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By Kumara Rachamalla

North American mining companies are lagging behind their global competitors in participating in the outstanding opportunities in India. The Indian government has liberalized foreign equity participation in the mining sector by up to 50% and, in some cases, even higher. Delegates from Europe, North America and South Africa learned this at an information seminar held in London, England, Attendees were welcomed by L.M. Singhvi, the UK's high commissioner for India. He introduced a government of India delegation headed by B.P. Baishya, minister of steel and mines. Singhvi is an eminent jurist and leading constitutional expert. He reiterated the soundness of India's legal system. He also outlined the recent Investment Protection Treaty between India and the United Kingdom. Baishya emphasized thee geological diversity and strengths of India's domestic market with its population of more than 920 million people the second largest in the world after China and its reservoir of skilled labor. He also outlined the potential of India's untapped natural resources. The private sector is the backbone of the Indian economy. It accounts for 75% of gross domestic product (GDP). The current minimum program of the new United Front government envisions 12% growth in the industrial sector, 7% in GDP and direct foreign investment of US$10 billion a year. "Mining is an area that can attract a sizable part of this investment," Baishya said. "Projected growth of the Indian economy will require increasingly large quantities of basic raw materials, such as coal and base- and precious-metals to meet the needs of domestic and export markets." Administration of India's mining sector is divided into the Ministry or Mines for regulating and developing the country's mineral resources, five public sector Mining Enterprises, the Geological-Survey of India (GSI), the Indian Bureau of Mines (IBM)and 25 states and seven Union Territories. The GSI is the second oldest (founded in 1851) and the third largest organization of its kind in the world, Baishya said. It has geologically mapped more than 90% of India's 3.2 million kmz (1.2 million sq miles) at a scale of 1:50,000. Several promising mineral projects have emerged from regional exploration programs conducted by GSI and the Mines and Geology State Governments. IBM recently completed a national mineral inventory. It covers 13,000 deposits/prospects of 61 nonferrous minerals. GSI also compiled a similar inventory on 61 coal fields. India is attractive to exploration companies for several reasons. These include favorable geology, accessible locations and a large mineral database. India also has many experienced geoscientists with well-equipped and efficient laboratories, Baishya said. Secretary to the Ministry of Mines A.C. Sen emphasized the largely untapped-geological and mining potential of India. He also discussed the new vistas that have opened up opportunities for exploration and mining. India has large quantities of mineral reserves, Sen said. Its vast Precambrian Shield - like those in Canada and Australia - is endowed with gold, platinum group and base metals, as well as coal and industrial minerals. Annual mineral production is valued at more than US$7 billion. Sen pointed out that India is the largest single consumer of gold. And domestic gold prices command at least a 20% premium above international prices. Recent diamond, gold and base-metal discoveries and prospects uncovered by GSI have generated investment interest from abroad, he added Delegates heard that the Indian Constitution gives the central government the job of framing legislation and the regulation and development of minerals. This ensures that mineral laws are uniform throughout the country. However, the right to grant mineral concessions, such as prospecting licenses and mining leases, rests with the minerals' owner. In India's case, that is the state government. The Indian government has formulated several guidelines that regulate the granting of prospecting licenses for large areas. ? The central government will consider the requests of state governments for the granting of prospecting licenses for areas exceeding 25 kmz (9.6 sq miles). But the license must include a provision to conduct aerial prospecting of the area. ? Any prospecting licensing area should not exceed 5,000 kmz (1,930 sq miles). for a single license. And the total area held by one company should not exceed 10,000 km2 (3,861 sq miles) for the whole country. ? The grant of larger areas will be linked to a mini- mum expenditure commitment on physical targets. State governments will monitor these expenditures. ? The granting of large areas for prospecting will be linked to a schedule of relinquishment.

Jan 1, 1997

By R. A. Holmes

Although manganese is a metallic element and is widely dispersed in nature, it never occurs except as a compound in combi¬nation with other elements. Use of such compounds in the production of glass is known to have occurred in early Egypt. The dioxide of manganese was considered a compound of iron until 1774 when C.W. Schule first recognized it as an element. In the same year, a Swedish mining engineer, J.G. Gahn, became the first to isolate manganese. In 1856 development of the Bessemer process of steelmaking gave economic importance to manganese, and later, in 1882 Robert Hadfield discovered the benefits of high manganese steels. GEOLOGY Physical Properties Elemental manganese is a silver-gray metal, resembling iron but harder and more brittle and used primarily in alloys (both ferrous and non-ferrous) and a wide variety of chemical compounds. Some of the physical properties of manganese include: melting point¬1 245°C; boiling point-2 150°C; density at 20°C-7.43 g/cm3; specific heat at 25.2°C-O.115 cal/g; latent heat of fusion-63.7 cal/g; hardness on Mohs scale-5.0; linear coefficient of thermal expansion from 0 to 100°C-22 x 10-6. Mineralogy There are over one hundred minerals that contain manganese. These minerals vary from those with compositions that are pre¬dominantly manganese to those having only minor percentages of manganese. Distribution of Deposits Manganese ore deposits are found worldwide and were formed in various geological environments, but only a rather limited num¬ber of deposits have high grade manganese ore in sufficient quan¬tities to be mined and utilized economically on an industrial scale. It is worth noting the fact that almost all of the significant deposits can be classified into two types of deposits: marine chemical sed¬imentary deposits, and residual (secondary) enrichment deposits. There are, however, a much larger number of geological types but these are not of commercial significance at this time. Sedimentary deposits are the most common and are usually stratiform or lenticular. Manganese minerals were formed by a chemical process during the deposition of marine sediments. They usually contain manganese oxides and carbonate minerals, some¬times interbedded together or with other sedimentary rocks such as limestone or shale. Examples of this type of deposit are the Russian ore bodies of Nikopol and Tchiatoura, as well as the Kalahari deposits in South Africa and deposits of Groote Elyandt in Aus¬tralia. Residual deposits were formed in a different way: by alteration of existing manganese deposits or by concentration of the manga¬nese minerals when other minerals were washed away by weath¬ering or ground water processes. The Nsuta deposit in Ghana, the Amapa deposit in Brazil, the Moanda deposit in Gabon, and nodules in the residual clays of the US Southern Appalachians are examples of this type of geological process. In the case of the Ghana and the Amapa deposits, this is only true for the outer layers of the deposit containing oxide minerals, the inner part being comprised of car¬bonate minerals including manganese carbonate, probably from marine origin. Some sedimentary and residual-type deposits have been metamorphosed, giving rise to small high grade ore bodies. These deposits are regionally metamorphosed, occurring in mar¬bles, slates, quartzites, schists, and gneisses. Some of these deposits, such as the Franklin, NJ, deposit are rich enough to be commercial without secondary enrichment; however, most of the exploitable deposits have been secondarily enriched. Due to the diversity and complexity of manganese deposits, both with respect to deposition and chemistry, a wide range of impurities are almost invariably present in the ores. Table 1 shows that the reserves (i.e., a measured resource that can be economically and legally extracted) are estimated at 814 x 106 tons of contained Mn, which equates to more than a 100 year supply at the current level of production. The reserve base is made up of the marginally economic reserves and sub-economic re¬sources and are some 4.5 times greater than the proven reserves. Also apparent from study of Table I is the fact that some 75% of manganese reserves are found in two countries, the USSR and South Africa. On the other hand, North America (i.e., the United States and Canada) have few significant deposits. The largest deposit, or at least one of the largest, in the United States is located at Chamberlain, SD. This deposit is sedimentary

Jan 1, 1994

By Bobby C. Carson

INTRODUCTION The National Park Service is entering the seventh year of monitoring caves for the presence of radon and radon daughter products. The purpose of this paper is to summarize the radiation monitoring program at Mammoth Cave National Park, and to present some of the results of this program. Mammoth Cave National Park completed five years of collecting data on May 1, 1981: although Mammoth Cave encompasses approximately 361 km of underground passageways, this paper will concentrate on only a 2.2 km section of the cave known as the Historic Tour. Included in this paper is a discussion of the methods the Nations Park Service uses to protect employees from exposure to alpha radiation. MONITORING METHODS The National Park Service monitors cave atmospheres utilizing the procedures provided by the Mine Safety and Health Administration in their Radiation Monitoring Training Manual (Anon., 1976). This procedure is described as the Kusnetz Method (Kusnetz, 1956) of radon daughter monitoring. Due to the length of the tours at Mammoth Cave, it has been determined to be the most practical procedure. The Historic Tour is a 2.2 km (1.4 mile) loop through passageways ranging in size from 18 m high by 12 m wide, to 0.9 m high by 0.6 m wide. Seven five minute walking samples were taken for this cave tour by drawing at least 10 1 of air through a 25 mm fiberglass filter utilizing a Monitaire Sampler Pump. The radon daughter concentration levels were determined using an alpha scintillation counter to measure the alpha activity on the filter paper. The Monitaire Sampler Pump was calibrated each day prior to monitoring the cave tour and the scintillation counter was calibrated by procedures described by the Mine Safety and Health Administration (Beckman, 1975) at six month intervals. Guidelines established by the National Park Service and approved by the Mine Safety and Health Administration require weekly sampling when the average working level exceeds 0.30 (NPS-14, 1980). A working level is an atmospheric concentration of radon (Rn-222) daughters which will deliver 1.3 x 10 5 MeV of alpha energy per liter of air in decaying through Ra C' (Po-214). The Historic Tour has continually exceeded the 0.30 working level average and has been monitored weekly. Generally, only radon daughter working level data has been collected on the Historic Tour due to limited personnel. However, other special measurements of the uncombined fractions of radon daughters with wire screens, tsivoglou method for radon daughter sampling (Thomas modification, 1970), and thoron daughter monitoring. These special measurements have not been routine due to time limitations involved in radon daughter sampling of other occupied portions of the cave. SUMMARY OF DATA The Historic Tour has been the most consistantly monitored tour since elevated levels of alpha radiation were found to exist at Mammoth. Cave. It is also the only natural entrance to the main sections of the cave and provided an opportunity to study man made actions upon the natural entrance. For these reasons the Historic Tour was isolated for study. Beginning October 10, 1977, and ending November 20, 1977, a pilot project was undertaken involving the Historic Tour and the practice of covering the natural entrance to this tour with sheet metal in the winter months. The purpose was to study radiation levels on the Historice Tour while the covers were on and off the natural entrance. In this pilot project, comparisons were made with incast air with covers on and off the entrance, and outcast air with covers on and off the entrance. TABLE 1 Incast air Mean W.L. Covers on . . . . 1.46 W.L. Increased 54% Covers off. . . . 0.67 W.L. when covers on Outcast air Mean W.L. Covers on . . . . 1.33 W.L. Decreased 5% Covers off. . . . 1.40 W.L. when covers on The natural entrance was artificially covered in the winter months (Yarborough, 1978) to protect the visitor from the extremely cold incast air, in the first four years of monitoring. The data in Table 1, illustrated in Figures 1 and 2, shows that this action increased the radon daughter working levels on the Historic Tour by 54% when the covers were on the entrance and the airflow was incast. While the air flow was outcast at the natural entrance, it made little difference as to whether the entrance was closed or open. Some interesting findings were observed when

Jan 1, 1981

By Wade E. Cooper

INTRODUCTION How accurate are past and present employee radondaughter exposure records of underground uranium miners employed in the United States? This often-debated question is essential for future substantiation of safe exposure limits. An apparent discrepancy between company-reported exposures and Mining Enforcement and Safety Administration (MESA) projected exposures was detected in 1977. For these reasons a need for an updated comparison of these exposure data was indicated. This paper gives some of the conclusions of the earlier study and compares more recent exposure records compiled by the Atomic Industrial Forum, Inc., with projected exposures based on sampling by Federal mine inspectors. EARLIER STUDY In its 1977 Annual Report (U.S. Department of the Interior, 1978), MSHA's predecessor, the Mining Enforcement and Safety Administration (MESA), reported that there was "an apparent discrepancy between Federal inspection results and company records." Both company records and MESA's projections from samples taken during routine Federal inspections indicated reductions in the average exposure of underground uranium miners from 1975 to 1977, but the MESA projections were over 4 times higher than the company-reported averages. This apparent discrepancy however, was based on a comparison of exposure data reported for all U.S. underground uranium miners. This projection more closely represented the average exposure of U.S. underground uranium mine production workers who worked 1,500 hours or more during the year. Exposures of such workers are reported each year by the Atomic Industrial Forum, Inc. (AIF) in summaries of exposure data reported to the AIF by uranium mining companies throughout the United States. (The AIF exposure summary for 1979 appears as tables A-1 and A-2 in the appendix of this paper.) Assuming that the average exposure for each exposure range category is the midpoint of each exposure range category, table 1 compares the estimated average exposures for U.S. underground uranium mine production workers who worked underground 1,500 hours or more each year in 1975 through 1977 with the exposures projected by MESA for those years. [Table 1. - Average Exposure and Projected Average Exposure for U.S. Underground Mine Production Workers Who Worked Underground 1,500 Hours or More During the Year. Company, MESA?' Reported- Projected Year (WLM) (WLM) 1975 1.59 5.68 1976 1.84 4.64 1977 1.68 4.08 1 Atomic Industrial Forum, 1976, 1977, 1978. 2 U.S. Dept. of the Interior, 1978.] Table 1 indicates that, even after adjustment to ensure better comparability an apparent discrepancy between Federal inspection results and company reported exposures for 1975-1977 exists; however, the apparent discrepancy diminished over the 3 years. Slade, 1977, explained some of the discrepancy between company records and MESA projections of miners' average radon-daughter exposures as follows: 1) Concentrations of radon daughters in some work areas can vary greatly during any one day. A variation from 0.3 WL to 17.0 WL has been measured in the same stope on the same day. 2) Seemingly simple abatement problems indicated by the regular Federal and State inspections were solved simply by manipulating the mine ventilation. 3) The methods used by mine operators to compute cumulative exposures were such that high radiation readings were seldom or never reflected in the records. For example, a work area sampled on Monday indicated a radon-daughter level equal to 0.2 WL and this was recorded. It was sampled again on Wednesday of the following week and the level was 2.2 WL. The miners were withdrawn or told to fix the ventilation, and when this was accomplished the area was sampled and found to be at 0.2 WL again. Although the miners could have been working in the higher concentration up to 6 days, this reading might never be reflected in their records. If it was recorded, only a fraction of the day on which it was discovered would be entered into the cumulative exposure calculation (time-weighted average). 4) Some of the mines visited used a mine average radiation concentration, and every employee working underground was given the same exposure per unit of time spent underground. As a result of the 1977 study, more stringent sampling and recordkeeping standards were proposed and public hearings held in 1977. The resulting new and revised health standards on radon-daughter sampling and exposure recordkeeping became effective August 30, 1979 (Mine Safety and Health Administration, 1979). Prior to these new regulations, radondaughter sampling requirements were on an "as often as necessary" basis (Code of Federal Regulations, 1978). The new regulations required practically all active work areas in underground mines to be sampled at least once every 2 weeks, with many areas requiring weekly sampling. They also required calendar-year exposure records of all underground

Jan 1, 1981

By J. Lovering, H. Wilhelm, P. Siewert

Introduction The Afton property is located 290 km (180 miles) by air east-northeast from Vancouver and 14 km (8.7 miles) west of Kamloops, a city of 60,000 people, in south central British Columbia, Canada. The mine is adjacent to the Trans-Canada Highway at an elevation of 670 m (2198 ft) above sea level. The ore body is a porphyry copper deposit that has undergone supergene alteration. The major economic minerals in the supergene zone are native copper and chalcocite with chalcopyrite and bornite in the hypergene areas. The grade is 1% with an overall copper distribution - 70% native, 25% chalcocite, and 5% chalcopyrite with bornite and covellite. The ore also contains important but variable amounts of gold and silver. The mill was designed to treat 6350 t/d (7000 stpd). Semiautogenous grinding was selected to minimize capital cost and because of the expected high clay content of the ore, which would have caused problems in a conventional crushing and screening plant. Test work indicated that a recovery of 87% was possible in a circuit incorporating both flotation and gravity separation. Flowsheet Run-of-mine ore is crushed in a 1.06 x 1.65-m (3.5 x 5.4-ft) Allis Chalmers gyratory crusher set at 228.6 mm (9 in.), closed side setting. The surge pocket, below the crusher, is emptied by a Hydrastroke feeder onto number one conveyor, which discharges onto a 180,000-t (198,416-st) coarse ore stockpile. Six Hydrastroke feeders on two conveyors withdraw the crushed material from the bottom of the pile. These two conveyors, in turn, discharge onto the belt feeding the semiautogenous mill. The live storage in the stockpile is approximately 22,000 t (24,250 st), sufficient for three days' mill feed. Primary grinding is accomplished in an 8.5-m (28-ft) diam by 3.7-m (12-ft) long Koppers (Hardinge Cascade) mill (Fig. 1) containing a 10% ball charge and driven by a 4000-kW dc variable speed motor. The mill dis¬charge is pumped by a 10 x 12 G.I.W. pump to a 1.22 x 4.88-m (4 x 16-ft) stationary screen sloped at 20°. Screen oversize returns to the semiautogenous mill (SAM), and the undersize flows by gravity to the ball mill discharge pump box. Secondary grinding is performed in a 5-m (16.4-ft) diam by 8.84-m (29-ft) Koppers overflow ball mill driven by a 3430-kW synchronous motor through an air clutch. The mill is in closed circuit with a Krebs Cyclopac containing 10 635-mm (25-in.) cyclones and the cyclone overflow, at 35% solids and 65% to 70% -200 mesh, is flotation feed. In order to limit the buildup of native copper, circulating in the secondary grinding circuit, a portion of the underflow from the cyclones is processed in a circuit containing screens, cyclones, and shaking tables to produce a finished metallic copper concentrate. Primary mill variable speed drive The overall waste to ore ratio at Afton was 4.5:1. The mining was to be done with only three shovels, which meant that it was highly unlikely that more than one of them would be in ore at any one time. The resulting inability to blend the mill feed made it impossible to prevent wide swings in the grade and grindability. The variable speed do drive motor installed on the semiautogenous mill was selected because of the extreme variability of the Afton ore body. This variability has persisted throughout the lifetime of the mine. There are times, however, when due to ore conditions, the mill is operated at full speed (78% of critical) for extended periods of several shifts duration. There are other times when the mill speed may be changed several times in a 12-hour shift due to changing ore conditions. When ore is processed that contains a fairly large proportion of fine native copper, the primary mill speed and, consequently, the tonnage may be reduced to improve the secondary grind and to maintain an acceptable grind and recovery. High clay ores require less mill speed and more dilute grinding densities. In the latter case, the slower primary mill speed also helps to minimize damage to the mill liners. Approximately 57% of the time the mill operates between 90% and 100% of full speed or between 71% and 78% of critical. The variable speed is also used for inching during mill relines.

Jan 1, 1987

By R. W. Carn, D. Vincent

During the next 15 years, US coal production is expected to double, with the increased production evenly divided between the East and the West. Along with greater production, coal export markets should increase dramatically from East, West, and Gulf Coast ports. The annual overseas export capacity of US coal-loading terminals is expected to rise from 147.1 Mt (162.1 million st) in 1981 to a minimum of 278.1 Mt (306.6 million st) in 1985, according to the US Maritime Administration. Increased coal production and use will lead to more development of import and export terminals, a vital link in the coal transportation chain. With continually escalating capital costs and the competitive markets that the terminals will serve, a well designed and efficient terminal is necessary. This article begins a two-part series that presents concepts presently used in coal export terminal design. Part I looks at site selection factors and equipment needs, while Part II will examine environmental considerations in building a terminal as well as typical capital and operating costs. The world is nearing the end of the oil era. In a few years oil will not be available to sustain the growth rate and increasing standard of living we have known in our lifetime. The big question is what energy era are we moving into? With the decline of readily available oil reserves and rapidly increasing prices, many countries are trying to switch to alternate energy forms. While intensive efforts to find new oil reserves continue, alternate energy sources such as natural gas, coal, synthetic fuels, nuclear, hydroelectric, solar, and wind power are being developed. Recent indications are that coal is expected to bridge the energy gap over the next 25-30 years until the technology and economics of the alternate energy forms reach satisfactory levels. Use of coal for energy is receiving strong attention due to its long-term availability (200-300 years minimum), relative ease of development, and its low cost per unit of power produced. By the year 2000, it is expected that 25% of world energy supply will be met by direct coal combustion and possibly another 5-10% by synthetic fuel from coal. Coal's expanding share in the world energy market, along with an increase in coking coal requirements, will result in a large increase in the world's seaborne coal trade. Recent statistics and projections for the future are shown in Table 1. This phenomenal development rate includes increases in both coking and thermal coal requirements. Because of the rapid increase of seaborne coal trade during the last 10 years and the even greater projected increase of trade to 2000, various sectors of the coal industry are faced with enormous technical challenges and huge investments in equipment, land, transportation systems, and port facilities. Very large bulk terminals are under development throughout the world. Latest surveys indicate that there are about 30 new coal export and import terminals under consideration and at least 30 existing terminals have expansion programs planned or underway. With the high cost of borrowed capital and rapid inflation rates there is great emphasis on new planning and design techniques to minimize capital and operating costs of coal transportation systems. Terminals A total coal supply system can be considered to consist of one or more mines; a train, barge, truck, or other haulage system; an export terminal; a fleet of bulk carriers; a receiving terminal; and possibly, local inland distribution networks that include barges and railways. Terminals, though only a small link in the total transportation system, play a key role in overall system efficiency. At ports or inland distribution centers, terminals act as transportation links bringing trains, ships, barges, or trucks together for cargo transfer and temporary storage. A well-designed terminal can provide maximum independence between two modes of transportation and optimum freedom for intermodal interference. A terminal acts as a buffer between the two transportation modes by providing sufficient storage capacity so a ship need not wait for its cargo on, for example, a train-by-train basis, but can load immediately from the ready stock. Similarly, a train need not wait for a ship to unload its contents but can dump immediately into storage. A terminal also can be used to properly mix various types of coal to satisfy a buyer's requirements. Consider the relative value of various production and transport segments for a typical steam coal

Jan 6, 1983

By P. Wearden, K. Bryner, K. Vrana, V. Castranova, R. Dey, R. Reist, J. Blackford

INTRODUCTION Proliferation and enhanced synthesis of collagen by pulmonary fibroblasts have been shown to be key steps in the development of chronic silicosis (Goldstein and Fine, 1986). The regulation of lung fibroblast proliferation by cytokines released from alveolar macrophages may be an important pathogenetic mechanism in the development of the fibrotic process (Kelley, 1990). One cytokine, platelet-derived growth factor (PDGF), promotes fibroblast proliferation by inducing the movement of quiescent (Go) cells into the C1 phase of the cell cycle (Chen and Rabinovitch, 1989). Others regulate the rate of transition of fibroblasts from Gl into the S phase (Leof et al., 1982). These two classes of cytokines have been termed, respectively, competence and progression factors. One approach used to examine the release of cytokines from macrophages is the fibroblast proliferation assay in which fibroblasts are exposed to culture supernatants from macrophages exposed to various stimuli. In most of these assays, the supernatant contains fetal calf serum which provides the competence factor(s) necessary to facilitate the proliferation of fibroblasts (Bitterman et al., 1982; Bitterman et al., 1983; Elias et al., 1988). Recently, a fibroblast proliferation assay using plateletpoor plasma (lacking competence factor(s)) as a substitute for fetal calf serum has been described (Kuman et al., 1988; Bauman et al., 1990). In this assay, the release of a competence-inducing PDGF-like growth factor from rat and human macrophages can be distinguished from other cytokines that act as progression factors. In order to obtain more consistent results and with the ultimate goal to be able to discriminate between the effects of competence factors as opposed to progression factors, we have conducted experiments to determine the appropriate concentrations of plasma and PDGF required for imparting competence in the fibroblast proliferation assay. We tested lung fibroblast cells obtained from explants of rat lung tissue and also a fetal human lung fibroblast cell line obtained from American Tissue Culture Collection (ATCC153). MATERIALS AND METHODS Fibroblasts Specific pathogen-free, male Sprague-Dawley rats were use in some studies. Animals were given a lethal intraperitoneal dose of sodium pentobarbital. Fibroblasts were isolated by chopping the lung in enzymes that digest the connective tissue but liberate lung cells for further study (Rabovsky et al., 1989). After digestion, the remaining lung tissue suspension was filtered through two layers of sterile gauze and centrifuged to recover lung fibroblasts. These were resuspended in culture medium that contained 10% fetal calf serum and distributed to culture plates for growth. In other experiments, a human fetal lung fibroblast cell line, obtained from American Type Culture Collection, Rockville, MD, 20852, was used instead of rat lung fibroblasts. In these cases, a 1 ml ampule containing human fetal fibroblasts was plated into a tissue culture flask containing medium plus 10% fetal calf serum. For both types of fibroblasts, culture medium was changed 3 times per week and cultures were incubated at 37°C until confluent. Harvested rat and human lung fibroblasts were quantified using an electronic cell counter equipped with a cell sizing attachment (Coulter Electronics, Inc., Hialeah, Florida). Tritiated Thymidine Incorporation The basic procedural outline of Kumar et al. (1988) was used with modifications to evaluate tritiated thymidine incorporation into fibroblast DNA following exposure to PDGF and plasma. Both rat and human lung fibroblasts were plated at 50,000 cells/ml at a density of 250,000 cells/25cm2 culture plate. Cells were quiesced for 4 days with 2% rat plasma. As the assay was refined, fibroblasts were quiesced in plasma-free media for 48 hrs, since the mitogenic activity of 2% plasma was variable. Test medium was applied for a period of 6 hrs, followed by a 24 hr tritiated thymidine (lµCi/ml) labelling period in plasma-free media. Medium alone was used as a negative control and media with 10 or 20% fetal calf serum was used as the positive control for rat and human fibroblasts, respectively. Cell Quantification and Measurement of Mitogenesis Twenty-four hours after the addition of tritiated thymidine, the fibroblasts were washed with 5ml of fresh serum-free media, centrifuged and resuspended in phosphate-buffered saline. The cells were dissolved in 0.5m1 of O.1N NaOH and radioactivity determined in a beta counter. Incorporation of trititaed thymidine as an index of DNA synthesis was expressed as DPM/fibroblast. RESULTS In the present study, we quantified mitogenic potential by monitoring the incorporation of tritiated thymidine as

Jan 1, 1991

By William T. Jr. Renfroe, Donald C. Gipe

INTRODUCTION Historically, pollution control in the metal-ore mining industry has been very limited. Unless mine water contained large quantities of solids, it was generally discharged without any treatment. If treatment was used to control solids, it was principally the provision of a settling basin in the form of a tailings impoundment used in conjunction with an associated metal ore dressing facility. Recently, however, a growing awareness of the adverse environmental impacts of mine drainage, coupled with strict environmental laws, has prompted the mining industry to look at new technologies and to refine the existing methods to further treat the wastes generated. This industry is unique in that waste loadings are extremely variable, and a "typical facility with typical waste loads" does not exist. Consequently, one waste- water treatment system cannot be utilized on an industry wide basis; rather, each treatment system must be designed specifically for the pollutants in each individual discharge. Public Law 92-500, the Federal Water Pollution Control Act (FWPCA) Amendments of 1972, became effective on Oct. 18, 1972. This law completely restructured Federal laws and philosophies underlying the Federal approach to water pollution control. Prior to the 1972 amendments, the principal Federal regulatory tool had been water-quality standards based on a designated use for a particular body of water. The concept was that waste disposal into water bodies is a desirable and acceptable use of the water body if it does not interfere with other beneficial uses. This had the effect of requiring various degrees of treatment and, consequently, various economic hardships on industries de- pendent upon their location. In many waterways. it is very difficult to quantitatively relate discharges to water quality. The 1972 amendments changed the basic philosophy, as stated in the Senate Committee report on the bill, to ". . . no one has the right to pollute . . . that pollution continues because of technological limits, not because of any inherent right to use the nation's waterways for the purpose of disposing of wastes." Pursuant to Sections 301, 304(b), and 306 of the FWPCA Amendments of 1972, the US Environmental Protection Agency (EPA) was required to establish effluent standards applicable to all industrial discharges. These standards must be based upon the application of the "best practicable control technology currently avail- able" (BPT) and the application of the "best available technology economically achievable" (BAT). The BPT and BAT levels must be achieved industry-wide by July 1, 1977, and July 1, 1983, respectively. WASTE SOURCES The waste-water situation in the mining segment of the ore mining and dressing industry is unlike that encountered in most other industries. Most industries (e.g., the milling segment of this industry) utilize water in the specific processes they employ. This water frequently becomes contaminated during the process and must be treated prior to discharge. However, in the mining segment, process water normally is not utilized in the actual mining of ores (exceptions are hydraulic mining operations and dust control), but it is a natural occurrence that interferes with mining activities and must be removed before mining can commence. Water enters mines by ground-water infiltration and surface runoff, and it comes into contact with materials in the host rock, ore, and overburden. The underground mine must pump large quantities of ground water to prevent flooding of the mine. Water from surface mining operations generally occurs as a result of surface runoff of rainwater. Generally, mining operations control surface runoff through the use of diversion ditching and grading to prevent, as much as possible, excess water from entering the working area. Nevertheless, some surface runoff does come into contact with the working area and may become contaminated. The quantity of water from an .ore mine is unrelated, or only indirectly related, to production quantities. De- pending upon its quality, the mine water may require treatment before it can be discharged into the surface drainage network. The variability of water quality from mines can best be demonstrated by looking at Table 1. This table shows the range of pollutant concentrations in untreated discharges from three different categories of mines (as categorized by EPA in the development of BPT and BAT effluent standards for the metal-mining industry). Data for this table were obtained during EPA's preparation of effluent standards for this industry. The parameters shown on the table are the pollutant parameters of primary interest in this industry; blanks in the table indicate that data were not available, and the parameter is not expected to be present in significant quantities. Other pollutant parameters are present in mining waste water, but they are either incidentally removed in the treatment process or are found only in trace amounts. The three categories comprise more than 90% of the metal production value in the United States and approximately 95% of the total mine discharges. It is important to note that not all parameters are found in significant concentrations at all locations. IMPACT ON WATER QUALITY One of the most troublesome mine-drainage problems is acidity. Although generally associated with coal mining, acid mine drainage frequently occurs from other types of mines. Although the exact mechanism of acid mine drainage is not fully understood, it generally is believed that pyrite (iron sulfide, FeS,) is oxidized by oxygen (Eq. 1) or ferric iron (Eq. 2) to produce ferrous sulfate (FeSO4) and sulfuric acid (H2SO4) . The mining of ores associated with pyritic material exposes the pyrites to water and oxygen and grossly accelerates the natural oxidation processes, resulting in the significant production of acid mine drainage.

Jan 1, 1982

By C. A. Rowland

Introduction Ball mills are lined drums, either cylindrical in shape or modified cylinders that have either one or both ends of the shell, consisting of conical sections, that rotate about the horizontal axis. Fig. I I shows a cylindrical mill, Fig. 12 a conical ball mill, and Fig. 13 a Tricone ball mill (Hardinge tradename). Steel or iron grinding media, generally in the shape of spheres, are used to grind the ore to the specified product size. In order to obtain more contact area for grinding and to simulate the shape of worn balls, balls have been made with two concave surfaces diametrically opposite each other. Some concentra¬tors, such as Erie Mining Co., have used slugs cut from worn and broken rods to supplement the balls in ball mills and save money otherwise lost as rod scrap. Cylindrical and conical shapes have been tried instead of balls, but balls remain as the most common shape grinding media used in ball mills. Ball mills were a logical development from the earlier pebble mills that used hard natural pebbles such as flint pebbles or sized ore pebbles (obtained from the ore itself) as grinding media. In the early 1900s36 it was found that when cast iron or cast steel balls were used in place of flint or ore pebbles, the mills drew more power and gave greater production capacity. Advances in technology have resulted in the manufacture of ball mills up to 18 ft diam inside shell, drawing up to 8,000 hp. Ball mills are employed to grind ores, especially the more abrasive ores, to finer sizes than can be produced economically in other size¬reduction machines such as roll crushers, hammer mills, and impactors. Ores can be ground dry-dry grinding-or in a slurry-wet grinding-using ball mills. Dry grinding nominally refers to less than I %v moisture by weight. If the moisture content increases by several percent, dry grinding capacity is significantly reduced as shown in Table 17. The usual range of solids content in wet ball-mill slurries is from 65 to 80% by weight. Wet grinding is used to prepare the feed material for unit opera¬tions such as flotation, magnetic separation, gravity concentration, and leaching that require a slurry of liberated valuable mineral and unwanted gangue particles. Dry grinding" is employed to produce feed for agglomeration, pelletizing, and pyrometallurgy processes that require feed that is dry or nearly so and for finely ground industrial mineral products used in the dry state. Dry grinding is also used when minerals cannot be dewatered economically to the required moisture level or when the ground product reacts unfavorably with liquids. For example, cement clinker must be ground dry. Dry grinding requires about 30% more power than wet grinding for comparable size reduction .28 The total power required in a dry¬grinding ball-mill plant including drying may be double that required for a wet-grinding plant. Grinding-media and liner consumption in dry grinding reported as pounds of metal consumed per kilowatt-hour per ton of ore" is 10-20% of that used in wet grinding. The Wabush pellet plant, Point Noire, Que.3o reported ball consumption dropped from 6.3 lb per ton of ore ground to 2.5 lb per ton of ore ground when they converted from wet to dry grinding, and a 30% increase in power consumption. A number of comparisons made on wet and dry grinding of cement raw materials show metal consumption in dry grinding to be 10% of that in wet grinding. The capital costs for wet grinding are generally lower than for dry grinding. When thickening and filtering of the wet-ground product are required, dry grinding may have a lower capital cost. With open-circuit grinding the ball-mill discharge passes directly to the next processing step without being screened or classified and no fraction is returned to the ball mill (Fig. 14). In closed-circuit grinding the ground material, undersize, in the ball-mill discharge is removed either using a screen or a classifier with the oversize being returned to the mill for additional size reduction (Fig. 15). The over¬size material that is returned to the ball mill is called the circulating load. Open-circuit ball-mill grinding requires more power than closed¬-circuit grinding for products containing similar amounts of top-size material. The less the amount of oversize allowed in the product, the longer the ore must remain in the ball mill when grinding in open circuit. This increases the production of extreme fines and thus the consumption of more power. The power required for open-circuit ball-mill grinding can be estimated using the multipliers listed in Table 18 and knowing the power required for closed-circuit grinding to yield the desired product particle size. For example, assuming the desired grind size is 90% passing some specific top size, open-¬circuit grinding would require 1.40 times the power to achieve similar results as closed-circuit grinding.

Jan 1, 1985

By Janet Briggs, Michael Vertes, J. F. Shirley, R. E. Cutherbertson, Alexander Sutulov, J. D. Vincent

Molybdenum, the 42nd element in the periodic table, is now a very important metal in industry of the United States and the world. Chemically, it is a very complex element having valences of 0, +2, +3, +4, +5, and +6, and can exist as a mixture of these valences in many compounds. It was discovered and separated by Scheele in 1778. At the present time, molybdenite is the only important commercial source of molybdenum. Molybdenum occurs in some 12-14 different recognized minerals, but practically all of the production is molybde¬nite. As late as 1915, it was considered a rare metal. Germany made extensive use of it during the first world war, which resulted in the establishment of a molybdenum industry in the United States at Cli¬max, Colo., in 1917. It was not until the mid-1920s when its introduc¬tion to steelmaking developed for the US automobile industry that the metal gained important acceptance. The first notable example was the Willis-Saint Claire automobile introduced in 1921 by C. Harold Willis, formerly chief metallurgist for Ford Motor Co. Molybdenum has also been produced in the US from wulfenite (PbMoO4) from Arizona and Nevada, and a small amount of molyb¬denite was produced up to the early 1970s as a coproduct from two molybdenum-bismuth mines in Quebec, Canada. The first byproduct production of molybdenite from a copper con¬centrating operation was achieved at Cananea, Sonora, Mexico, and was followed shortly by the Utah division of Kennecott Copper Co. in the mid-1930s. This section of the Handbook is divided into four main subdivi¬sions: Chapter 2, which deals with the concentration and recovery of molybdenite from primary molybdenum ores; Chapter 3, which deals with the recovery and separation of molybdenite as a byproduct from copper production; Chapter 4, which deals with the conversion practices employed for converting molybdenite to the oxide or metal form; and Chapter 5, which covers the uses of molybdenum in steel¬making, chemicals, pigments, lubrication, catalysts, etc. The Free World molybdenum production, for the past eight years from all sources, is shown in Table 1. 2. Primary Molybdenite Ores General In 1978 there were four primary molybdenite recovery plants in operation: Climax, Endako, Quests, and Henderson. For over half a century the Climax mine has been by far the largest and most important source of molybdenum in the world. During the early 1970s a few primary molybdenite mines in British Columbia were shut down due to adverse economical conditions. The largest of these was British Columbia Molybdenum, formerly owned by Kennecott Copper Corp. The Climax plant and operations are described in detail and the opera¬tions of the other plants are summarized following a general discussion of recovering molybdenite from primarily molybdenite ores. Recovery of molybdenite as a byproduct will be taken up in Chapter 3. Geology Molybdenite production from mines in which it is the sole or preponderantly important mineral is confined to ore bodies in granites or igneous-type rocks or pegmatite dikes. More than one state of mineralization is sometimes present as at Climax, York, Hardy, and Questa. In the case of Climax, important byproduct recovery of tung¬sten, tin, pyrite, and monazite is also attained. The molybdenite is commonly disseminated very finely in the siliceous granitic gangue but does occur in stringers and blebs in silicified mineral solution channels. Treatment of Primary Molybdenum Ores All primary molybdenum ores are low in molybdenite, the content ranging from less than 0.20% MoS2 to a high of about 0.40% MoS2. A high resistance to grinding is a common ore characteristic combined with a habit of very fine dissemination of molybdenite in the siliceous portion of the ore. Because of low MoS2 content, high hardness, and fine dissemination of molybdenite in primary ores, it has been necessary to develop a special mill treatment which yields an accepta¬ble recovery at a reasonably low milling cost. Essentially, the treatment consists of intensive rougher flotation at a relatively coarse grind with production of a low grade rougher froth containing a substantial percentage of the molybdenite in the form of low grade middling particles, followed with multistage regrinding to liberate the locked molybdenite and several stages of flotation cleaning to produce a finished grade concentrate. Rougher flotation on pulp ground to ap¬proximately 35 mesh is common practice, even if complete liberation of molybdenite from siliceous gangue may require regrinding to 200 mesh. The flotation of middling particles with a very low content of molybdenite in the rougher flotation step has been made possible by the application of a combination of reagents first developed at the Climax mill and used with some modification in the treatment of all primary molybdenum ores. The reagents commonly employed and their functions in flotation are:

Jan 1, 1985

By Paul M. Jr. Musgrove, Donald C. Moore

Conventional Smelting Practice Conventional copper smelting practice varies from smelter to smel¬ter, but generally consists of some or all of the following unit processes: roasting, smelting, converting, and fire refining. Roasting. Copper sulfide concentrates can be smelted directly or after an initial roasting step. Roasting is used in some smelters because roasting prior to smelting increases smelting capacity, less energy is required to melt hot roaster calcines than wet sulfide concen¬trates, roaster off gases are high in Sot concentration, 5-15% SO2, and some volatile impurities are removed from the concentrate prior to smelting. However, many smelters do not use roasters, because the problems associated with handling hot dry calcines outweigh the advantages mentioned. Concentrate roasting is performed in multiple hearth or fluid bed roasters. If the moisture is low, roasting can be performed autogenously, usually at 500-600°C. High roasting temperatures are avoided because excess oxidation of the iron compounds may lead to magnetite formation. Magnetite is detrimental to reverb operation because mag¬netite can combine with refractory minerals to form a highly viscous slag. This slag prohibits efficient matte-slag separation and leads to excessive copper losses. Also, magnetite can settle through the matte layer, deposit on the furnace bottom, and consequently reduce furnace capacity. Roasting is carried out only on sulfide concentrates prior to smelt¬ing in reverb or electric furnaces. For smelting processes, such as the flash and continuous that rely on the exothermic heat of oxidation of the sulfur minerals, roasting is not practiced. Reverberatory Smelting. The predominate copper smelting fur¬nace for the past 50 years has been the reverb. These furnaces are typically 100-120 ft long, 30-35 ft wide, and 12-15 ft high. A typical furnace layout is shown in Fig. 2. Refractory brick linings cover all internal surfaces of the furnace. Originally the flame was directed to reverberate or reflect off the furnace ceiling and melt the feed material. Current practice is to direct the flame down the furnace length to melt the concentrate. A method of charging the concentrates or calcines, generally along the side walls to minimize refractory erosion, is incorporated in the furnace design. The copper concentrates, calcines, and fluxes charged into the reverb undergo a series of complicated reactions as the temperature of the mixture increases. The reaction of the iron and copper sulfides with the oxygen in the furnace produces a molten Cu25-FeS mixture called matte. Copper smelting metallurgy is based on the fact that sulfur has a greater affinity for copper than for iron and most other common metals. Therefore, in a system containing copper, the copper will preferentially remain as a sulfide compound until all of the other metals have been oxidized. The oxidized metals combine with silica to form a silicate slag that floats on the matte and is removed from the system. Reverberatory furnace smelting chemistry can be approximated by the following chemical equations: FeS2 + O2 - FeS+ SO2 (1) The formation of FeS ensures that any copper present other than as sulfides will be reduced by the relationship: CuO2 + 2FeS + O2 - CuS + 2FeO+SO2 (2) or 2Cu +FeS - Cu2S + Fe (3) As the molten charge travels down the furnace, continued oxida¬tion of the iron minerals and sulfurization of the copper minerals occurs. When all of the copper has been converted to sulfides, the iron sulfides can then be further oxidized as: FeS + (3)2 O2 FeO + SiO2 (4) The FeO reacts with the silica added as flux in the furnace charge. A simplified equation is: FeO + SiO2 -FeO SiO2 (5) The iron silicate slag formed is skimmed from the surface at the end opposite the burners. The copper content of reverb slag is usually less than 0.6% Cu and is discarded. Matte is removed along the side wall and is taken to the converter for oxidation of the remaining sulfur and iron. The main objectives in reverberatory smelting are to produce a molten Cu2S-FeS matte containing 30-60% Cu and a throwaway slag. Production of matte permits complete conversion of all copper minerals into copper sulfides, which can migrate because of specific gravity differences, through the lighter slag layer. Also, the molten matte droplets collect the noble metals, gold and silver, as the matte settles in the furnace. The large settling area of the furnace provides enough separation time to produce a low grade slag, which can be discarded without further processing. High heat losses are associated with reverberatory smelting be¬cause of the large volume of gases sweeping through the furnace. Therefore, an outside source of heat is required to keep the smelting reaction going. Natural gas, fuel oil, or pulverized coal are used as this heat source.

Jan 1, 1985

By Robert G. Beverly

INTRODUCTION Standards limiting the annual exposure of United States uranium miners to radon daughters were established in 1967 at 12 Working-Level-Months (WLM). The standard was reduced by a factor of three, to 4 WLM, in 1971. Currently, the standard is again being examined to determine if it should be changed. Since 1967, Union Carbide has calculated individual monthly exposures in company and contract-operated mines located on the Colorado Plateau. Although it has been possible, by extensive ventilation control measures and accurate routine sampling, to meet the current exposure standard, there are many miners whose exposures closely approach the 4 WLM standard for any given year. However, it was noted that for miners who work for any extended period of years the [average] exposure was much less than the standard. The primary purpose of this paper is to show that, in effect, any annual exposure standard to radon daughters results in a long-term exposure considerably below that standard. Further, most miners, due to their job assignments and/or employment habits, only receive a small fraction of the standard. HISTORY OF EXPOSURE STANDARDS Prior to 1967, radiation protection in uranium mines was fundamentally based on a radon daughter concentration guide. In 1960, the American Standards Association published mine and mill radiation protection standards (ASA-1960). The Colorado Department of Mines, in 1961, adoped a standard which followed the ASA Standard and provided that if concentrations exceeded 10 Working Levels (WL), the area was to be shut down until corrective action was taken; if between 3 and 10 WL, corrective action was to be initiated; between 1 and 3, additional samples were to be taken and individual exposures evaluated; and if below 1 WL, conditions were considered to be controlled. In 1967, the U.S. Department of Labor issued the first exposure standard which called originally for limiting annual exposures to 3.6 WLM but which was later changed to 12 WLM. The complicated regulatory developments leading to this standard have been described elsewhere (Beverly-1969, Rock & Walker-1970). Effective July 1, 1971, this exposure standard was lowered to 4 WLM per year, which is the current standard. Over the past year, there has been speculation about the potential risk to uranium miners working at the present standard. A recent NIOSH Study Group Report (NIOSH-1980) concluded: "There is also strong evidence that a substantial risk extends to and below 120 WLM of exposure." The 120 WLM corresponds to a miner working in uranium mines for 30 years, a rare occurrence, at an exposure rate of 4 WLM per year, an even rarer occurrence. On the other hand, the General Accounting Office, in a recent Report to the Congress (GAO-1981), was very critical of reports by NIOSH on general low-level radiation risks. The GAO recognized that”...important questions remain unanswered about the cancer risks of low-level ionizing radiation exposure;" and recommended that Congress enact legislation giving statutory authority to an interagency committee to coordinate Federal research on health effects of ionizing radiation exposure. The International Commission on Radiation Protection at its March, 1980 meeting recommended limiting the inhalation of radon daughters to 0.02 J per year, equivalent to 0.4 WL, which on an annual basis would be 4.8 WLM and noted it is common to reduce this figure by 20% for allowance in the case of uranium miners for external and/or dust exposure(Sowby-1980). This is essentially equal to the present standard of 4 WLM. As earlier uranium miner exposure studies are reevaluated, and as new studies are conducted, it is important that the relationship between regulatory standards and the resulting actual exposures be recognized. UNION CARBIDE URANIUM MINING EXPERIENCE Union Carbide started mining Colorado Plateau uranium-vanadium ores in the late 1920s for the contained vanadium values. In the early 1950s, the Atomic Energy Commission contracted Union Carbide to produce uranium at mills located in Uravan and Rifle, Colorado. The company now has over fifty years of mining experience in the area. Some mines are operated as company mines and others are operated by private mining companies under a contractual arrangement. Ventilation, sampling, and exposure calculations are carried out the same in contract mines as in company-operated mines. Data presented in this report do not differentiate between company or contract employees and include all employees who worked underground any portion of a year in Union Carbide mines from 1967 through 1980. At the peak of uranium mining activities in 1970, there were 577 miners employed at year end (285 company employees and 292 contract) and 52 mines in operation (8 company-operated and 44 contract mines). Contract mines varied from two-man operations up to 15 employees. Company mines were generally the larger operations and employed from 20 to 100 miners.

Jan 1, 1981

By Friedrich Steinhäusler

RADON SOURCES AND CONTROL MEASURES IN THE MINING ENVIRONMENT Most of the contamination of the mine atmosphere by radon 222 is due to radon emanating from solid or fractured ore surfaces of walls, roof and floor. Also radon gas emanates from broken ore either from storage in backfilled mined-out areas as applied in e.g. shrinkage stopping methods or from ore spillage along intake airways mainly due to the use of trackless haulage. To a lesser extent water itself can represent an additional source of radon, which emanates into air from open drainage ditches or seepages along intake airways. The contribution from water can be controlled effectively by isolating the water from the primary intake air system, e.g. by diverting the water through pipes and/or sealing of seepages by grouting. However, control of radon emanating from rock surfaces creates a major technical problem with significant impact on the economic aspects of mining operations, if adequate radiological conditions must be maintained. Basically this can be achieved by suppressing the emanation process itself, confining already emanated radon or by removal of radon from the mine atmosphere. Extensive research has been carried out on the rate of radon emanation as a function of barometric pressure changes (Pohl-Rüling and Pohl, 1969). It could be shown that the radon supply consists of a permanent and variable component. The former results from the surface of the rock and depends mainly on the emanating fraction of its radium 226 content; the latter originates from within the rocks and is a function of the suction effect of decreasing barometric pressure, rock porosity and fissures. The practical application of this barometric pump effect for depressing the rate of radon emanation, e.g. by pressurizing the mine atmosphere, is limited due to high costs for providing a sink for absorption of radon and air as well as lack of permeability in most uranium ore bodies (Schroeder et al., 1966). Mine air cleaning by removal of radon can be achieved with the use of cryogenic methods, chemical removal, adsorption into charcoal beds, use of a gas centrifuge or general ventilation techniques. Technical problems have so far prevented the application of any of these methods other than ventilation. It is common practice to use the age-of-air concept, i.e. fresh air is delivered to the worker as directly as possible and removed quickly afterwards thereby maintaining the air "young". Engineering principles for quantity distribution of air through underground working areas are straightforward for general mining situations where radon constitutes an environmental contamination problem. However, in cases of high uranium ore content this concept may result in high costs with regard to installation and energy requirements for effecting both frequent air changes as well as sufficient heating of the air in cold seasons. Taking into account that the investment in ventilation systems is a major cofactor for the overall ore production costs this can be a limiting and decisive component in the assessment of the economic feasibility of specific mining operations and mineral reserves in general. Effective control of the radon flux from the rock surface prevents the initial contamination of the mine air with radon directly at the source. A radon diffusion barrier for practical application in mining requirements should fulfill the following requirements: - reduction of radon emanation rate by at least an order of magnitude - high mechanical strength - ease of sealant application onto surface to be coated - water resistant - low fire hazard - resistant to temperature changes encountered in mines - high cost efficiency in relation to exposure reduction achieved (direct and indirect costs) - low degree of maintenance. In the past several materials have been tested as sealants for controlling the emanation of radon from surfaces of rock and building materials. Epoxy paints reduce radon emanation rate only by a factor of 2 to 6 (Auxier et al., 1974; Eichholz et al., 1980; Keith Consulting Engineers, 1980). Although it is possible to prevent the escape of more than 99 % of the radon to the environment with gel seals over 80 mm thick (Bedrosian et al., 1974), practical applicability is very limited. Multilayer coatings of epoxy resins with various additives require meticulous preparation and flawless application of seamless four-layer coatings in four days to impede radon diffusion (Culot et al., 1976), otherwise results from this method have not been totally satisfactory (Leung, 1978). Aluminium foil laminated with polyethylene and paper on each side is under test as radon barrier but results are not available yet (Ericson, 1980). However, this method has the inherent disadvantage that possible malfunctioning electrical installations can cause fire or electrical shock through the sealant. Polyurethane foam coatings have been used on stoppings as very effective sealants. It does, however, represent a potential danger of spontaneous ignition and it is expensive (Rock, 1975). Thus, there is still need for a material which has similar properties as outlined above. In the following results are reported from investigations on the suitability of various materials as radon diffusion barriers.

Jan 1, 1981

By A. S. Paschoa, A. W. Nóbrega

INTRODUCTION There are now over twenty years since the radiological characteristics of the brazilian regions of Araxá, Tapira and Barreiros, three locations adjacent to each other, in the state of Minas Gerais, Brazil, started being surveyed by investigators from the Instituto de Biofísica da Universidade Federal do Rio de Janeiro (IB/UFRJ), Pontifícia Universidade Católica do Rio de Janeiro (PUC/RJ), and New York University (NYU) (Roser and Cullen, 1958; Cullen et al, 1980). The importance of the Araxá apatite – Ca5,(P04)3(OH, F, Cl) - and pyrochlore - NaCaNB206F - for the production of large quantities of fertilizers and niobium was early recognized (Roser and Cullen, 1958). Interesting data have been gathered and published throughout the years on the contents of naturally occurring radionuclides in geologic materials, soils, grass, foods and waters of the Araxá region (Eisenbud, et al., 1964; Penna Franca, et al., 1965a, b; Roser et al., 1966; Penna Franca et al., 1970; Cullen, 1977; Penna Franca, 1977; Cullen and Paschoa, 1978; and Cullen et al., 1980). The radioactivity of the Araxá region is associated with mineral deposits of niobium rich pyrochlore and phosphate rich apatite. There are in this region mineral ores with 3 to 5% pyrochlore with 2 to 2.5% Th as Th02 and 50 to 150 ppm U308 in the matrix, while the apatite deposits contain up to 150 ppm U308 and a thorium concentration similar to that of the pyrochlore deposits (Paschoa and Palacios, 1981). As an indication of the high thorium content of the Araxá soils the 228Ra and 224Ra activity concentrations were reported to range from 10.6±0.2 to 62.4±0.7 pCi228Ra/g soil in 7 samples and 2.1±0.1 to 104±1 pCi224Ra/g soil in 19 samples (Cullen, et al., 1966). The uptake of radium isotopes by edible roots vegatables and fruits growing in the soils of the Araxá region can be illustrated by the data listed in Table I. The biological availability of natural radium isotopes in some segments of the Araxá soil allows a large variation in the 228Ra, 224Ra and 226Ra concentrations in vegetables and edible roots, as can be seen in Table I. This fact makes quite difficult a quantitative local assessment of the radiological implications of mining the Araxá mineral deposits of pyrochlore and apatite for production of niobium and phosphate fertilizers, respectively; since one cannot easily separate the naturally occurring from the technogically enhanced radionuclide contents of foods. The position of the city of Araxá inside the contour of the state of Minas Gerais appears in the upper part of Figure 1, which shows the outlined map of Brazil with the positions of the cities Brasilia and Rio de Janeiro also indicated. The lower part of Figure 1 is a representation of the Araxá region in an expanded scale, which indicates the locations of the pyrochlore and apatite deposits in relation to the city of Araxá, as well as the nearby hydrographic basins. The distance between the city of Araxá and the pyrochlore deposit is about 8 km, and between the pyrochlore and apatite deposits is 4 km. This paper deals tentatively with the radiological implications of the industrial operations taking place in the Araxá region for the exploration of the pyrochlore and apatite deposits. However, one must bear in mind, firstly, that the radiological implications of these industrial activities are by far too complex to be covered adequately by the limited amount of data to be presented in this paper, and secondly, that such implications cannot be considered of local character only. INDUSTRIAL OPERATIONS The commercial exploration of the Araxá deposits of pyrochlore and apatite started only few years ago, motivated by the increasing demand of niobium and phosphate fertilizers in Brazil and the world. As a consequence of the industrial operations in Araxá, a redistribution of the uranium, thorium, and radium originally present in the local deposits of pyrochlore and apatite started occurring in the seventies, with possible radiological implications for the Araxá region and its immediate surroundings, not to mention the destinations of the end products of such industrial operations. A literature review on the radioactivity associated with the extration industries of selected minerals was made by the USEPA (Bliss, 1978). The low level radioactive wastes of the industries for copper ore mining and rare metals processing have been object of particular attention (Fitzgerald, Jr., 1976; Eng, et al., 1979), but the short and long term implications of the radioactivity associated with the niobium industry were also subjects of concern (Knight and Makepeace, 1978). Recently, high 232Th and 226Ra concentrations in samples from the tin mining industry in West Malaysia were reported (Hu et al., 1981). A great deal of attention has been dedicated to the implications of the redistribution of radionuclides originally present in the mineral ores used by the phosphate fertilizer industry (Moore, 1967; Menzel, 1968; Spalding and Sackett, 1972; Eisenbud, 1973; Guimond and Windham, 1975; Guimond, 1976; Roessler,

Jan 1, 1981

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By Clayton J. Parr

Introduction This chapter contains brief discussions of various environmental protection requirements that relate to underground mining operations. Environmental disturbances at an underground mining operation can result from subsidence; water discharges; waste dumps; construction and operation of access roads and utility lines; construction and operation of surface facilities such as maintenance shops, bathhouses, and storage yards; and emanation of dust and noise from surface crushers. Construction and operation of a concentrator or washing plant may result in the emission of air pollutants, the discharge of water pollutants, the creation of noise, and disturbance of the surface. Tailings ponds can be the source of fugitive dust.1 This chapter is not intended to provide a detailed discussion and analysis of laws and regulations dealing with environmental protection. Rather, its purpose is to provide the engineer with a basic awareness of the existence and nature of such laws and regulations, as well as the procedural requirements that must be followed in complying with them. The body of law relating to environmental protection has grow" very rapidly and should continue to do to for some time. Because many of the laws have been enacted recently, numerous court decisions are being rendered to resolve disputes over their interpretation. Hence, the reader is cautioned to be alert for subsequent modifications of statutes and regulations, and new case law. Rules and regulations pertaining to environmental protection are implemented at all governmental levels. The most widely known laws are those enacted by the federal government that have nationwide applicability. However, separate requirements exist in each state, county, and municipality. Because of their general applicability, federal laws are discussed most extensively in this chapter. Ownership of the property is the most significant factor considered in ascertaining what rules govern the conduct of an operation thereon. If the land is held under lease, reference to the lease terms must be made in the first instance to determine what obligations must be met in order to prevent default and possible loss of the property. If the land is held under a lease from the federal government, the operator is subject not only to compliance with the lease terms, but also to a large body of laws and administrative regulations that pertain generally to the conduct of mining operations on land held under federal leases. Although operations on unpatented mining claims, the legal title to which remains in the federal government, are not subject to the same rules and regulations that are applicable to operations conducted pursuant to federal leases or permits, they soon will be governed by a special set of regulations that provide for protection of surface resource.2 Operations conducted on lands leased from a state usually are subject to numerous environmental protection requirements specified in the lease terms, in addition to rules and regulations promulgated by the state agency having jurisdiction over mining on state lands. Operations conducted on privately held lands are subject to fewer such requirements. Leases from private parties sometimes have environmental protection and reclamation requirements written into them, but generally to a far lesser extent than governmental leases. Operations conducted on properties owned by the operator are subject only to those laws and regulations that have general applicability without regard to land ownership. COAL SURFACE MINING CONTROL AND RECLAMATION ACT OF 1977 Introduction On Aug. 3, 1977, the Federal Surface Mining Control and Reclamation Act of 1977 was signed into law.3 It governs coal-mine operations on private lands, as well as on public lands. The Act is pervasive in its scope and is extremely long and complex. The basic purpose of the Act is to control and minimize the environmental effects of surface coal mining. Surface coal-mining operations are defined as activities conducted on the surface of lands in connection with a surface coal mine and surface impacts incident to an underground coal mine.4 The Act is administered by the Secretary of the Interior through a new agency named the Office of Surface Mining Reclamation and Enforcement.5 The Act contains detailed environmental protection standards and reclamation requirements, and it establishes a permit system for all surface coal-mining operations. Mining in certain areas and under ceri-in conditions is restricted or prohibited, and a mechanism for enforcement by the states is provided. Stiff penalties are provided in the event of noncompliance. Implementation Schedule Nonfederal Lands: As required by Section 501 of the Act, interim regulations setting mining and reclamation performance standards based on and incorporating standards set out in Section 502(c) were adopted effective Dec. 13, 1977.6 They will. be incorporated as amendments to Chapter VII of Title 30, Code of Federal Regulations. Permanent regulatory procedures for surface coal-mining and reclamation operations performance standards, which were directed to be promulgated by Aug. 3, 1978, were published in proposed form on Sept. 10, 1978. 7 They govern surface coal-mining operations in any state until a permanent state or federal program is adopted. As of Feb. 3, 1978, all new operations, and as of May 3, 1978, all existing surface coal-mining operations, on lands on which such operations are regulated by a

Jan 1, 1982

By Richard L. Bullock, Bruce Kennedy

Vince Lombardi once said, "Practice doesn't make perfect, perfect practice makes perfect." When it comes to building a mine that will operate at the optimum level for the set of geologic conditions from which it was developed, Lombardi's remark might be paraphrased to describe the problem: planning won't guarantee the best possible mine operation unless it is the best possible mine planning. Any sacrifice in the best possible mine planning introduces the risk that the end results may not reach the optimum mine operation desired. This section addresses many of the factors to be considered in the initial phase of mine planning. These factors have the determining influence on the mining method, the size of the operation, the size of the mine openings, the mine productivity, the mine cost, and, eventually, the economic parameters used to determine whether or not the mineral reserve even should be developed. A little-known fact, even within the metal-mining community, is that room-and-pillar mining accounts for most of the underground mining in the united States. According to a 1973 study on noncoal mining (Anon., 1974), more than 76% of the producing mines [of over 1089 t/d (1200 stpd) capacity] produced approximately 70 000 000 t (77,000,000 st) or 60% of the nation's underground tonnage of material by room-and-pillar mining. That same year, 96.8% of the nation's under- ground coal mines produced 262 950 000 t (289,911,000 st) of coal extracted from room-and-pillar mines (Anon., 1976). Thus, nearly 333 000 000 t (367,000,000 st) of the United States' raw material is produced from mines using some form of the room-and-pillar mining system. Because approximately 90% of all mining in the United States is done by some variation of room-and- pillar mining, it is appropriate to give special emphasis to the effects of the various elements of mine planning on room-and-pillar mining. The relationship of these elements to other mining methods will become apparent as the elements are described in later sections herein. TECHNICAL INFORMATION NEEDED FOR PRELIMINARY MINE PLANNING Assuming that the reserve to be mined has been delineated with diamond-drill holes, the items listed in the following paragraphs need to be established with respect to mine planning for the mineralized material. Geologic and Mineralogic Information The geologic and mineralogic information needed includes the following: 1) The size (length, width, and thickness) of the areas to be mined within the overall area to be considered, including multiple areas, zones, or seams. 2) The dip or plunge of each mineralized zone, area, or seam, noting the maximum depth to be mined. 3) The continuity or discontinuity within each of the mineralized zones. 4) Any swelling or narrowing of each mineralized zone. 5) The sharpness between the grades of mineralized zones within the material considered economically minable. 6) The sharpness between the ore and waste cutoff, including whether this cutoff can be determined by observation or must be determined by assay or some special tool; whether this cutoff also serves as a natural parting resulting in little or no dilution, or whether the break between ore and waste must be induced entirely by the mining method; and whether or not the mineralized zone beyond (above or below) the existing cutoff represents submarginal economic value that may be- come economical at a later time. *7) The distribution of various valuable minerals making up each of the minable areas. 8) The distribution of the various deleterious minerals that may be harmful in processing the valuable mineral. 9) Whether or not the identified valuable minerals are interlocked with other fine-grained mineral or waste material. 10) The presence of alteration zones in both the mineralized and the waste zones. Structural Information (Physical and Chemical) The needed structural information includes the following: * 1 ) The depth of cover. 2) A detailed description of the cover including: the type of cover; * the structural features in relation to the mineralized zone; * the structural features in relation to the proposed mine development; and * the presence of and information about water, gas, or oil that may be encountered. 3) The structure of the host rock (back, floor, hanging wall, footwall, etc.), including: * the type of rock; * the approximate strength or range of strengths; * any noted weakening structures; * any noted zones of inherent high stress; noted zones of alteration; the porosity and permeability; * the presence of any swelling- clay or shale interbedding; the rock quality designation (RQD) throughout the various zones in and around all of the mineralized area to be mined out; the temperature of the zones proposed for mining; and the acid generating nature of the host rock. 4) The structure of the mineralized material, including all of the factors in item 3 plus: * the tendency of the mineral to change character after being broken, i.e., oxidizing, degenerating to all fines, recompacting into a solid mass, becoming fluid, etc.; * the siliceous content of the ore; the fibrous content of the ore; and the acid generating nature of the ore. Economic Information The needed economic information includes: *1) The tons of the mineral reserve at various

Jan 1, 1982

By Edie Trimmer

Utah statute defines sovereign lands as "those lands lying below the ordinary high water mark of navigable bodies of water at the date of statehood, and owned by the state by virtue of its sovereignty." The lands within the bed of Great Salt Lake (GSL) are, by this definition, sovereign lands, acquired at statehood in 1896 in accordance with the "equal footing" doctrine, granting each state control and ownership of navigable waters and the lands underneath those waters within its borders. Under public trust doctrine, the state, as trustee for the people, bears responsibility for preserving and protecting the right of the public to use of the waters for navigation, commerce, fishing, recreation, and wildlife habitat. Also by statute, sovereign lands are defined as "state lands," to be managed by "multiple use sustained-yield principles." The Division of Forestry, Fire and State Lands is given management authority for sovereign lands and, as manager, has responsibility to prepare comprehensive plans, initiate studies of the lake and its resources, implement comprehensive plans through state and local entities, and coordinate the activities of various divisions within the Department of Natural Resources (DNR). The Division of Forestry, Fire and State Lands also has responsibility for management of mineral leasing on sovereign lands. The many resources on the lake--water, minerals, wildlife, recreation, archeological and historical values--are managed by as many state agencies which occasionally creates conflicts. The brines of GSL contain several ions that crystalize into valuable minerals during evaporation. The major ions in the lake are, in order of relative abundance, chloride, sodium, sulfate, magnesium, and potassium. Mineral products which are currently extracted from lake brines are sodium chloride, magnesium chloride brine which can be sold as flake magnesium chloride or further processed into magnesium and chlorine gas, and potassium sulfate. Mineral products which have potential for extraction include gypsum, sodium sulfate, and trace amounts of lithium, boron, and bromine. The GSL contained an estimated 4.3 billion short tons (st) (3.9 billion metric tons [mt]) of dissolved salts in 1998. Utah Geological Survey (UGS) estimates of the dissolved salt content in GSL have fluctuated from 4.0 to 5.5 billion st (3.6 to 5.0 billion mt) due to the dynamic conditions in the lake as salts are precipitated and redissolved, and due to the diversion of brines from GSL, such as the West Desert Pumping Project. The lake has four areas of varying salinity, separated by dikes or other man-made structures: north arm and Stansbury Bay brines at near saturation (25 to 27 percent total dissolved solids [TDS]); the main body of the south arm with concentrations ranging from 7 to 15 percent TDS as lake elevations fluctuate; the waters in Farmington Bay at approximately 3 to 5 percent TDS; and Bear River Bay at <1 to 7 percent TDS. The percent TDS in Bear River Bay fluctuates with lake level, and changes in Bear River inflow. The transfer of salts from the south arm to the north arm has raised questions about the viability of the mineral and brine shrimp industries. The UGS and the U.S. Geological Survey (USGS) continue to monitor salinities at designated sites on the lake to document changing lake salinity. A recurrent theme is that placement of dikes and diversions can have significant and rapid impacts on various conditions in the lake. Hydrocarbon resources on the lake are significant, but presently undeveloped. The hydrocarbons are low gravity (4 to 9 degree API) and tar-like, contain high nitrogen concentrations, and up to 12 percent sulfur. The unusual characteristics of the oil have been the subject of studies by chemists at Weber State University and University Louis Pasteur de Strasbourg. However, these resources are difficult, and at present, uneconomic to extract using current technology because of the nature of the hydrocarbons, and production in "an offshore, highly saline environment." Oolitic sand deposits make up many of the beaches and shorelines around the lake. Because of their high calcium carbonate content, oolites have been used by Magnesium Corporation of America (MagCorp) and its predecessors for acid neutralization and dike construction. Oolites are also used in very minor amounts in flower drying. The Utah Division of Oil, Gas and Mining reports up to 130,000 st (118,000 mt) mined annually by MagCorp from U.S. Bureau of Land Management (BLM) lands adjacent to GSL. Currently, there are twelve producing mineral leases which generated slightly more than $1,000,000 in royalties during calendar year 1998. IMC Kalium Ogden Corp. (IMC Kalium) produces potassium sulfate and magnesium chloride from brines concentrated through solar evaporation in Bear River Bay and Clyman Bay. By-product sodium chloride is transferred to IMC Salt, which packages and sells the salt. MagCorp produces magnesium metal from brines concentrated in Stansbury Bay. Cargill Salt produces sodium chloride from brines provided by MagCorp under a lease agreement. Morton Salt produces salt at the southeast end of Stansbury Island. Lastly, North Shore Limited produces cocentrated brines for use in dietary and mineral/vitamin supplements near Spring Bay in the north arm of the lake. Producers of magnesium, potash, and salt from GSL contribute significantly to the value of metals and industrial minerals in Utah. Together these companies contribute approximately $240 million in gross value, or 18 percent of the value of the state&apos;s nonfuel mineral production. Most of this production is exported.

Jan 1, 2001

By Robert H. Merrill

INTRODUCTION In most cases, the design of an underground mine is based upon the premise that the ground either will cave or will be stable. This chapter concerns the design of a mine in ground that will cave readily or with some as¬sistance, such as by long-hole drilling and blasting. Some of the more widely used caving systems of mining are panel caving, block caving, sublevel caving, and large pillar recovery. Some of the less widely used systems are glory-hole, top slicing, and induction caving. Al¬though the common practice of pillar robbing is not usually considered to be a caving system, this subject will be treated as a part of this chapter. BASICS OF CAVING Caving systems are most successful in ground that will cave in sizes that will flow through openings and grizzlies, and will easily load in cars or on belts for haul¬age. The ground most likely to cave well is highly frac¬tured and contains breaks, flaws, or other discontinui¬ties that form planes of weakness. Also, caving action can be greatly enhanced if the host rock itself is low in compressive, shear, and tensile strength. Ideally, a cav¬ing system of mining is best employed when the criteria for caving is a feature of the ore body and the develop¬ment drifts, haulageways, and drawpoints can be mined in a highly competent rock beneath the mineralized zone. However, the development is often in the same, or similar, fractured rock and the openings require sub¬stantial artificial support to assure stability. Several clues can be assembled to identify potential caving ground; however, for borderline cases, no sure method has been devised to date. The diamond-drill cores taken for exploration can provide an excellent clue provided drilling is performed carefully by experienced drillers. For example, if the ground is cored in such a manner that the breaks in the core are caused more by failure of the rock than by whipping core barrels, plugged drill bits, or other drilling causes, and the intact core lengths are consistently long [say, 0.6 to 3 m (2 to 10 ft) of unbroken core], there is little reason to believe the ground will cave without considerable as¬sistance. This is especially true for rocks with compres¬sive strengths above 34.5 MPa (5000 psi) and tensile strengths above 2.1 MPa (300 psi). On the other hand, if core recovery is low (below 80%) and the recovered ore is broken in small pieces and the breaks are along obvious weaknesses in the rock, the chances are excel¬lent that the ground will cave. This is true even when the rock between the defects has high compressive and tensile strength. Another clue has already been mentioned, that is, the measurement of the physical properties of the rock and the natural planes of weakness or defects in the rock. The planes of weakness in the rock can often be detected from outcrops, cores, or other exposures of the rock under consideration. Some rock types are known to be strong and will sustain large, unsupported open¬ings and would be difficult to cave intentionally. Yet the same rock type can also contain unbonded or weak planes of weakness or fractures, and in these locations the rock would undoubtedly cave with little assistance. Therefore, although the inherent strength of the rock is a factor in caving, the natural defects in the rock are more often the deciding factor. DESIGN CONCEPTS For the most part, the design of openings for caving ground is a problem of the interaction of openings over a relatively large area of the mine. To illustrate, Fig. 1 is a simplified section of a series of openings along the grizzly level or draw level of a block caving or panel caving development, and above this opening is a simpli¬fied section of a room-and-pillar arrangement on the undercut level. At this stage of the development, the stresses around the openings on the grizzly level are only moderately influenced by the openings on the undercut level and vice versa. Therefore, the stresses around the openings are approximated by the stresses around single or multiple openings in rock, the values of which are de¬scribed in the literature (Obert, Duvall, and Merrill, 1960; Obert and Duvall, 1967). Once the pillars on the undercut level are blasted (Fig. 2), the situation changes abruptly. The undercut opening (prior to caving) now can be approximated as an ovaloidal opening above the grizzly drifts and this opening tends to shield the vertical stress field. As the caved stage is drawn the stope approximates a much larger rectangular or square opening filled with rock, and if the rock is not sustaining a major portion of the stress field, this opening can be considered (for en¬gineering purposes) to be empty and the stresses that interact between the larger and the smaller openings take on a totally new perspective (see Fig. 3). Next, let the material cave to the surface, and let the caving ma¬terial sustain some stress, but much less than if the ma¬terial were intact. This condition is similar to a soft inclusion in a rigid body and has been treated in the literature (for example, Donnell, 1941). At this point in time, the grizzly drifts are subjected to the stress con-

Jan 1, 1982

By S. L. Bessinger

Introduction Extensive testing in the past ten years has shown that where a uniform natural gamma background is present in the strata bordering a seam, the thickness of the boundary coal left in place after mining can be determined by measuring the attenuation of that radiation (Nelson and Bessinger, 1989). Measurements made by the authors in underground mines in Pennsylvania, West Virginia, Ohio, Illinois and Kentucky and by others in Wyoming and New Mexico have shown the presence of such a gamma background (Nelson. 1989). Natural gamma coal-thickness sensors of several configurations have been tested in mines owned and operated by the Consolidation Coal Company (Consol) in Pennsylvania and West Virginia (Nelson and Bessinger, 1988). This paper describes the installation of a natural gamma coal-thickness sensor on an operating continuous miner. Previous tests had shown that the NGB-1000 coal-thickness sensor manufactured by American Mining Electronics, Inc., of Huntsville. AL, is an accurate, mine-worthy instrument. This large gamma detector consists of a sensing head and a control panel. The sensing head contains thallium-doped, sodium iodide scintillating crystal, which is coupled to a photomultiplier tube. The control panel contains the electronic components required for calibration, count conversion and display to the operator. Methods Conditions at a Consol mine in northern West Virginia require that 10 to 15 cm (4 to 6 in.) of coal be left at the roof boundary of continuous miner development sections. This roof coal is required because the shale of the immediate roof is friable and unstable. In the past, operators have used a dirt band that is usually visible near the top of the seam as a guide in maintaining the proper cutting horizon. However, this is not always reliable. Earlier observation showed that the actual thickness of the coal left on the roof varied widely; further, it was noted that occasional, accidental excursions into the immediate roof required supplementary roof control measures, such as installation of planks or center bolts. Thus, it was concluded that operators needed a better source of guidance for control of the cutting horizon, and a roof-coal thickness sensor was scheduled for installation. The NGB-1000 sensor was installed on a Joy 12CM10 continuous miner in June 1988. The sensing head was mounted on the cutter boom of the miner, and the control panel was mounted in the operator&apos;s cab. Power for the sensor was initially derived from an intrinsically safe battery power supply. Initial measurements with the sensor showed that the calibration was the same as that used in earlier tests at two other mines, indicating the uniformity of the natural gamma background above the Pittsburgh seam. Operating personnel were initially skeptical of the instrument&apos;s accuracy, and were hesitant to use its readings as a guide in maintaining a proper cutting horizon. Because gamma attenuation, the instrument&apos;s operating principle, is somewhat abstract, attempts to demonstrate the instrument&apos;s accuracy by explaining that principle were generally ineffective. It was found, however, that an operator could usually be convinced of the usefulness of the instrument by placing a large piece of coal of fairly uniform thickness over the instrument&apos;s sensing head and allowing the operator to see that the instrument reading increased by an amount very near his estimate of the thickness of the piece. The mine was provided with seven battery power supplies and a charging station. The charging station was kept in the lampman&apos;s office, and the mechanic on each shift was instructed that he was responsible for two battery power supplies each day: a freshly charged one to be taken in at the beginning of his shift and a depleted one to be brought out at the end. This system worked well for a few weeks, but eventually some battery power supplies were left in use so long that their batteries were discharged too deeply to allow recharging. In addition, transport and recharging of the batteries represented an additional task for the mechanics, who were already very busy. Consequently, a request was filed with MSHA to allow the sensor to be powered through intrinsic safety barriers by an electronic power supply connected to machine power. The permit was granted, and the sensor was connected to machine power. After the sensor was connected to machine power, the only operating problem experienced was occasional failure of cables. A supply of the required cables was made and delivered to the mine so damaged cables could be quickly replaced. Much of the cable damage could be eliminated by slight modifications to the miner during a rebuild, so that cables could be installed in more protected locations. After the sensor had been in operation for about two months, a survey was made to determine its effect on continuous miner operations. In previous research, coal thickness measurements made in 88 locations by the natural gamma method were compared to measurements made in the same locations by observing drill cuttings and by inspections of drill holes with a borescope. That research showed that the gamma method is at least as accurate as the other two methods (Nelson and Bessinger, 1989) and is also much easier to use. The object of the survey described here was not to assess the accuracy of the natural gamma measurements. but rather to determine the effectiveness of the sensor output as a guide for the operator in maintaining control of the cutting horizon. Thus a smaller, hand-held gamma detector

Jan 1, 1992