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By C. W. Fitzgerald

INTRODUCTION Continuous mining machines are used in regular flat ore beds having uniform thickness. Since coal is the most common type of material occurring in such beds, most continuous miners are used for the extraction of coal from underground mines. The first continuous mining machine for under¬ground coal mining was introduced in 1948. In one machine, the continuous miner combines the functions performed by a conventional face drill, a cutting ma¬chine, and a loading machine. It breaks or digs the material out of the surrounding solid and transfers the material back through the machine chassis by means of a chain conveyor. The conveyor carries the material to a haulage system, usually a rubber-tired vehicle known as a shuttle car. Generally, continuous miners are used in seam heights from 890 to 3660 mm (35 to 144 in.). For seam heights of 1070 mm (42 in.) or less, the haulage system may be a multiple-bridge conveyor or battery-powered tractors that haul low-bed trailers. Proj¬ects are being undertaken by private industry and through government contracts to develop a continuous haulage system for all seam heights. As of 1976, the majority of continuous miners were of two types. The first type is the drum miner, which mounts the cutting elements on a series of drums across the cutting head of the machine, sometimes using a cutting chain to cut clearance for the gear drive. In the drum-type machines, the cutting head rotates parallel to the material face. The second type is the auger miner, which utilizes cutting elements mounted on two auger-type cutting heads. One head is located on each side of the machine, with the heads mounted per¬pendicular to the material face. Chain-head ripper-type miners/miners having rotary cutting elements with oscillating motion and twin-head boring-type miners have all been popular within the last 15 years. Although these are considered obsolete, some machines are still being supplied for specialized applications. GENERAL DESCRIPTION This chapter describes only continuous mining ma¬chines available as manufacturers' production models as of September 1976. Figs. 1 through 11 illustrate some typical machines currently in use. Cutting Elements The cutting elements are the interface between the machine and the material being mined. In describing the cutting elements, consideration must be given to the differences between the drum- and auger-type machines, the bit-tip speeds, and the bit types. Drum-Type Continuous Mining Machines: Three of the four principal manufacturers of drum-type miners use multi-motor designs that provide separate a-c motors for driving the cutting elements. Gear trains are used to reduce the speeds of the motors to the rotational speeds of the cutting elements. One of the three suppliers drives the cutting elements of its low-seam models with a heavy-duty drive chain that is equipped with cutting bits; all other functions of this machine have separate motor drives that are either electrical or hydraulic. The fourth principal manufacturer generally provides two a-c electric motors, utilizing gear reducers, drive shafts, and hydraulically operated engaging clutches to power all of the operating functions, including the cut¬ting elements. Some drum-type miners use trim chains, equipped with cutting bits, to cut clearance for the gear drives. These trim chains vary from two 305-mm (12-in.) chains to one 760-mm (30-in.) chain per machine, de¬pending upon the manufacturer's design concept. In all machines other than the low-seam models that use chains to drive the heads, the trim chains are driven by sprockets on the drive shaft of the final cutterhead drum; therefore, the trim chains are not used for power trans¬mission to the cutting elements. In addition to cutting clearance for the gear drives, the trim chains provide uniform bit spacing all the way across the head and assist in moving the cut material to an area where it is either loaded directly into the miner's conveyor or is easier to load into the conveyor. On a drum-type mining machine, the cutting ele¬ments all rotate from the top to the bottom. Therefore, the recommended procedure is to raise the cutterhead to the maximum seam height, sump the entire machine forward to maximum depth by means of a crawler drive, and then shear downward until the cutter elements reach the floor level. The maximum sump depth varies ac¬cording to the drum diameter, the type of material, the floor conditions, and the operator's abilities. It normally varies an average of 460 to 760 mm (18 to 30 in.). Cutter drums are supplied with two types of cutting bit holders. Weld-on holders are attached permanently. Pin-on holders are replaceable at the face. Cutter-drum diameters over the bit tips vary from 660 to 965 mm (26 to 38 in.), depending upon the chassis height of the machine and the range of cutting heights. Auger-Type Continuous Mining Machines: As cut-

Jan 1, 1982

By Stanley J. Gasior

Most bituminous coal mined near Eastern industrial areas requiring high Btu pipeline gas is caking and therefore unsuitable for fixed-bed pressure gasification by present techniques. If the caking property of such coal can be destroyed the resulting fuel can be gasified in a fixed bed at pressure with steam and oxygen as a preliminary step in making a high-Btu gas. To find a practical way to convert strongly caking coals of granular or small lump size to noncaking fuel was the object of this study. In previous work strongly caking coals of 1/8- to 1-1/2-inch size was converted to a noncaking coal char by treatment in a static bed in about 3 hours [1/]and coal of 18-100 Tyler mesh in a fluidized bed in about 5 minutes.[2/] An inert gas such as `steam, nitrogen, carbon dioxide, or a mixture of all three with a small amount of oxygen was used in both studies. In our work a strongly caking bituminous coal from the Pittsburgh seam of 4-8 Tyler mesh size was converted to a-noncaking form in free-fall in about 2 seconds by dropping it through a countercurrent flow of steam containing 4 to 7 percent oxygen at 1,150°-1,200° F and 300 psig. This method should also be applicable to smaller-sized material.

Jan 1, 1965

By J. A. Dos Santos

High angle conveyor (HAC®) systems in open pit mining offer many advantages over the traditional, truck only, haulage systems, including: (1) superior energy efficiency, (2) less dependency on petroleum products, (3) less sensitivity to inflation, (4) less labor, (5) less excavation for the amount of ore recovered, and (6) less ramp construction and maintenance costs. Previous studies had revealed the technical and economic feasibility of high angle conveyors. This paper presents the Sandwich Belt HAC® as the vital link to optimization of in-pit conveying systems. HAC development to date is discussed, beginning with studies in 1979, to the first large scale prototype in 1983, to the first commercial unit in 1984, into "the latest commercial installations. Discussion continues with the HAC systems to suit the various open pit mining operations and the corresponding characteristics of the system components, particularly the high angle conveyor. The range of HAC systems from the low capacity (up to 1,814 t/h), low lift (10.7 to 36.6M), highly mobile to the moderate capacity (1,814 to 3,629 t/h), moderate lift (36.6 to 76.2M), intermittently mobile to the high capacity (greater than 3,629 t/h), high lift (greater than 76.2M), semi-mobile, are discussed along with some possible combinations of these.

Jan 1, 1989

By C. E. McFarland, N. W. Kirshenbaum

Cortez Gold Mines is located about 75 miles southwest of Elko, NV. Activity in the mining district dates from 1862. The current operation developed from a 1963 joint venture of Webb Resources, The Bunker Hill Company, Vernon F. Taylor, Jr., and American Exploration & Mining Co. Three years later, the USGS discovered gold on the property following a regional geochemical survey. Drilling in 1966 and 1967 identified 3.4 million st of oxide ore grading 0.28 oz Au/st, and a 1700 st/day countercurrent decantation mill was commissioned in 1969. Heap leaching started in 1971 on low grade ore and continues to the present. The mill was shut down in 1976 following exhaustion of known oxide reserves. Mobile equipment was sold and the proceeds committed to local exploration. In 1980, a new plan of operations was developed starting with existing low grade and newly discovered mill ore. The mill resumed operations in 1981 and has been operating to date. A series of metallurgical innovations, culminating in the commissioning of a new fluid bed roaster in the summer of 1990, has enabled treatment of a series of increasingly refractory ores. Current partners in the venture are Kennecott Corporation, Vernon F. Taylor, Jr., and Placer Dome U.S. Inc., manager of the operation. Idaho Mining Corp. holds a 7.5% carried interest.

Jan 1, 1992

By Harry M. Edenborn, Richard W. Hammack

Experiments were done to determine if a mushroom compost-based sulfate reduction system could be used to treat nickel-contaminated mine waters. Sulfate reduction systems were established in columns containing acid-washed mushroom compost. Simulated mine waters containing 2000 mg/L sulfate and 50 to 1000 mg/L nickel were adjusted to pH 4.5 and pumped through the columns at flow rates between 15 and 25 mL/hr. During the first nine days of operation, virtually all influent nickel was removed in the columns by sorption and ion exchange mechanisms. The nickel removal rate then dropped to 18 to 30 mg Ni/day (7.8 to 12.8 nmol/ g-substrate/day), probably due to low sulfate reduction rates. When sodium lactate was added to the inflow, a sustained and seven-fold increase in the nickel removal rate was observed. Bacterial sulfate reduction rates measured in mushroom compost-based wetlands exposed to coal mine drainage are 2 to 20 times faster than the rates observed in this study. Low sulfate reduction rates in the columns may have been due to the inhibition of microbial activity by high nickel concentrations and the removal of labile organic compounds from the mushroom compost substrate during acid-washing. Results of this study indicate that bacterial sulfate reduction can effectively treat nickel concentrations up to 500 mg/L if labile carbon is nonlimiting.

Jan 1, 1992

By R. W. Seedsman

Mandalong Mine was originally planned to operate with panel widths of 250 m and a predicted surface subsidence in excess of 3 m. The consent conditions applied by the NSW government on this mine plan were very tight and, following a change in ownership, were considered by the new operator to present a number of unacceptable business risks An acceptable risk profile for the mine layout restricted surface deformations to less than 500 mm vertical subsidence, less than 7 mm/m tilt, and less than 5 mm/m strain. Empirical subsidence prediction methods indicated that it would be necessary to reduce extraction panel widths of 80 m - an uneconomic option. The application of a number of rock mechanics concepts identified that the local overburden geology could be exploited such that panel widths of up to 175 m could be possible. The 125 m width was selected as a low-risk introduction of this new subsidence prediction method. The extraction of the first 2 longwall panels has demonstrated the validity of the prediction subsidence method and has provided the opportunity to refine the selection of key input parameters. The impact of mining through a zone of thrust faults was successfully predicted. The mine is now planning the introduction of 160 m wide panels, and will review the associated subsidence impacts prior to any further increases in width.

Jan 1, 2006

By Weston Bourret

The need for wartime supplies of uranium for the Manhattan Project prompted a 1944 study of uranium in South Africa. Uraninite had been reported by R.A. Cooper in 1923 in a Witwatersrand heavy- mineral concentrate, and the search for uranium, therefore, included the district's 30 operating mills. The method of investigation was simple: take bulk concentrate samples from the Wifley tables and determine their mineralogy. Samples were taken on a daily basis, although weekly composites also were taken of some mills. In 1944, most mills had simple gravity circuits and produced (I) a gravity concentrate, and (2) an all-slime [-75 pm (-200 mesh)] product. The tonnage throughput was known, and from that, head assays could be calculated. Concentrate samples were sent to the United States for chemical analyses. A Geiger counter set up in a Johannesburg hotel room provided preliminary data. The early data showed uraninite to be widespread. Microscopic, mineragraphic, and siezing tests were all under- taken in Johannesburg. Ultimately, tests led to a high grade (+90%) uraninite concentrate. Platinum and other heavies were the impurities. In seven months, the investigations had reached the stage where it was possible to forecast an annual production of 5.4 to 7.3 kt (6,000 to 8,000 st) of U3O8. The estimate is still realistic. Additional details are found in Bourret (1981).

Jan 1, 1991

By Juan L. Sepulveda

An MPSI KW-20 Tower mill has been tested at the Butte Molybdenum plant of Anaconda Minerals Company, as a substitute for the conventional 1.5x3.0 m (5x10 ft) tumbling ball mill currently used for regrind purposes. The Tower mill was operated in both open and closed continuous circuits. These results were complemented with a series of batch grinding tests to determine some of the operational features of this mill. The relative grinding efficiency of the Tower mill with respect to the tumbling ball mill was determined and the applicability of this new mill to a plant/heavy duty environment was assessed. Experimental results obtained on the regrind of moly concentrates indicate that approximately 40% energy savings can be realized by using the Tower mill instead of the ball mill.

Jan 1, 1984

By G. G. Marino, J. W. Mahar

This paper provides an utierstanding of the betivior and potenttil damage of homes resulting from sag-type mine subsidence. This is done with extemive research of numerom case histories in Illinois collected by the authors. The howe response to substience-itiuced ground displacements are summarized and evaluuted. A damage level criteriu, specifically developed for residenttil structures, is presented based primarily on repair costs. Damage levels and the associated subsidence profile characteristics for each case history are compared.

Jan 9, 1985

By S. C. Sun

A process was developed to recover coal, clays and pyrite from coal wastes. The process consists of fine grinding followed by coal flotation and pyrite flotation leaving the clays in the flotation pulp. A bituminous coal refuse, containing l0% S-30% C*, was treated by this method to yield a coal product containing 4% sulfur, 10% ash; a pyrite product containing 45% sulfur (84% FeS2), 1% carbonaceous material and a clay product containing 2% sulfur (3.5% FeS2). The coal yield was about 89%. The pyrite yield was about 77%. The process steps may be entirely flotation, or gravity separation (hydrocyclone) may be used to increase the pyrite: coal ratio in the flotation feed. Cost estimation shows a profit of $2.28 per ton of low pyrite grade refuse (Labor, maintenance, overhead and plant depreciation were not included). The development of this process consisted of three parts: 1) exploratory tests, 2) optimization tests and 3) confirmatory tests.

Jan 1, 1968

By Daniel Su

This paper describes recent ground control challenges encountered at the longwall recovery areas of one of Bailey Mine?s two longwall faces and the proactive measures taken by CONSOL management and engineering to mitigate them. The challenges derived primarily from the following causes: (1) irregular roof coal and draw slate thickness, (2) presence of thick sandstone at the recovery area and extremely weak rocks below the sandstone, (3) close proximity to a regional anticline, (4) low shield set pressure and tip load, and (5) ineffective primary roof support that led to progressive failure of the headgate roof. The presence of the anticline contributed to extensive fracturing of the sandstone. Excessive pot-out in advance of the shields frequently occurred near mid-face due to the presence of thick draw slate and irregular roof coal near the recovery area. This, in turn, resulted in face heights that exceeded the maximum reach of the shields. Subsequently, during longwall meshing, when the advance rate slowed considerably, fractured sandstone would often break loose, causing the shields to yield, thereby reducing the height and inhibiting shearer movement. To improve the headgate and face conditions within the longwall recovery area, several actions were taken, including: (1) assessing site-specific roof geology by conducting in-mine roof coring around the recovery areas and by mapping immediate draw slate and roof coal thicknesses within surrounding entries, (2) increasing pump pressure and thus shield set pressures (3) diligently monitoring shield set pressures and longwall extraction horizon, (4) installing additional cable bolts and roof screening to prevent progressive failure of the immediate headgate roof, (5) fully-grouting supplemental cable bolts to improve headgate support stiffness, and (6) gluing the headgate roof within the recovery areas with polyurethane. As a result of these collective efforts, longwall recovery delays at Bailey Mine were reduced from 21 days in 2006 to nearly zero in 2009.

Jan 1, 2009

By D. W. Park

Underground openings resulting from coal mining activities have been a serious cause of subsidence problems in the Warrior coalfield of Alabama, which is the 11th largest coal-producing state in the United States (National Coal Association, 1986). The problem will be even more serious in the future because 17 Gt (19 billion st) of Alabama's 22 Gt (24 billion st) of coal reserves are classified as an under- ground resource (Tolson et al., 1982). Subsidence data in this area has been scarce com- pared with other coalfields (Conroy, 1980; Dunrud and Osterwald, 1980; Gentry and Able, 1978; Hunt, 1979; King and Gentry, 1979). This paper discusses the results of a study of subsidence overlying two longwall mines that contrast each other. One of them (Mine #1) is a shallow longwall mine where the coal strength is very high, and the other mine (Mine #2) is extremely deep with a low strength of coal.

Jan 1, 1987

By Donald L. Everhart

Economic geologists faced with the task of evaluating phosphate rock deposits find it necessary to deal with a unique and a lengthy list of geologic, chemical, and economic factors before they can determine whether the deposit does, or does not, constitute "ore". The underlying fact is that phosphate rock concentrate sells at a price per ton that is among the lowest for all mineral commodities. A recent going price for a standard grade (i.e. 72% BPL) phosphate rock concentrate F.O.B. Tampa is $6.00 per long ton. This compares with sand and gravel's price of about $5.00 and $8.00, respectively, per ton. In the case of many, if not the great majority of mineral deposits, the evaluating geologist needs to concern himself principally with the fundamental and classic considerations of tonnage and grade. Barring some very unusual metallurgical problem any geologist that finds say 50 million tons of 1% copper rock, or substantial tonnages of 0.7% U308 sandstone, pretty well knows he has an ore body! Certainly the evaluation of phosphate rock starts with tonnage and grade. Any rock with less than about 10% P205 in the ground is generally ruled out as ore, on the basis of grade alone. Since nearly all phosphate rock operations must be of the large-tonnage, low-unit-value type, the tonnage of new reserves, to be considered as "ore", ought to be in the 100-million + class. But having outlined' a deposit of phosphate rock having acceptable

Jan 1, 1971

By Vincent J. Murphy

INTRODUCTION Geophysical applications to engineering projects are usually considered for reconnaissance exploration, and for some preliminary design phases. Recent experience indicates a growing need and a greater number of requests for measurements during construction. This is particularly true where seismic velocity values are related to strength properties of rock. Typical project objectives during construction are: dike and tunnel realignments required by surprise conditions encountered during foundation preparation; checks on grouting effectiveness which require velocity measurements before and after grouting; and dynamic moduli measurements along the walls of tunnels and deep excavations. This paper considers one of the on-site seismic applications that is rather simple and straightforward, and yet has few references in the literature. The specific objective is the measurement of seismic velocities (both "P" waves and "S" waves) for determining the thicknesses and dynamic moduli values of loosened zones of tunnel wall rock. PRE-EXCAVATION SEISMIC METHODS The seismic methods suitable for engineering project studies are well known and especially suited for routine site reconnaissance; these are illustrated in Fig. 1. The seismic refraction profile with a

Jan 1, 1997

By G. Newman, C. D. Sellars, E. A. Portfors

Power for mine operation in remote locations is usually provided by diesel generators. For gold mines where processing is normally desirable on-site, power requirements can be significant and the operating cost for diesel generation then becomes an important factor in mine feasibility. If a suitable site is available, a small hydroelectric power facility near the minesite can provide a significant portion of the mine power requirements. The hydrological, engineering and economic factors involved in small hydroelectric power development are described in this paper. Capital costs are minimized by run-of-river hydro development which does not require a dam and reservoir and has less environmental impact. During periods of low flow, diesel generators are used to make up the power demand. The hydrologic analysis required to determine unit energy costs is described and technical requirements for linking the hydroelectric power facility with diesel generation are outlined. Case studies are presented demonstrating the economic advantages of a small hydro installation. One case study is for a proposed gold mine in British Columbia and the second is a large operating gold mine overseas, for which the hydroelectric power scheme is presently under construction .

Jan 1, 1987

By M. J. Zdunczyk

Located approximately 50 miles northeast of Knoxville, Tennessee, Short Mountain Silica mines the Clinch Formation, an orthoquarzite, for silica sand. In 1980, a Knoxville resident and area contractor visualized mining Short Mountain for silica sand. Because APG Industries, the main market for the glass sand product, was only 48 kilometers from the site, many silica sand companies became interested in the property. In 1986, J.D. Nicewonder began to develop the site, and in May of 1988, APG Industries received the first sand shipment.

Jan 1, 1992

By H. E. Rudi

The advent of open nit mining permitted the development and processing of minerals never before thought possible. New equipment was needed; new techniques were devised; new methods were called for. This development and acquisition of new techniques, methods and equipment did not, however, stir the souls of our industry and educational leaders sufficiently to a recognition of the problem of people, especially the future development of the people needed in open-pit mining. The people problem has been and will be with us as long as there are two of them. The development of people will be required for the same reason. The problem of developing future personnel for open-pit mining is not what must be developed; that is known to all of us and can be briefly stated ashore and better of everything." As I see it, the real problem lies in determining (1) what personnel are really critical, (2) why are they critical and (3) how can we develop them.

Jan 1, 1965

By Alain A. Sonntag

The MBS Solvent Extraction System utilises a unique design of mixer- settler having very high capacity per unit area and great flexibility. The development of the mixer-settler is traced through its various stages wherein the independence of mixing and pumping, reduction in settler area requirements and simplicity of plant layout were achieved. The laboratory, pilot plant and full scale tests contributing to the development of the machine are described. The existing KREBS SX plants in Europe and Africa are described as well as the plants presently under design. Results of recent testwork are also given.

Jan 1, 1981

By R. H. Parker, B. A. Wills, D. G. Binns

Introduction With the increasing need to mine lower grade ores, high energy-related costs in comminution are of major concern. Very fine grinding is needed to liberate the fine mineral particles in these ores. Not only is this expensive, but it leads to greater losses in slimes. Gravity concentration techniques become unacceptably inefficient for particle treatment below 25 to 50 µm, and even flotation fails in the ultra-fine range. Ore from the South Crofty mine near Redruth, Cornwall, was subjected to thermal pretreatment in the hope that differential thermal expansion of the minerals would lead to intergranular cracking, and thus enhanced liberation. South Crofty ore contains about 1% tin as cassiterite associated with a complex assemblage of quartz, chlorite, tourmaline, and hema¬tite in granite and slate. Recovery of cassiterite can be as low as 70%, as overgrinding occurs during stage reduction of the material to 1 mm (primary grinding) and 180 µm (regrinding). Cassiterite is present in grain sizes ranging from submicroscopic to + 10 mm, the bulk being predominantly in the range of 50 µm to 3 mm. Experimental Flat polished sections of the ore were photographed in reflected light using a Vickers M17 microscope. The sections were then heated in a Carbolite LMF4 muffle furnace. The heating rate and temperature were monitored by a thermocouple immersed in the bed of particles. Previous work by Sherring (1981) on the same material showed that a 55% reduction in grinding resistance occurred when the material was rapidly heated and cooled through the quartz inversion tem¬perature (573°C), where a volumetric expansion of 0.86% occurs. The samples were therefore heated to 650°C at 26°/min, and were then water-quenched to room temperature. The effect of heat treatment on the mineralogy and fracture network was assessed by examining the same area after treatment. Results Most of the sections examined showed that extensive transgranular cracking occurred as a result of heat treatment. Such cracking, although weakening the rock and hence reducing the work index, would in no way enhance the liberation of the cassiterite from the host minerals. Intergranular cracking, which would lead to enhanced liberation, was difficult to discern under reflected light, but there was evidence of such cracking in some of the sections examined. Figures 1 and 2 are examples of typical sections before and after heat treatment. Figures la and lb - show cassiterite in hematite, Fig. lb, in normal incident illumination, illustrating clearly that after heat treatment, transgranular fracture is evident in the cassiterite. If any intergranular cracking has occurred, it is not evident in this photograph. Figures 2a and 2b show cassiterite in quartz before and after treatment. Fracturing in both minerals after treatment can clearly be seen, and there is evidence of intergranular fracturing. However, protruding "arms" of cassiterite are severed by transgranular fracture. Subgrains of quartz have also been isolated by trans¬granular fracture. There is major transgranular cracking across the wider sections of cassiterite. Discussion The effect of rapidly heating South Crofty tin ore to 650°C, followed by water-quenching to room temperature, has been studied by observing the fracture networks in mineral grains. The choice of 650°C as the heating temperature was influenced by the work of Sherring (1981), who found considerable reduction in grinding resistance after thermal pretreatment of the same ore. Scheding et al. (1981), however, have shown, by means of a crude calculation, that heat treatment cannot be justified solely on the basis of reduced grinding costs. The cost of heat treatment far out-weighs that of grinding, resulting in the combined costs for heat treatment being over six times that for grinding unheated material. The most economically attractive aspect of heat treatment is the possibility of enhanced liberation of the valuable mineral due to increased intergranular rather than transgranular fracture. Grinding costs would be greatly reduced by improved liberation at coarser sizes, and the costs of ancillary processes, such as dewatering and tailings disposal, would be reduced. The most significant economic effect, however, would be in improved metallurgical efficiency. Improved liberation would increase concentrate grades, and recoveries would be higher, particularly in the case of ores where high slime losses are produced due to the excessively fine grinding required to produce adequate liberation. Manser (1983) has shown that only a 1% increase in tin recovery, at the same concentrate grade, would be sufficient to offset heat treatment costs on South Crofty ore. Heavy liquid analysis and shaking table separation have been used to evaluate the effect of thermal treatment on processing (Binns, 1984; Scheding, 1981; Sherring, 1981). However, treatment of unheated and heated samples of South Crofty ore, ground to the same product size, by such methods have shown little evidence of improved metallurgical efficiency after thermal pretreatment. The large amount of transgranular fracturing revealed by reflected light studies helps to explain this lack of improvement. Nevertheless, from the evidence of some of the fracture work, it is surprising that no improvements in metallurgical

Jan 1, 1988

By M. Rieber

The potential for the wide-spread commercialization of superconductivity, the disappearance of all resistance to a direct current (D.C.) electric flow, began in 1986-87 with the development by Bednorz and Muller at IBM's Zurich Laboratory of an yttrium -barium -copper oxide conductor which attained zero resistance at 90t:. Although the status of superconductivity is still pre-bench scale, it is useful to forecast its probable effects on the copper industry. I t should not be forgotten that at their inception, the impacts of transistors, semi-conductors, thin films and fiber optics on the electric/electronics sector of this industry, were largely underestimated. Superconductors are not metals, but a class of ceramics. They represent an efficiency increase of one or more orders of magnitude. If their development is to be a boon to the copper industry, rather than a reduction of metal demand through miniaturization or material substitution, vastly expanded copper containing electrical uses must be in prospect. In this paper the projected uses of high temperature superconductivity are examined, followed by the remaining problems to be overcome. Assuming success, a time horizon for maturation is estimated. The penalties for the non-recognition of new markets and material substitution are shown on the basis of past forecasts. Finally, the application of successful superconductivity to the copper industry over the forecast period is estimated.

Jan 1, 1990