hind the mined area. Fig. 2 illustrates the principle of longwall support. There are two distinct types of longwall mining. The longwall retreating system is the most common in the United States, South Africa, and Australia. The longwall advancing system is the most common in the re¬maining industrialized countries such as Britain, Ger¬many, Poland, and the Soviet Union. The longwall retreating system comprises a mining panel, where the roadways that form the head entry (end at which mineral is delivered from the face con¬veyor to a stage-loader conveyor) and the tail entry are predriven before the longwall units starts production. In the longwall advancing system, the longwall unit makes the head and tail entries as the face advances, utilizing various techniques to prevent the entries from being destroyed as they are left in a consolidating gob area. Fig. 3 illustrates the comparative approaches of the longwall retreating and advancing operations. As with most other underground mining systems, there are many variations on these two basic systems, using combinations of the systems that are dictated by

Jan 1, 1982

By Peter Hall

The need for large volume covered storage has increased drastically in the last ten years. Increased world demand for such products as thermal coal, potash, metallic concentrates, etc., availability of land for large open storage and environmental issues have brought about this need. This paper discusses briefly demand, environment, types of covered storage and the Alternative Storage System "Eurosilo", a proven system for storage of dry bulk mateials in units up to 100,000 m3.

Jan 1, 1982

By J. R. Brett

Utah Development Company (UDC) operates five large opencut coal mines in the Bowen Basin of Central Queensland, Australia. In total, over 20 million tonnes of coking coal are exported annually from the five mines. Typically, the coal seams dip at 5 percent, are 2 to 10 metres thick, and sub-crop some 20 to 40 metres below surface. In the main, the operations are single seam, but multiseam (2, 3 or 4 seams) operations do exist and will form an increasing proportion of the mining activities as depths increase. Draglines are used as the major stripping equipment, highwall depths currently ranging from 20 to 65 metres. Whilst project feasibilities were based on stripping to a 60 metre highwall, it is now proposed to continue opencut mining to depths exceeding 100 metres. This will entail large quantities of prestripping (ahead of the draglines) and overcoming problems associated with steepening seam dips, including haulage of coal from the pits.

Jan 1, 1982

By R. W. Thompkins

"PREFACEPart 1 of this paper, published in the September issue, covered the occurrence of uranium, radiation from uranium ores, the uranium series, radioactive theory, radioactive equilibrium and application to the radon series, the working level, the theory of emanation and diffusion, transport by thermal diffusion of radon emanating from porous rocks, and definitions.The October issue carried Part 2, which dealt with radiation theory applied to uranium ores, other radon migration factors, factors affecting the radon gas release rate, the measurement of radon gas emanation rates, the tunnel concept, variation of emanation rate with grade and estimating radioactivity from grade analysis.Part 3, in this issue, concludes the series by discussing the techniques for controlling radiation and providing some example calculations.17.0. Techniques for Controlling Radiation17.1. Altering the Factors In the Diffusion EquationIt has been noted that radon gas emanates from the surfaces of exposed ore to the mine atmosphere. The diffusion laws establish the rate of this flow. The fundamental relationship is that the rate of flow (J) is a function of the pore concentration and the distance from the surface. It is expressed as: []If any of these terms is varied, either naturally or artificially, the rate of gas emanation will be altered.Altering the Diffusion CoefficientThe literature describes many ways to lower the radon gas emanation rates which are dependent on changing the diffusion coefficient:1. A layer of ice on top of a tailings dam or on a soil.2. Humid clayey materials and swamped soils over a radon source."

Jan 1, 1982

By Tweedy. D. H.

"The paper describes the rapid development of, and operational experience in the utilization of, pneumatic low-pressure pipeline hoisting systems to vertically convey raw coal and rock through shafts, supplementing shaft capacity-the greatest lift to date being 535 m.A typical installation has raised over 250,000 tonnes in less than one year from a depth of 413 m through a 350/ 400-mm pipeline at a rate of 85 tonnes per hour. Operational experience on seven systems has provided valuable data on performance and operating costs.The main advantages of the systems lie in the flexibility of the pipeline to occupy existing free space in shafts and underground roadways, the relatively low first cost, the short delivery and installation time, and the economical running cost.Conclusions are therefore reached on future applications for example, the use of boreholes in older mines where new production technology has outstripped shaft or slope capacity, inter-level hoisting to implement concentration on transport roads and perhaps full mine output from relatively small mines.IntroductionIt is seven years since the basic art and practice of the pneumatic hoisting of minerals was described. The technique was developed in Western Canada by Radmark Engineering and found favour in the United Kingdom with the National Coal Board. Engineering and commercial feasibility trails were conducted in collaboration with the Mining Department and Mining Research and Development Establishment of the N.C.B. There was also considerable involvement with the British Health and Safety Executive.During the last five years, rapid progress has been made in the installation of new systems in the United Kingdom, Hungary and other countries. The operational experience gained makes it possible to report ""in depth"" on these installations, which cover a wide range of different physical mining applications."

Jan 1, 1982

By Heinrich Siemes, Christa Hennig-Michaeli

The effect of temperature and orientation on the plastic behaviour of an impure titaniferous magnetite from Palabora (South Africa) has been investigated. Prismatic specimens were axially compressed along [001], [110] and [111]. The strain rate was about 3.10-6 s-1. Trace analyses of glide lines revealed a plastic anisotropy due to glide on several non-equivalent crystallographic planes. At 200 °C {111}<112> twinning, {111}<110> slip and {100}<011> slip have been operative. At 400 °C {111}<110> slip and {100}<011> slip are primary glide modes. {110}<110> slip seems to be a secondary glide mode in the [001] specimens at 400 °C. {III}<112> twin glide generates isolated sharp lamellae. {111}<110> slip produces smooth surface striae. {100}<011> slip is characterized by coarse and partially wavy slip bands. {I10}<110> slip displays faint remarkably equidistant glide lines. The critical resolved shear stresses (CRSS) decrease with rise of temperature. The CRSS ratio of slip on [100) to slip on {lll} increases from 1.1 at 200 °C to 1.4 at 400 °C giving rise to a distinct flow anisotropy of the magnetite at 400 °C.

Jan 1, 1982

By Cox LT

A short term tyre study of trucks operating in a multiple pit surface coal mine showed significant potential for cost savings. The tyre study, conducted on-site over a three months period covered a mixed fleet of 18 trucks ranging from 27 tonne to 77 tonne capacity operating on overburden removal, coal haulage and watering duties. Some existing plant history records and a variety of tyre related measurements during the course of the study allowed the current state of tyre operations to be quantified. Specific attention was given to determining ger>ral trends from historical records and conducting detailed studies of individual lug damage, pressure monitoring and rectification, as well as tyre maintenance procedures. Important haul road parameters including camber, gradient, curve radii and superelevation, constructional materials and surface clean-up were appraised for their influence on tyre wear. Recommendations for achieving long term cost effective tyre utilisation included increased tyre maintenance, personnel development, compliance with haul road design parameters and more stringent clean-up Procedures.

Jan 1, 1982

By S. A. Tims

INTRODUCTION The Falconbridge mine ore body extends about 1.6 km (1 mile) in length and the deepest developed ore is on the 6050 level below surface. The ore zone varies in width from a few inches to over 30 m (100 ft) and the average width is 4.9 m (16 ft). Access levels are driven in the ore at 53.3-m (175-ft) intervals. The principal method of mining is overhand longitudinal cut-¬and-fill. Prior to 1962 timber square-set stoping as a secondary extraction method was used for about 15% of the total production. Undercut-and-fill was intro¬duced at Falconbridge in 1962 as a potential replace¬ment for the square-set method in heavy ground. The undercut-and-fill method was developed by Inco in the 1950s, its principal application being to transverse pillar mining. Falconbridge made modifications to this method. A feature of the mine is the No. 1 flat fault which dips 0.79 rad (45°) towards the northeast. The main characteristic of the fault is the presence of large swells of ore directly under the plane of the fault. The ground under the fault area is highly fractured and associated with massive sulfides. In the past, the ore under the fault was recovered, with difficulty, by either tight cut¬and-fill or square-set stoping. In the 1970s these meth¬ods were supplanted to a large degree by the under-cut-¬and-fill method. An advantage of the current undercut-and-fill method which uses cemented fill compared to the cut-and-fill and square-set methods is the reduction of dilution due to better control of the walls. At Falconbridge mine, it is estimated that the grade of ore produced by undercut¬and-fill is improved by approximately 10% over other methods. Where undercut-and-fill is used in very weak ground, a much greater improvement in grade can be expected. Table I shows mining production for 1974. The undercut-and-fill method was first used at Falconbridge during 1962. The first longitudinal stope was prepared for undercutting by laying down laminated beams the length of the stope and installing a lagging mat floor on top of the beams. Unconsolidated tailings fill was poured on top of the mat floor. As the cut ad¬vanced under the floor, heavy posts were placed under the laminated beams at 1.8-m (6-ft) intervals. During 1966, a radical change was made to the method when tailings fill, consolidated with portland cement, replaced the unconsolidated fill. This development eliminated the laminated beams and heavy mat floor and greatly im¬proved the stability of the stope. This system, with minor variations, is currently used at Falconbridge mine. APPLICATION The undercut-and-fill method is used to mine in¬competent ground, sills or floor pillars under mined-out levels, or a block of ore isolated between levels. It is occasionally used to advantage in sequencing produc¬tion from various mining blocks. This is done by mining a block of ore cut-and-fill method and at the same time mining the ore block directly underneath by the under¬cut-and-fill method. The undercut-and-fill mill holes at Falconbridge are either boreholes, stripped timbered raises, or steel mill holes. Boreholes and rock raises tend to slough in heavy and broken ground which increases dilution when sloughing exceeds the ore width outline and also in¬creases the difficulty of moving down to start the new cut. For example, in one installation, a 1.2-m (4-ft) diam borehole sloughed to a size of 3.7 x 5.5 m (12 x 18 ft). The undercut-and-fill method usually requires a mill hole extending from the level below the ore to the top horizon of the ore block. The customary methods of providing a mill hole are: 1) A borehole is driven from level to level through the ore block and a chute installed on the bottom level (Fig. 1). 2) An existing raise is used as a mill hole. If the raise is timbered, a steel mill hole is installed inside the timber and tailings fill poured around the steel mill hole (Fig. 1). 3) An existing steel mill hole, situated at one end of a mined-out stope, is used as the mill hole for an ad¬jacent undercut-and-fill ore block. The mill hole posi¬tion is determined when planning the mining sequence of the first stope (Fig. 2).

Jan 1, 1982

By Dennis N. Poffenroth

One concern of the Mine Safety and Health Administration is the structural integrity of steel wire ropes used on mine personnel hoists and elevators. Failure of such ropes has resulted in loss of life and property. Although thorough and regular visual examinations are performed, hoist ropes may still fail from undetected internal wear, corrosion, and broken wires. To complement visual examinations, nondestructive rope inspection techniques have been developed to determine deterioration and removal criteria. The techniques use electromagnetic and permanent-magnet test instruments designed by a Canadian firm. This paper describes the operating principles, field and laboratory testing of these nondestructive testing instruments, and discusses the feasibility of using this type of testing as a supplement to visual examination and destructive tests.

Jan 1, 1982