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By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

9.1 INTRODUCTION There are approximately 70,000 abandoned coal mines nation¬wide, which is about 35 times the number of underground coal mines presently operating. The US Bureau of Mines estimates that there are over 8 million acres of undermined land due to the extraction of coal, metals, and nonmetals. Subsidence has affected more than 2 million acres, and more than 99% of the subsidence is related to underground coal mining. There is reason to believe that some of the remaining 6 million acres of the undermined land have a high potential to subside. The expansion of housing, highways, commercial structures, and other facilities has required the use of many areas that are underlain by abandoned coal mines, and this growth will continue. Many subsidence problems are derived from the collapse of aban¬doned underground coal mines. There are numerous abandoned mine workings in the anthracite fields of northeastern Pennsylva¬nia, in the Appalachian bituminous fields, the Illinois Basin, the Rock Springs, Wyoming area, and other areas of the United States (Gray et al., 1976). Various room and pillar patterns of mining have been used in the dipping anthracite seams and the nearly flat-lying bituminous seams with considerable variation in the per¬centage of coal extracted. The progressive deterioration of pillars, mine floors, and mine roofs after long exposure to air and water may later result in the collapse of strata over the mine entries, the crushing of the remaining coal pillars, or the bearing failure of the mine floor beneath the coal pillars. Subsidence then results as the collapse reaches the ground surface in the form of differential strains, depressions, cracking of the ground, and sinkhole devel¬opment. Subsidence over active longwall mines, which occurs concurrently with mining or is completed within a short period following coal extraction, has been studied extensively over the past decade. On the other hand, subsidence over abandoned coal mines receives little attention by the researchers, mainly because it is difficult to predict and takes place decades after mining has ceased. The techniques of investigating the subsidence events over abandoned coal mines are similar to those employed for active mines except that at the outset, it is necessary to determine whether or not the subsidence events are mining-related (Chugh et al., 1986; Cummings and Singh, 1986; Peng and Hsiung, 1986). This calls for the identification and confirmation of abandoned mine workings under or near the affected surface structures. Gen¬erally old mine maps, if available, are acquired, the surface bore¬holes are drilled for confirmation of the accuracy of the mine maps and determination of the potential for continued subsidence. Sub¬sidence monuments are established and periodic surveys con¬ducted to determine the amount and trends of surface movement in and around the affected surface structures; surface boreholes are used to investigate the integrity of the underground structures (i.e., roof, coal pillars, and floor) by TV camera. They are also used for monitoring the vertical and horizontal movements of the subsur¬face strata by Sondex (FPBX) and inclinometer (PFBI), respec¬tively, for determining the continuity of the subsidence events. Tape extensometers, crackmeters, etc., are used to monitor the development trends of major cracks in the structures or on the ground. These data are used to identify the causes of the subsidence events. Finally, abatement methods are selected to stabilize the structures. It must be noted that most subsidence events over abandoned coal mines are reported after the fact, and investiga¬tions are begun some time after reporting, that the subsequent measured movements are generally much smaller than that at¬tained in the active mines, and that due to lack of knowledge about the damage conditions and the exact location of the abandoned mine workings with respect to the affected surface structures, the precise causes of surface subsidence or surface structural damages are in most cases very difficult to identify. 9.2 TYPES OF SURFACE SUBSIDENCE According to Gray et al. (1977), after examining 354 incidents of subsidence above abandoned mines in the Pittsburgh metropol¬itan area, the subsidence features have a mean diameter (i.e. the average of long and short dimensions) from less than 1 ft to 1600 ft, with 84% less than or equal to 15 ft; the subsidence features have a depth ranging from less than 1 ft to 48 ft, with 89% less than 25 ft; 66% of the subsidence features are deeper than they are broad. Nearly 59% of the subsidence features occur with overbur¬den less than 50 ft thick and 81% less than 100 ft thick. No subsidence features occur with overburden thicker than 450 ft (Fig. 9.1). Occurrence of subsidence incidents varies from imme¬diately to more than 100 years after mining (Fig. 9.2). In analyzing the characteristics of approximately 3000 chim¬ney subsidence features along the Colorado Front Range, Matheson and Eckert-Clift (1986) examined historical aerial photographs on 4- to 14-year intervals between 1937 and 1967 and found that the majority of observable surface subsidence features occurred within 30 to 40 years after mining. According to Gray et al. (1977), the most prevalent subsi¬dence features over abandoned mined land are sinkholes, with depths of more than 3 ft, and troughs or sags, usually less than 3 ft deep. Sinkholes are steep-sided pits, while troughs are shallow depressions much wider in area than sinkholes. A. SINKHOLE SUBSIDENCE A sinkhole is caused by the collapse of a mine roof that works its way upward. If it is not arrested during the process it will eventually reach the surface and emerge as a sinkhole. The process is governed by the thickness and character of the overburden, the width and height of the mine openings. Sinkholes are usually 3 to 20 ft deep and may be 2 to 40 ft in diameter, although most are fewer than 16 ft across (Gray et al., 1977; DuMontelle and Bauer, 1983). Newly formed sinkholes have steep sides with straight or bell-shaped walls. At times, they appear to be conical in profile with the apex upward. If the topsoil collapses, the top portion will widen to form an hour-glass shape. Sinkhole subsidence usually occurs over abandoned mines less than 165 ft deep (Hunt, 1979). Matheson and Eckert-Clift (1986) found that chimney sink¬holes are likely to occur when the ratio of overburden thickness to mining height (h/m) is less than 4 to 5. When h/m is between 5 and 10 to 11, the potential occurrence of chimney sinkholes decreases rapidly. When h/m is more than 10 to 11, less than 10% of the mine openings that collapse will induce sinkholes on the surface.

Jan 1, 1992

By Richard J. Ebens

The trace element geochemistry of the cherty, clay-rich carbonate residuum. Covering much of southern Missouri and adjacent parts of Arkansas was investigated during the early 1970's as part of a statewide geochemical survey of Missouri-1/2/3 The study was environmentally oriented and focused on the character of the "natural' geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1 000 parts per million (ppm) or more of Pb, Zn and Ba. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in figure 1 indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician Formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units--Bonneterre Dolomite, Potosi and Eminence Dolomites, Roubidoux Formation, and the Jefferson City, Powell and Cotter Dolomites. Six localities were selected randomly from over the Gasconade Dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background

Jan 1, 1977

By Chris A. Fleming, Daan Seymore

The Golden Jubilee Mine in South Africa is the first mine in the Western World to introduce resin-in­pulp technology for the recovery of gold from a cyanide-leached slurry. Located in the Eastern Transvaal region of the country, the plant processes approximately 375 metric tons• of ore per day, of which over 90 per cent is mined in an open pit adjacent to the metaflurgical plant, while the remaining 5 to 10 per cent is trucl<ed in from two small high-grade deposits, located about 10 kilometers from the plant. The gold grade in the open pit is about 1 git, while the values m the two high-grade deposits range between 5 and 15 git. At the current treatment rate, the proven reserves m the three locations will last about another 10 years. Early testwork indicated that the ore was not well suited to conventional filtration processing or to heap leaching because of its high cfay content, and that carbon-m-pulp (CIP) would probably prove to be the most viable metallurgical route. Tcstwork to establish the feasibility of CIP for this application was carried out at Mintek in 1985. This worl<, which was based on batch-leaching and carbon-loading tests, indicated that CIP was technically feasible, but that both the rate of gold loading and the gold-loading capacity of the carbon were inferior to those obtained with Witwatersrand ores. The poor adsorption characteristics of the carbon were attributed to the abnormally high concentration of natural organic compounds (humates and fulvates) in the ore. Batch tests were therefore also carried out with strong-base resins, since resins are known (1] to be far less susceptible to organic fouling than is activated carbon. The results were very promising, and pointed to fast kinetics and a high loading capacity. (Fig. 1 ). Notwithstanding these findings, it was decided to recover gold with activated carbon at Golden Jubilee Mine because the CIP process was, by then, well established in South Africa and elsewhere in the world, whereas resin-in-pulp (RIP) was unproven technology. The r>lant was commissioned in September 1986 and, for 18 months, operated in a carbon-in-leach (CIL) configuration, i.e. with a total of . six leach tanks, the last four of which contained activated carbon. As could have been predicted from the laboratory results, the performance of the CIL plant was generally poor, low gold loadings on the carbon (less than 1500 git) and high losses of soluble gold (0.2 to 0.3 p.p.m.) bemg characteristic of the operation. Consequently, it

Jan 1, 1990

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

8.1 INTRODUCTION Under room and pillar mining if the extraction ratio is low, (i.e., < 50%) a pillar can support the overburden without collapse and there will be no surface subsidence whatsoever. In other words, if the pillar sizes are properly determined to support the overburden weight, the roof strata cannot cave and there will be no surface subsidence, provided the entry width is small and properly reinforced so that a continuous roof fall leading to piping or the total collapse of the roof does not occur. Under the prevailing mining conditions in the US if only development mining is em¬ployed in room and pillar mining (i.e., no pillar extraction), the extraction ratio is usually less than 50% and generally no surface subsidence will occur. If pillar mining is employed, surface subsidence may occur, depending on the extraction ratio and the size of the remnant pillars (or stumps) left in the gob. The higher the extraction ratio, the smaller the stump pillars left behind in the gob. As the size of the stump pillars becomes smaller, the possibility of their support¬ing the roof without collapse is smaller. If all stumps are crushed completely, the roof will collapse and surface subsidence will occur. Again, whether surface subsidence occurs, it depends on whether the remaining stumps can support the overburden or not. Therefore, since the weight of the overburden which the stump pillars have to support is proportional to the mining depth, a fixed¬size stump pillar can support the overburden at a shallow depth, but will collapse under greater depths. In this context extraction ratio is not the only parameter for determining whether surface subsidence will occur. Furthermore, during pillar mining a pillar can be cut away in many ways to finally reach the same extraction ratio. Therefore, the safest way to determine whether or not sur¬face subsidence will occur in room and pillar mining is to compare the strength of all stump pillars in the panel of interest with the weight of the overburden that the pillars have to support. If they are too small to support the overburden, surface subsidence will occur. Under normal mining practices the characteristics of surface subsidence profiles vary with methods of mining. In the US there are two types of coal mining methods, i.e., longwall and room and pillar mining. Longwall mining is indisputably a total extraction method, but different percentages of coal recovery are achieved in room and pillar mining. Recovery is much higher for sections with pillar extraction than those without pillar extraction. For those without pillar extraction, coal recovery increases with increasing entry width or decreasing pillar size. For those with pillar extrac¬tion, the total coal recovery depends on the amount of coal ex¬tracted from each pillar which is in turn dependent on the roof conditions and methods of pillar extraction. Bauer and Hunt (1981) analyzed the characteristics of the subsidence profiles due to the three types of mining methods in Illinois. Fig. 8.1 shows the range of locations associated with the following profile features: profile edge, maximum tensile curvature, maximum slope, maxi¬mum compressive curvature, and maximum subsidence. The scat¬ter of the data increases with decreasing percentages of extraction. The profile is much sharper for longwall mining and becomes milder as percentage of extraction decreases; for longwall mining the profile edge is always beyond the opening, while in room and pillar mining it may fall within. The location of maximum tensile strain moves outward, while that of the maximum compressive strain moves inward as the percentage of extraction decreases and the maximum subsidence decreases with decreasing extraction ra¬tio. The general concepts discussed in this chapter may not apply to the entry development sections between two longwall panels where roof caving in both panels are very severe and overlap each other. Consequently, it must be treated differently. 8.2 CASE STUDIES Kohli et al. (1982) described several case studies on surface subsidence due to room and pillar mining with pillar extraction. In one case, the seam was 6 ft thick and ranged from 650 to 800 ft deep. The panel was 720 ft wide (Fig. 8.2) and was developed with seven entries. Each entry was 18 ft wide and the pillars were driven at 88 ft center-to-center between entries and crosscuts. Dur¬ing retreating, two additional pillars (pillars J and K) were devel¬oped, two blocks ahead of the pillar line, to take the barrier pillar, 160 ft wide, between panel No. I and No. 2. Pillar extraction began by taking a central slice across the whole panel, beginning from pillar 0 through J. After that, each pillar was sliced by 6 wing-slices, 3 on each side, leaving 4 stumps, 2 on each side. One of each pair was triangular shaped approximately 6 ft wide by 26 ft long. The other was trapezoidal, 26 ft long by 16 ft and 10 ft wide at the upper and lower bases, respectively. This resulted in a total extraction ratio of 85% within the panel. Obviously the stump pillars were not large enough to support the overburden and col¬lapsed completely. The final subsidence profiles were rather smooth, just like those under total extraction (Fig. 8.3). The max¬imum subsidence on Lines A and B was about 3 ft for an average mining height of 6 ft, but the locations of the maximum subsidence were skewed toward panel No. 1. Along Line D, the central por¬tion was flat-bottomed, but the maximum subsidence was only 2 ft, mainly because Line D was only two blocks from the right-side edge. Notice in every profile there was heaving of as much as 0.3 ft, beyond the edges above the solid coal. In another case (Fig. 8.4), the seam was 6 ft thick under 650 to 850 ft cover. Three panels, 710, 400, and 350 ft wide, respec¬tively, were extracted side by side. The entries were 18 ft wide and the pillars at 88 ft center-to-center between entries and crosscuts. Pillar extraction was practiced the same way as described previ¬ously. According to the mine maps, there was no barrier pillar left between the adjacent panels. Thus after extraction the three panels formed a common gob of more than 1460 ft wide. This was re¬flected in the development of the subsidence profile along the county road. After Panel No. I was mined, the maximum subsi¬dence of 1.2 ft occurred at Monument No. 1. This was increased to 1.6 ft after the extraction of Panel No. 2 and went to a further 1.8 ft after the extraction of Panel No. 3. The final maximum subsidence of 4.3 ft occurred at Monuments No. 6 through No. 8, while the remaining portion of the county road showed rebound. However it must be noted that the subsidence profiles were rather bumpy and fluctuated, perhaps an indication of the irregular stump pillars that were left and crushed at various time after mining. Moebs (1982) surveyed the subsidence profiles of four room

Jan 1, 1992

By H. Wagner

INTRODUCTION One of the most important tasks facing the opera¬tors of deep-level gold mines is supporting the roof strata in the immediate vicinity of the stope faces. Some of the deeper gold mines operate at depths ex¬ceeding 3.0 km; the extreme mining depths and the tabular nature of the gold-bearing reefs result in high stress concentrations at the abutments of the stope faces. These, in turn, lead to extensive fracturing of even the hardest quartzites. Although most of these fractures develop in a stable manner, there have been many instances where rock failure occurred suddenly, resulting in a considerable amount of damage to the underground workings. Such violent rock failures commonly are known as "rock bursts," and they are typical of many hard-rock mines that are operating at great depths. To safeguard mining personnel and equipment against the effects of stable and unstable rock failures, adequate support must be provided to the fractured roof as quickly and as effec¬tively as possible. Traditionally, timber props and timber packs have been utilized to provide roof support in gold mine stopes up to 1.5 m wide. Both methods of support suffer from the fundamental disadvantage of generating sup¬port only in reaction to movement of the strata. Since unrestrained strata movement often results in a break¬down of the frictional forces that preserve the stability of the intensely fractured roof strata, it is essential that stope supports have rapid load-bearing characteristics. An equally important requirement is that the roof sup¬ports be able to accommodate displacements of the rock mass. A third requirement of roof supports in gold mine stopes is that they must be able to accommodate the high rates of stope convergence typical of rock bursts. The hydraulic prop is the obvious choice for sup¬porting the stope faces in deep-level hard-rock mines. Since hydraulic props can be installed with a positive setting load, they can provide roof support that is independent of strata movement. By equipping the hydraulic props with yield valves, they meet the require¬ment of being able to accommodate rock-mass displace¬ments. However, the requirement of being able to accommodate high rates of stope convergence without overloading or damaging the rock strata was rather difficult to meet. The first introduction of hydraulic roof supports into gold mine stopes showed that conventional hy¬draulic props, such as those used in coal mines, were completely unsuitable for a number of reasons, in¬cluding: 1) Since drilling and blasting are used to break the ore in gold mine stopes, supports installed close enough to be effective are exposed to the vigor of the blast. Conventional coal-mining props suffered extensive blast damage, despite being protected by a blast barricade. 2) The environmental conditions in gold mine stopes favor corrosion. The high quartz content and the abrasiveness of the ore further aggravate the problem by damaging the protective coatings on the hydraulic surfaces. Simply changing the design of gold mine props to have the piston project from the bottom rather than from the top resulted in dramatic improvements in the service life of the props. The improvement was attributable to external water in the stope washing the fine quartz grit away from the hydraulic surfaces, rather than concentrating the particles between the cylinder and piston of the prop. 3) The most obvious deficiency of conventional coal mining props was their inability to yield rapidly enough to accommodate the high convergence rates of rock bursts without overloading the supports. As a result of the first disappointing results with hydraulic supports in narrow hard-rock stopes mined by drilling and blasting, a comprehensive research and development program was undertaken. The hydraulic props resulting from several years of research and exten¬sive field trials are described herein. To date (1978), more than 130,000 of these props have been manufac¬tured and used successfully in a number of mines and under a variety of geological conditions. Although many of the detailed specifications are directed spe¬cifically to gold mining conditions, there are a number of points that are of general importance in the design of underground support systems for tabular excavations in hard-rock mines of any type. SPECIFICATIONS FOR RAPID-YIELDING PROPS Yield Requirements The yield requirements for hydraulic props in deep¬level hard-rock gold mines are summarized in the fol¬lowing paragraphs. Slow Yield Load: The design of the yield valve must allow the yield load to be preset to between 300 and 400 kN. This flexibility is necessary to account for the differences in the strength properties of the roof and floor strata. Drop in Load Under Slow Yield: The drop in sup¬port load during a slow yield should be as small as possible, but it should not exceed 5% of the slow-yield load. Rapid-Yield Load: The load exerted by the prop during periods of rapid stope convergence should be as close as possible to the slow-yield load. The rapid-yield load should not exceed the slow-yield load by more than 15% at yield rates of 1.0 m/s. The limitation of the upper valve of the yield load is important; it deter¬mines the hydraulic performance of the yield valve. The hydraulic props that satisfy this criterion have performed well, even under severe rock bursts. Load Changes Under Rapid Load: On some of the earlier rapid-yielding props, conditions of dynamic in¬stability were observed; these conditions resulted in a complete breakdown of support resistance. It is essen¬tial that the force oscillations during rapid yielding be as small as possible, not exceeding -*5% of the average rapid-yield load at a yield rate of 1.0 m/s. Longitudinal Resilience: During rock bursts, vibra¬tions of the roof and floor sometimes are observed. To

Jan 1, 1982

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

10.1 I~RODUC1&apos;ION In order to understand the effects of surface subsidence due to underground mining on the ground water hydrology, a brief introduction to the rate, direction, and general pattern of ground water movement in aquifers is necessary. Ground water movement in an aquifer is controlled by aquifer lithology, structure, permeability, water levels, topographic set¬ting, and pumping patterns (Sgambat et al., 1980). Fig. 10.1 shows an idealized flow pattern of ground water in a hilly terrain (Coe and Stowe, 1984). The hill is made up of a series of interbedded permeable aquifers (mostly sandstone) and confining units (mostly shale). Since sandstone is porous and permeable, ground water moves through it downward and laterally until it reaches the interbeddings of the confining unit immediately below. The ground water then moves laterally along the interbeddings, be¬cause the confining unit is less permeable, and comes out as springs on the surface at the intersection of the aquifer and the confining unit. If there is a ground water table in each aquifer, it is a perched water table (e.g., right side of Fig 10.1). On the other hand, if the ground water recharge is larger than can be drained out, the interbedded permeable aquifers and confining units will be filled with water and a ground water table will exist only at the uppermost permeable aquifer. This is the semi-perched water table (e.g., left side of Fig. 10.1). There are three ground water flow systems: shallow, interme¬diate, and deep. Shallow flow systems occur near the land surface and the surface drainage system. The flow depth ranges from a few to several hundreds of feet with a resident time of a few weeks to a few years after entry, depending on the permeability of the rock strata. In a shallow flow system, the drainage coincides with sur¬face run-off water divides. Intermediate flow systems occur in minor hills some distance away from surface drainage. The travel distance ranges from several hundreds to several thousands of feet, the travel time from several months to tens of years. These systems are less dependent on seasonal recharge or discharge. In deep flow systems, the ground water travels from thousands of feet to miles, and transfers between surface basins. The travel time ranges from several tens to several hundreds of years. The systems generally receive recharge over a large area on major mountains and dis¬charge into major stream systems. 10.2 CASE STUDIES There are two ways of monitoring water wells for studying the effects of dewatering due to underground mining, i.e., water level and piezometric level. The water level is the water elevation in a water well that has penetrated two or more aquifers. Thus it indi¬cates the combined flow of all the aquifers that the water well intersects during the time of measurement. The piezometric level is the level to which water rises in a piezometer cased to a specific aquifer, and coincides with the pressure level of water in that aquifer. Therefore piezometric level reflects the water condition in that particular aquifer to which it is confined, while the water level represents the total yield of a water well without regard to the source of water. For instance, if a water well intersects two aqui¬fers, and if the lower aquifer is affected by undermining and goes dry while the upper aquifer does not, the water level in the well will drop, but may never go dry. Under this condition, the data obtained by monitoring the water level do not indicate the true behavior of the individual aquifers, especially if they are located at widely separated heights. 10.2.1 WATER LEVEL EFFECTS The effects of underground mining on ground water flow have been the subject of many investigations (Booth, 1984; Coe and Stowe, 1984; Owili-Eger, 1983; Rauch et al., 1984). However, the problems are very complicated, as have been exemplified by the conflicting conclusions reached by various investigators in var¬ious case studies. The most dramatic examples were described by Coe and Stowe (1984). In the first case (Fig. 10.2), the coal seam was 5 ft thick with cover from 400 to 800 ft deep. The majority of water supplies came from a sandstone aquifer in most ridge tops. When the face of the Panel No. 3 passed under it, dug well WW-1, which was approximately 670 ft above the coal, went dry, while the water level in Borehole B- 1, which was located at the center of Panel No. 4 with the bottom reaching the coal seam, also dropped about 160 ft. In Panel No. 5, the developed spring DS-2, which originated from the surface of the sandstone/shale contact and was approximately 750 ft above the coal seam, stopped when the face passed under it. The water level in the dug well W-2, which was dug into the shale about 700 ft above the coal seam dropped about 22 ft, but the well did not go dry. Therefore, nearly every spring or well located above or near the active panels was affected by longwall mining. In the second mine (Fig. 10.3), the seam was 5 ft thick with cover from 200 to 400 ft thick. Dug well DW-2, the bottom of which was 35 ft deep and approximately 330 ft above the coal seam in a perched sandstone aquifer, went dry right after the longwall passed under it. Drilled well W-3, in the sandstone aquifer, about 220 ft above the coal seam, experienced a 75-ft drop in water level when the face was under it, but showed some re¬covery after the face had passed under it. Ponds P- I and P-2 were developed in shale bedrock and were spring fed. They were not affected by longwall mining. Pond P-3 was developed in the sand¬stone aquifer, and its water table showed some decline when the face passed under it. Spring ST-1, which originated from Pond P-1 and was fed from several tributary valleys, was affected by longwall mining, probably because it passed across the panel edges where tension cracks drained off the water. Although the overbur¬den in case 2 was about half of that in case 1, the majority of water resources were not affected by longwall mining. Coe and Stowe (1984) attributed this discrepancy to: (1) a higher percentage of clays, shales, and claystones, resulting in a lesser degree of inter¬connected fracturing, and (2) a less rugged topography resulting in a less localized ground water flow system. Therefore, in order to assess the effects of underground mining, the premining ground water flow patterns, topographical, lithological, and meterological conditions must be known. Booth (1984) also monitored the ground water effects of a longwall mine (Fig. 10.4) in a seam 5 ft thick and 300 to 800 ft deep. There were three major aquifers: Morgantown sandstone approximately 30 ft thick and 630 ft above the coal; Saltsburg sandstone of up to 50 ft in thickness and 520 ft above the coal; and

Jan 1, 1992

By Garry Keizer

The Cortez gold mine is located in Lander County, NR, about 130 km (80 miles) southwest of Elko. Cortez Gold Mines (CGM) is a joint venture between Placer Dome US Inc. (60% operator) and Kennecott Corp. (40%). The claims that supply oxide and refractory ore feed to the Cortez milling facility are located in both Lander and Eureka counties. These include the Cortez and Horse Canyon deposits, all operated by the joint venture. Other deposits outside the CGM claim area include Fire Creek, Buckhorn, Gold Bar and Tonkin Springs. These deposits all occur along the Battle Mountain-Eureka Trend. History Mining in the Cortez district began in 1862 with the discovery of high-grade silver ores. Grades in the hundreds of ounces per ton occurred in an area 7 km (4.5 miles) southeast of the current CGM mill site. Underground mining through the Garrison and Arctic tunnels provided the bulk of the ore production between 1864 and 1895. To a lesser extent, the Mill Canyon area also supplied ore until 1930.

Jan 1, 1993

"There is no fast way to add more than 6.4 km (4 miles) of track, most of it underground, to the Sound Transit Link Light Rail system in the Seattle, WA area. But increasingly busy traffic through a number of neighborhoods made it necessary. That’s the challenge a joint venture of several tunneling contractors faced when starting its portion of the $1.9-billion Northgate Link Extension in 2013, and the project is still keeping it busy three years later. Completion will mean just a 14-minute ride from Northgate to downtown for an extra 60,000 passengers by 2030. The transit lines will join 57 km (36 miles) of new and under-construction light rail lines running north, east and south from Seattle.Voters tired of long commute times approved the project, a key part of Seattle’s regional mass transit system, in 2008. The extension is part of a plan adopted by the Sound Transit board in 1996 to connect the region’s major activity centers. The Northgate project is meant to relieve one of the area’s most congested traffic areas by connecting the Northgate, Roosevelt and University District neighborhoods to existing routes including downtown Seattle and SeaTac airport.Sound Transit dedicated more than a quarter of the project’s estimated expense — $440 million -— to general contractor JCM Northlink LLC to handle the tunnel segment. JCM is a joint venture formed in 2013 between Jay Dee Contractors of Livonia, MI, Frank Coluccio Construction of Seattle, WA and Michels Corp. of Brownsville, WI. The contractors regularly collaborate on large-scale projects and brought experience working on a different section of the extension they completed in 2013.JCM project leaders knew immediately that the schedule would be demanding. They had to work around the clock six days a week to meet deadlines. The extension runs through highly populated areas, meaning contractors had to follow strict city noise restrictions between 10 p.m. and 7 a.m. Plus, supervisors needed to manage the scheduling of nearly 250 employees to keep work running smoothly.“We had eight different operations and four different subcontractors trying to work in one tunnel,” said Chris Wood, JCM site project superintendent. “It’s important that we keep work linear. That’s the biggest challenge: keeping a path in and out so you can pour concrete, keep electricians and carpenters going, and accomplish everything else.”"

Dec 1, 2016

By J. Drew

Oregon Resources Corporation (ORC), an industrial minerals mining company, is currently engaged in the development of a new process design that will allow the unique paleo-beach placer deposits of southwestern Oregon to be mined efficiently and economically, creating an additional domestic source for chromite, garnet, and zircon. Engineering design has been guided by the variable geology and mineralogy of the paleo-beach placer deposits as well as the need for a dry tailing scheme that resolves a lack of water resources at the placer locations and at the same time eliminates the need for slurry settling ponds, typical of paleo-beach placer operations in North America. Metallurgical study of the placer material was grouped into four distinct samples based on marine terrace deposition, geological facies, and mineralogy. Because the metallurgical samples represented the extremes likely to be encountered in all future Oregon paleo-beach placer ores, the process design is highly dynamic and will successfully adjust to meet the production needs. Water availability at the mining area is seasonal and will not support a typical placer operation?s water requirement needs for heavy mineral concentration. For this reason, ORC has developed a plan to construct the ore processing facilities near Coos Bay, where a municipal source of water is available. Raw ore will be transported from the mining sites approximately twenty miles one way to the processing facility with return loads hauling tailings back to the active pit. The need to haul dry tailings, return dry material to the pit for reclamation, and limit the amount of water being purchased from the municipality has driven the design of a unique water reclamation system. BACKGROUND Location and Access Economic concentrations of ?black sand? or heavy mineral (minerals with specific gravity greater than 2.85) have been recognized and studied in marine placers from Coos Bay to the mouth of the Rogue River, a distance of approximately 75 miles along the southern Oregon coast (Hornor, 1918; Griggs, 1945)(Figure 1). ORC will begin mining existing reserves approximately 20 miles south of Coos Bay in a region known locally as Seven Devils. Ore will be trucked north to the processing site near Coos Bay via use of existing county and state roads, including U.S. 101. The available facilities at the processing site include highway, rail, municipal water and electricity, natural gas, and a deep water port. At the time of this writing, the rail line had been abandoned, but was being pursued by the International Port of Coos Bay. It is anticipated that this will be serviceable at some time in the future. History The southern Oregon marine placers have garnered the interests of miners since 1852, when present day beaches were exploited for gold. The beach deposits were small and irregular in nature and were easily washed away by the major storms the coast endures during the winter months (Hornor, 1918). [ ] In the 1920?s, deposits at the beach were followed upstream to their paleo-beach terrace origins (Pardee, 1934). These terrace placers were mined, but with little success, as the cost of mining and processing was greater than at the present day beach deposits. The greatest effort to understand and delineate the paleo-beach terrace placers came during World War II. As the need for a domestic source of steel hardening chromite was evident, the heavy mineral bearing placers of southwestern Oregon provided just that. The U.S. Geological Survey, under guidance from the Oregon Department of Geology and Mineral Industries, began exploration drilling of the paleo-beach terraces in 1940 (Griggs, 1945). This work was part of the investigation of strategic mineral deposits and would ultimately supply much needed chromite for the war efforts. The first mining efforts began in 1943 by the Humphreys Gold Corporation and the Krome Corporation (Griggs, 1945). Black sand

Jan 1, 2009

"There’s no fast way to add more than four miles of track, most of it underground, to the Sound Transit Link Light Rail system in the Seattle area. But increasingly busy traffic through a number of neighborhoods made it necessary. That’s the challenge a joint venture of several tunneling contractors faced when starting its portion of the $1.9 billion Northgate Link Extension in 2013, and the project is still keeping them busy three years later. Completion will mean just a 14-minute ride from Northgate to downtown for an extra 60,000 passengers by 2030. The transit lines will join 58 km (36 miles) of new and under-construction light rail lines running north, east and south from Seattle.Voters, tired of long commute times, approved the project, a key part of Seattle’s regional mass transit system, in 2008. The extension is part of a plan adopted by the Sound Transit board in 1996 to connect the region’s major activity centers. The Northgate project is meant to relieve one of the area’s most congested traffic areas by connecting the Northgate, Roosevelt and University District neighborhoods to existing routes including downtown Seattle and SeaTac airport.Sound Transit dedicated more than a quarter of the project’s estimated expense — $440 million -— to general contractor JCM Northlink LLC to handle the tunnel segment. JCM is a joint venture formed in 2013 between Jay Dee Contractors of Livonia, MI, Frank Coluccio Construction of Seattle, WA and Michels Corp. of Brownsville, WI. The contractors regularly collaborate on large-scale projects and brought experience working on a different section of the extension they completed in 2013.JCM project leaders knew immediately that the schedule would be demanding. They had to work around the clock, six days a week to meet deadlines. The extension runs through highly populated areas, meaning contractors had to follow strict city noise restrictions between 10 p.m. and 7 a.m. Plus, supervisors needed to manage the scheduling of nearly 250 employees to keep work running smoothly.“We had eight different operations and four different subcontractors trying to work in one tunnel,” said Chris Wood, JCM site project superintendent. “It’s important that we keep work linear. That’s the biggest challenge: keeping a path in and out so you can pour concrete, keep electricians and carpenters going, and accomplish everything else.”"

Jan 6, 2018

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a &apos;normal" weekend with the level during long-term shutdown in the old &apos;Casino" of the papermill of Niederschlema. This papermill as well as the &apos;Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein &apos;Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel &apos;Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a &apos;normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

By W. J. Schlitt

Introduction Heap and dump leaching are low-cost metal production techniques that are gaining in popularity among gold and copper producers. However, the flow of solution in heaps and dumps has received little attention in the literature. This is unfortunate since solution flow is one of the few parameters subject to operator control. Thus, solution management may well influence both operating costs and plant performance. Costs aside, there are two important aspects of solution flow to be considered -the metallurgical and the hydrological. Some of the metallurgical factors have recently been discussed (Jackson and Ream, 1980; Schlitt, 1984; Schlitt and Nicolai, 1987). These cover solution application rates and methods, irrigation rates, leach-rest cycles, and nonvertical solution flow. According to Caldwell and Moss (1985), however, the hydrological aspects may be more significant than the metallurgical ones. In particular, a phreatic surface will form when excessive flow leads to an accumulation of water and the associated buildup of pore pressure within the rock pile. Such internal flooding can either originate at the foundation or at some other zone of very low permeability. The latter condition gives rise to an impounded volume of solution, i.e., a perched water table (see Whiting, 1985). Caldwell and Moss point out that a rising zone of saturation is probably the most common cause of dump failures. Of course, such failures have even occurred in heaps that were carefully prepared for leaching (Milligan and Engelhardt, 1984). Thus, flooded conditions and perched water tables represent an important safety consideration as well as having an impact on metallurgical performance. The following sections describe a case history in which a perched water table developed within a copper leach dump. The description includes background information, solution flow rates, and metallurgical data. Then this situation is compared to one involving normal drainage. Description of the leach system The leach dump in the case history is located at a shutdown open pit copper mine. It was built in two lifts, with the second added some 20 years after the first. There is little detailed information available on the initial lift. It was built with rail¬hauled waste and was generally less than 30 m (100 ft) high at the crest. The area available for leaching was about 120 m (400 ft) wide and more than 380 m (1250 ft) long. The dump surface was prepared for leaching by dozing ponds approximately 12 m by 12 m by 2.5 to 3.0 m deep (40 ft by 40 ft by 8 to 10 ft). The ponds were leached by flooding with barren leach solution returned from a scrap iron cementation plant. Based on mill feed at the time, the average waste grade was probably close to 0.4% Cu. The first lift leached well. It accepted high flows, which together with the waste grade, produced a rich pregnant leach solution (PLS). Old records indicate a PLS of "50 lb Cu/ 1000 gal," or about 6 g/L. As later drilling would indicate, such a high tenor led to dissolution of considerable scrap iron that was returned to the dump. The iron then hydrolyzed and settled out on the pond bottoms. In addition, the waste settled substantially so that the unleached crest was 3.0 to 4.5 m (10 to 15 ft) above the ponded area. Eventually, the leachable copper was extracted and the dump became less permeable. Thus, the PLS tenor dropped until it became uneconomical. About ten years later, mining resumed and the decision was made to add another lift to the dump. This was done without giving much thought to a subsequent leach operation. Hence a 24-m (80-ft) lift was built on top of the original dump. The surface of the latter was not prepared in any way, e.g., by leveling and/or deep ripping, prior to over-dumping. Examination of subsequent drill cuttings indicated that the new lift contained about 0.2% Cu, with chalcopyrite and chalcocite (50:50) being the predominant copper minerals. Most of the chalcocite occurred as rimming on the abundant pyrite, with the pyrite to chalcopyrite ratio estimated at 10 to 1. The use of 100- and 120-ton trucks for haulage caused some waste compaction during emplacement. In addition, the host rock itself was relatively soft, being a porphyry intrusive material that was partially altered to clay. As a result of the initial compaction and clay swelling, the rate of water percolation from the new leach ponds was slow and the ponds often contained considerable standing water. Even frequent ripping failed to provide a sustained improvement in the percolation rate. The poor surface permeability was exacerbated by hydrolysis of iron salts which settled as a layer on the pond bottoms. Partly as a result of the permeability problems, metallurgical performance was not up to expectations. These had been based on laboratory tests which showed about 20% copper solubilization in two weeks. Continued copper extraction in the tests also suggested that a substantial percentage of the copper would eventually be recovered. However, the PLS grade in the actual operation peaked briefly at about 0.48 g/L Cu (4 lb Cu/1000 gal), then declined to a range of only 0.24 to 0.36 g/L Cu (2 to 3 lb Cu/1000 gal). The poor leaching was traced to a lack of oxidation of the sulfides. There were two principal observations supporting this conclusion. First, there was no evidence of any heat being generated within the dump. As discussed elsewhere (Schlitt and Jackson, 1981; Hiskey and Schlitt, 1982), pyrite oxidation is quite exothermic and the high pyrite content of the waste should have led to an increase in the temperature of the leach solution as it percolated through the rock pile. Second, there was no sign of any natural convective air flow through the dump.

Jan 1, 1987

By Derek. C. Arnold

In order to meet increasing demands for water supply in the next millennium, the Government of the Hong Kong Special Administrative Region have identified the need for a major new water treatment and transfer scheme. The project now under construction will comprise a 1.2 km long raw water aqueduct, a major new 0.25million cubic metres per day water treatment works (with space for upgrading to 1.2Mm³/d), a 12.1 km long treated water aqueduct and a primary service reservoir. In order to meet increasing demands for water supply in the next millennium, the Government of the Hong Kong Special Administrative Region have identified the need for a major new water treatment and transfer scheme. The project now under construction will comprise a 1.2 km long raw water aqueduct, a major new 0.25million cubic metres per day water treatment works (with space for upgrading to 1.2Mm³/d), a 12.1 km long treated water aqueduct and a primary service reservoir.

Jan 1, 1999

By M. Plesha, A. Zubelewicz, K. Connor, T. B. Belytschko, O&apos

INTRODUCTION The Northwestern University Rigid Block Model (NURBM) has been extended to explicitly model dilatant behavior of rock joints. By incorporating dilatancy, the resistance to shear displacement and slip, which previously was limited to simple Coulomb friction, along edge-to- edge contacts is now provided by contact roughness as well as frictional resistance. The explicit formulation is based on concepts presented by Plesha (1985) in which a plasticity-based phenomenological approach was used to model dilatant joint behavior. This approach has been modified to allow for joint closure upon reversal of shear deformation. After presenting a brief description of the new joint constitutive relationship, model behavior will be illustrated with two examples. The first involves a simple two-block model to simulate a cyclic direct shear test of a single joint. The second example involves simulation of a large cavern in a moderately jointed rock mass subjected to cyclic excitation. It will be shown that in- creasing joint roughness increases cavern stability by interlocking of adjacent blocks. JOINT CONSTITUTIVE RELATIONSHIP Within NURBM adjacent two-dimensional rigid blocks interact along their common edge-to-edge contact with contact forces transmitted at the contact midpoint. The contact constitutive relationship was originally restricted to a Coulomb friction yield criterion with no tensile stresses permitted along the smooth contact surfaces. The local coordinate system for any contact is described as shown in Figure 1 with the elastic stresses and displacements related by the linear- elastic constitutive law: [ ] The contact microstructure has been modeled as Patton-type saw-tooth surfaces such as shown in Figure 2. Sliding can be active along the right asperity with angle aR and length lR or along the left asperity with angle a and length lL. In general, aR and aL will difkr as will lR and lL but for purposes of this paper it has been assumed that these angles and lengths are identical. By a simple transformation, the microscopic asperity stresses can be expressed in terms of the macroscopic contact stresses [ ] If Coulomb friction is assumed along the asperity surfaces, the plastic slip condition will be [ ] where µ is the friction coefficient with the convention that compressive stress is negative. When F is negative, the behavior is elastic and when F is zero, slip can ensure. When plastic slipping does occur, it is assumed that the relative tangential and normal displacements between the two contacting rigid blocks are additively composed of recoverable elastic parts and nonrecoverale plastic parts, [ ]

Jan 1, 1986

By J. Wagner, P. Eger, G. Melchert

In 1991, LTV Steel Mining Company decided to close an open pit taconite mine in northeastern Minnesota, using a passive treatment approach consisting of limiting infiltration into the stockpiles and wetland treatment to remove metals. Over 50 million metric tons of sulfide-containing waste had been stockpiled adjacent to the mine during its 30 years of operation. Drainage from the stockpiles contained elevated levels of copper, nickel, cobalt and zinc. Nickel is the major trace metal in the drainages, and prior to closure annual median concentrations ranged from 1.5-50 mg/L. Copper, cobalt and zinc are also present but are generally less than 5% of the nickel values. Median pH ranges from 5.0 to 7.5, but most of the stockpile drainages have pH greater than 6.5. Based on the chemical composition of each stockpile, a cover material was selected. The higher the potential a stockpile had to produce acid drainage, the lower the permeability of the capping material required. Covers ranged from overburden soil removed at the mine to a flexible plastic liner. Predictions of the reduction in infiltration ranged from 40% for the native soil to over 90% for the plastic liner. Five constructed wetlands have been installed since 1992, and have removed 60-90% of the nickel in the drainages. Total capital costs for all the infiltration reduction and wetlands exceeded $6.5 million, but maintenance costs are less than 1% of those for an active treatment plant. Since mine drainage problems can continue for over a hundred years, the lower annual operating costs should pay for the construction of the wetland treatment systems within seven years.

Jan 1, 1999

By Morris Beattie

Arseno Processing Ltd. has developed process technology for the extraction of gold from refractory ores and concentrates. These ores, which do not respond to conventional extraction processes, are rapidly becoming the primary targets for mining companies wishing to become significant gold producers. It has recently been estimated that by the mid 1990's, the annual gold production from the Pacific Basin Epithermal Arc in southeast Asia could be as high as 550 tonnes. Within this area the Porgera deposit under development by Placer, Renison and MIM, and the Lihir Island deposit under development by Kennecott have been shown to be refractory and require pretreatment before the gold can be extracted. The additional deposits which are now starting to be developed in this area are also expected to require such pretreatment. The Arseno process promises to be an economical pretreatment for these ores and to result in the maximum achievable gold recovery. Numerous refractory deposits are also known in Australia, North America, China and other parts of the world. Samples from many deposits throughout the world have been tested using the Arseno process conditions and the consistently successful results have demonstrated its widespread application. The samples which have been tested include ores in which the gold is locked in pyrite, marcasite or arsenopyrite. The process is insensitive to the sulphur content of the feed. Feed materials containing from 1 % to 50% sulphur have been tested successfully with the process. In the Arseno refractory gold process, sulphide minerals are dissolved so that the contained gold becomes accessible for subs7qu7nt gold recovery. This gold recovery is achieved by cyamdat1on of the process residue. For many of the materials which have been tested, gold extractions in excess of 95% have been achieved.

Jan 1, 1998

By Mark L. Munsell

In recent years, miners have tried many approaches to maintenance such as: - Highly sophisticated computerized programs. - large planning groups designed to preplan and schedule work. - Contract maintenance. - Break down maintenance. Each of these approaches has faced the added constraints: - High replacement cost of equipment necessitating longer useful equipment life - management?s increased scrutiny of all indirect costs such as maintenance. Faced with these factors, what system of maintenance control is right? After visiting many properties recently, it appears there is no right system. Maintenance programs need to be tailored to the people and production constraints that occur at an individual property.

Jan 1, 1976

By G. Bierie

The conveying of bulk materials is prone to problems. These problems can reduce the capacity and efficiency of the plant, and even threaten the health and safety of the workforce. This presentation will look at new conveyor technologies systems, and how, used separately or in combination, they can improve material handling.

Jan 1, 2007

By Reinhold Schulze

Thiourea is being applied as an industrial leaching agent for gold and silver only in a few cases. A CIP technique was developed which eliminates most of the obstacles against its use. Mild conditions for precious metals stripping from the loaded carbon simplifies process conditions in comparison to cyanide CIP technique. Examples are given.

Jan 1, 1998

By Otis M. Clarke

Although most bauxite mined is used to make metallic alu¬minum, more than a million tons of bauxite is used annually in the United States in the manufacture of refractories, chemicals, abrasives, and cement. Raw material requirements are becoming increasingly difficult to meet because high-grade bauxite, required for some uses, is limited. The supply of raw material for high-alumina refractories is most critical because a low-iron bauxite is needed. Although there are substantial bauxite reserves in the world, bauxite suitable for use in refractories is limited to a few districts.

Jan 1, 1977