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By D. M. Range

Exxon is reclaiming an 80-hm2 (200¬acre) Wyoming uranium tailings basin in accordance with US Nuclear Regula¬tory Commission (NRC) requirements. The reclamation plan is designed to control the tailings and their radioactive emissions for at least 1000 years. Rigorous modeling and analyses of potential erosion by wind and water were used in planning. The use of riprap was minimized and the need for long¬term reclamation maintenance was eliminated. The resulting design is simi¬lar to regions stable from erosion cre¬ated by nature. The concepts used in the design and implementation have wider reclamation application. Highland history Exxon operated the Highland uranium operations from 1972 until mid-1984. Open pit and underground operations mined the roll front, sandstone ore body. The 2.5 kt/d (3000 stpd) mill employed conventional crushing, grinding, sulfu¬ric acid leach, CCD, SX, precipitation, and drying circuits. The mill produced 10 Mt (11 million st) of tailings. An¬other operator began commercial scale in situ leach operations in 1988. The tailings basin was constructed by deposition of mill tailings in a natural ravine dammed with a 30-m (100-ft) high by 1050-m (3400-ft) long clay core earthen dam built entirely downstream of tailings. Figure 1 shows the layout of the tail¬ings basin in 1984. Tailings were de¬posited along the dam face on the east side of the basin as well as in the west end and along the south edge until the early 1980s. This left a deep water pond in the middle of the basin. The deep pond was largely filled with tailings during the last four years of mill opera¬tions by the construction of two tailings causeways and the operation of a float¬ing tailings line. The pond water volume was mini¬mized by recycling tailings water to the mill and by operating a 32-L/s (500¬gpm) evaporation enhancement spray system during the last few years of mill operations. Regulatory requirements Several conceptual tailings basin rec¬lamation plans were prepared during the mine life. In September 1984, Exxon began preparation of a detailed plan in response to an NRC request. Water, Waste and Land Inc. (WWL) of Fort Collins, CO was selected to help pre¬pare the plan that was submitted to the NRC in December 1984. The design had to meet the require¬ments found in NRC regulations (NRC, 1985). These were based on regulations prepared by the Environmental Protec¬tion Agency (EPA, 1983) at the man¬date of federal law (UMTRA, 1978). The NRC regulations contain strict primary standards and a somewhat more flexible set of secondary standards. The primary standards include requiring that: • radiation from tailings be reduced to background levels; • radon gas emitted by the tailings be reduced to 20 pCi/m2/s from the surface of the reclaimed tailings pile; • tailings not be dispersed into the environment; and • these conditions endure without routine maintenance for at least 200 years and for 1000 years to the extent reasonably attainable. The lengthy time requirements pre¬sented a large engineering challenge. Radon is a radioactive, chemically inert gas with a 3.8 day half-life. It is released from uranium tailings by the radioactive decay of radium, which, in turn, is a radioactive decay product of uranium. Radon emissions from ura¬nium tailings can be controlled with an earthen cover thick enough that the radon gas decays to its solid, short-lived radio¬active daughters before the radon dif¬fuses through the cover material into the air.

Jan 1, 1989

By Peter Joralemon

Introduction The Sierra Nevada Foothill Gold Belt in California extends from Mariposa north for more than 420 km (200 miles). It includes the Grass Valley and Alleghany districts in the zone of Northern Mines and the famous Mother Lode to the south. The gold belt has produced more than 3.1 kt (100 million oz) of gold since its discovery in 1848. This remarkable production, at the present gold price, would be valued at more than $40 billion. The Law of Gambler's Ruin supposes that averages are based on systematic rather than random variations. It emphasizes the vast differences between finite and infinite probability. By applying this Law to this figure, it can be seen that the belt has produced an average of 15.5 t (500,000 oz) of gold per running mile. Many geologists believe, with some justification, that less than half the gold in this belt has been mined. Sixty per cent of the past production has come from lode deposits, the balance from Tertiary and Recent placer gravels. This paper will concern itself with the developing lode mines along the Sierra Nevada Foothill Gold Belt. As a further and final note regarding placer mining along this belt, Mark Twain suggested that placer miners, having lost their senses, had contributed liberally to the population of California's insane asylums. After a 40-year lull with little or no gold production, the lode mines of this famous belt are showing signs of renewed activity. Two open pit mines, the Sonora (Jamestown) and Carson Hill, are now in operation at substantial scales. Two others, the Goldenbell (Pine Tree-Josephine) and the Royal-Mountain King, are being readied for production on equally impressive scales. All are in the southern part of the Mother Lode. Exploration farther north, at Sutter Creek, Drytown, and Georgetown, is showing encouraging results in developing deeper lode deposits. This new activity has been sponsored to a large extent by junior Canadian mining and exploration companies. Most of the established American mining companies have looked the other way, not participating in this important development. One major reason for domestic disinterest in this opportunity is there is little chance for heap leachable open pit ore in this belt, excluding Carson Hill. And heap leaching has become fashionable with American mining companies today. Also, this great old gold belt must have been worked out, for why else is there so little activity here - Catch 22 logic. Another reason is that if a new deposit of size and grade similar to the mines in the past was found, high costs would preclude a profitable operation. And, finally, California environmentalists would not permit mining, even if a rich deposit were found. These may all be pefectly valid reasons not to explore for gold along this belt. But a look at the history of gold mining in California may suggest that the future for exploration and development of the gold resources in the Sierra Nevada Foothills can be greater than the past. It was during the Depression of the 1930s that this gold belt saw its maximum activity, with more than 3000 mines in operation. Some were father and son operations producing only a few tons of ore a day. Other mines employed hundreds of miners to hoist more than 900 t/d (1000 stpd). Some ores were of marginal grade. Others, like the Sixteen to

Jan 7, 1987

By G. J. Jansen

Introduction Coal petrography is the microscopic study and description of coal and other organic substances in terms of the microscopically recognizable organic constituents and of the rank of the material. Petrography studies of coal and other organic substances are useful in themselves, and have a number of practical applications. Some of these include determining coking quality of coal, estimating yields of methane from coal beds, and assisting in coal preparation. Petrography studies can also help in mine planning, in source rock studies, in combustion problems, and in synthetic fuel applications. Most recently, petrography has been found to be useful in studying sediment-hosted ore deposits. Microscopic constituents-macerals Under the microscope, coal is a heterogeneous assemblage of the coalified remains of woody matter and plant remains from the swamp or bog from which the coal was formed. The individually recognizable materials making up the organic mass are termed macerals by petrographers. They are the organic equivalent of the minerals that make up rocks. Coal petrographers can examine coal in two ways. Those interested primarily in the microscopically transparent macerals or in the study of very low rank materials, lignites for examples, like to use transmitted light microscopy where light passes through a microsection about 3 to 6 µm (0.0001 to 0.0002 in.) thick (Fig. 1). The dominant red color of the section in Fig. 1 is inherent, not a stain or produced by a filter. Distinct suggestions of textures in the reddish constituents can be seen. The red material is vitrinite and is thought to be the coalified remains of woody tissues. The yellow, lenticular substance is exinite. It is what remains of the waxy coating on a spore. Also seen are white dots of mineral matter and black opaque material that is better seen and described in polished sections. A polished microsection is prepared by grinding and polishing to a mirror polish a compact of coal and embedding resin or a coherent chunk of coal. The resulting microsection is then examined in a reflecting microscope. High magnification, 200X to 400X, and an oil immersion lens are needed to resolve fully the macerals of the organic matter so prepared. Figure 2 shows two macerals, an even gray vitrinite, unstructured and fusinite, a whitish, more or less cellular constituent. Fusinite is thought to represent more or less completely carbonized cellular material. Coal petrographers appreciate that coal is a heterogeneous material. Figure 3 is a view of another section. The generally structureless gray material is vitrinite. The bean-shaped dark constituents are exinites. The long, narrow, somewhat scalloped dark constituent is cutinite. Cutinite is thought to be derived from leaf coatings. Some coal petrographers would describe the large white, structureless mass as micrinite. Others might prefer the term macrinite. A few fragments of fusinite can also be seen. Figure 4 shows a coalified fungal remain, termed sclerotinite. It is surrounded by vitrinite and crushed fragments of fusinite all floating in a sea of embedding resin. Sometimes the microscope reveals the great heterogeneity of coal even on the microscope scale. Figure 5 shows a section with semi-fusinite, inertodetritinite, vitrinite, and micrinite. The nature of the heterogeneity may be judged by noting that the long dimension of the photomicrograph is about 70 µm (0.0027 in.). Measurement of rank Hoffmann and Jenkner (1932) discovered that measuring the reflectance of the vitrinite constituent of coal in a polished section gave a good measurement of the rank of the coal. In practice, a photomultiplier is used to mea-

Jan 9, 1987

By P. N. Eksteen, G. E. Greig

The South African mining industry currently installs about 6 km per month of plastic pipes for conveying compressed air and water. Plastic pipes have a number of desirable properties and are favoured by many mining engineers because of their low mass, which facilitates ease of erection and assembly, and their excellent corrosion resistance and durability. However, they are combustible, and if involved in an underground fire can contribute to the overall toxicity of the combustion gases emitted. Furthermore, plastic pipes have the potential to conduct a fire from one part of a mine to another, and, by so doing, exacerbate an already hazardous situation. The limitations of small scale tests and the need to establish a large facility where underground conditions could be simulated was recognized by the South African mining industry. This paper briefly describes the Industry's full scale fire test facility and presents the conclusions drawn as a result of a number of full scale combustibility tests conducted on pipes manufactured from unplasticized Poly Vinyl Chloride (uPVC) and High Density Polyethylene (HDPE). Based on these findings, some practical solutions to prevent the spread of fire along the length of pipe columns of various pipe diameters and suspension configurations are presented. Recently some new and exciting developments in the field of intumescent coatings have occurred which are proving effective in minimizing the incidence of fire underground on South African mines. The potential benefits of these coatings are briefly discussed. INTRODUCTION Plastic pipes used underground on mines are predominantly in the form of protective cladding on insulated chilled water pipes and as service pipes for the conveying of water and compressed air to the workings. Plastic pipes have found wide use in the aggressive environmental conditions encountered underground due to their favourable properties relative to steel; these include being lightweight, easily installed and virtually immune to corrosion. However, plastic pipes are inherently combustible and can contribute to the overall hazard in three ways: (i) they are easily ignitable and can provide the initial fuel for a fire. (ii) once ignited, they can propagate a fire along their length and by so doing ignite other combustibles in their path, thereby contributing significantly to the overall conflagration. (iii) they can contribute to the overall heat, smoke and toxic gas emissions from a fire. While conscious of the value of small scale fire tests, the South African mining industry recognized the need for full scale testing to determine the combustion characteristics of plastic pipes under conditions closely approximating those underground and, as a consequence, a full scale fire tunnel test facility was commissioned with the objective of being able to (a), accurately evaluate the potential hazard of the various types of the plastic pipes found in mines, (b) use this information to suggest suitable fire engineering practices to minimize the hazard and (c), to practically verify the efficacy of the practices suggested. THE FIRE TUNNEL The fire tunnel is situated on a mine in the Western Transvaal and is 40 m long with a cross section of 3 m x 3 m which is typical of many haulages currently in use in South African gold mines. The ventilation drawn through the tunnel is provided by a 110 kW centrifugal exhaust fan and air velocities can be varied from 0 to a maximum of 5,8 m/s.

Jan 1, 1993

By J. M. W. Mackenzie

Three major developments have dominated the recovery of gold from ores, viz; 1846 Elsner's studies of gold solubility in aerated alkaline cyanide solutions 1887 MacArthur and Forrest's process based on zinc precipitation 1980+ Large scale development of the CIP process The brevity of this list for a metal whose extractive metallurgy has been as widely studied as any metal reflects the fact that the chemistry of the gold recovery process, as opposed to the engineering has seen few significant breakthroughs in over a century of application. to The purpose of this paper is introduce a new chemistry for gold recovery which may, in the future, add significantly to the technology of gold recovery. This new chemistry is based guanidine functionality - (R2R2N) 2 C = N R2 on the In developing the use of the guanidine functionality for the recovery of gold as the aurocyanide anion from leach solutions and pulps, Henkel has, as its objective, the development of an effective gold selective ion exchange resin. As will be seen, the initial investigation of the properties of the guanidine functionality in the solvent form for gold recovery lead to some significant metallurgical developments of their own but, the foremost objective was, and remains, the development of a resin for gold recovery. As a part of this development project, Henkel has entered into association with Mintek, South Africa, and this ongoing association has the potential to develop RIP/RIS technology for the gold industry. More details on Henkel's gold recovery technology and Mintek's role in this programme are given in (1). an The application of ion exchange resins to gold recovery is not new and many reviews have been published on the advantages of this process, and on the current state of the art of gold recovery by ion exchange resins. (2)(3)(4) Widespread application of ion exchange resin based processes for gold recovery has, however, not occurred to date. Fleming and Cromberge (5), described some significant pilot plant RIP trials in South Africa and the partial success of these pilot plant trials held out the tantalising possibility of a breakthrough in gold recovery based on RIP technology. Despite this pioneering work, only two really significant applications of RIP technology for gold recovery exist today. These are the Golden Jubilee Plant in the Eastern Transvaal South Africa and the widespread, although in some respects technically backward, use of ion exchange for gold recovery in the former USSR. There are two main reasons for the current scarcity of RIP and RIS plants used for gold recovery. These are: 1. 2. The perceived problems of engineering RIP plants The lack of a resin having suitable chemistry for both gold loading and gold elution The problems of engineering RIP plants have been largely overcome in the last 25 years. Much of the development in RIP plant design took place in the uranium industry during the rapid expansion of that industry in the mid 1970's. In addition, the development of CIP engineering systems has lead to the development of in-pulp solid extractant based metallurgical plant designs which can, with limited modification, be adopted to RIP. The lack of a resin having suitable chemical properties has almost certainly been the main reason for RIP/RIS not being introduced into the gold industry over the past decade. Attempts to use RIP/RIS have focussed on either conventional we ak base or conventional strong base resins. The weak. base resins have the disadvantage that they

Jan 1, 1993

By Richard L. Bullock

INTRODUCTION TO COAL MINE PRACTICES To illustrate the trends and the proportions of the various methods used to mine coal underground, Fig. 1 is useful. However, of all the underground coal mining practices in the US, room-and-pillar mining still accounts for 96.5%. This percentage is divided as follows: conventional, machine cut-33.0%; conventional, hand cut -1.7% : and continuous mining-61 .8%. As was stated in Sec. 1, Chap. 7, the term "development" means different things to a coal miner and a metal miner. Though some of the ideas and concepts presented in the first two chapters of this section are valid for all types of mines, they apply particularly to mines using noncoal methods. In coal mines (and many other mines using coal mining systems) development refers to that phase of the mining cycle which advances several entries through the virgin seam, forming pillars by cross- cutting between the parallel entries. Because in the past, coal seams (as well as seams of potash, trona, oil shales, gilsonite, salt, and some uranium) have been much more uniform in their thickness, continuity, and quality, and because they often cover large areas, it has been possible to develop the room-and-pillar mine completely from a preplanned, systematic approach with a minimum of surface drilling information. As the better coal seams of the midwest and east are being depleted (or eliminated because of high sulfur), seams of poor quality and continuity will have to be mined. Also, seams that are not nearly so well known at greater depth will have to be taken as well as some of the steeply pitching and folded seams of the west. The American coal miner is only now feeling the need to do a much more complete job of prospecting the geology of the coal seams than he has ever done before. This is the only way that the risk of unknown rock conditions can be minimized. In many ways, surprises in the rock or coal seams are much more of a disaster to the coal miner developing a room-and- pillar mine than to his metal mining counterpart. Be- cause of the multiple entries, the complex set of the Federal Coal Mine Health and Safety regulations, and the totally dedicated design of coal mining equipment, coal mining systems do not have the flexibility to easily overcome changes in the seams mined or the rocks around them. Thus, if a wide parting of rock shows up unexpectedly in the middle of a coal seam entry being milled away by a continuous miner, it will either have to slowly "chew" through it, convert to a conventional mining system, or stop that particular entry. Likewise, if suddenly the coal disappears completely and in its place are the remains of an old creek bed, the "want" or "wash out" as it's called will usually have to be mined right along with the coal, by the same methods and with the same soft rock equipment, or the entry will have to be abandoned. For reasons such as these, coal seams today should be drilled from the surface just as thoroughly as the ore bodies in metal mines. Having once determined a more precise location of where the coal seams are located and which are suitable for mining, the engineer must still take a systematic approach to planning the mine as completely as possible. The initial series of entries must be planned to be driven either from the side hill portal or from the bottom of the shaft or slope. In either case, they should usually intersect the area close to the midpoint of the mass to be mined.

Jan 1, 1982

By Ted D. Maki, Joseph B. Taylor

INTRODUCTION FMC's Paradise Peak mine, located 13 Ian (8 miles) south of Gabbs, Nevada, became opera¬tional in April of 1986 (Figure 1). It was designed and built by Davy McKee, who was instrumental in equipment design and selection. It is the 7th largest gold deposit in the United States with defined reserves of 10.9 mt (12 million st) containing 34 million gms (1.1 million tr oz) of recoverable gold and 933 million gms (30 million tr oz) of recoverable silver. The mine extracts ore by the open-pit method, taking advantage of a 1.5:1 stripping ratio. The mill operates at a 3600 mt/day (4000 st/d) capacity. Crushing is done in three stages to achieve an ore size of minus 0.635 cm (0.25 inches). Grinding further reduces the ore to 85% minus 100 mesh. The ground ore is treated with cyanide in agitated leach tanks and then washed in the counter-current decantation (OCD) thickeners. Zinc is added to the clarified, deaerated pregnant solution to precipitate the precious metals. The precipitate is acid digested to eliminate excess zinc, filtered and retorted to drive off contained mercury. The retorted precipitate is then fluxed, melted and poured through cascade molds. Dore bars are cleaned for shipment and the slag is sent to an in-house slag treatment system. ELECTROSTATIC SEPARATION The dry electrostatic slag treatment system at Paradise Peak is the first installation of its kind. Electrostatic separation has been widely used in mineral processing since the early 1950's. A brief discussion of the theory behind the process is helpful to those not familiar to electrostatic separation. Charging and sepa¬rating slags, dry minerals or other materials by ion bombardment is the most common form of electrostatic separation. Millions of tonnes of minerals are processed each year by this method. In an ion bombardment separation, granular material is fed onto a grounded metal cylinder (or roll) and charged by a corona-producing electrode placed above the roll's surface (Figure 2). While both conductors and non¬conductors become charged, only the conductor is able to lose its charge. The charged non¬conductor, as it rests on the roll's surface, "sees" an oppositely charged image of itself in the metal surface. It is attracted to the image charge, becomes electrostatically pinned to, and moves with the roll's surface. The conductive particle also sees an image and is attracted to it. But upon touching the roll's surface, it discharges rapidly to the grounded surface and is thrown free from the roll's surface with a projectile motion. In the case of precious metal slags, the con¬ductive particles would be metallic prills of dore metal; and the nonconductive particles would be the slag, free of metal. Middling grains, generally, are in the form of a metallic prill encased in or incompletely liberated from slag. LABORATORY TESTING A number of precious metal slag samples have been tested in the laboratory. It has been found that the composition of slags vary widely from one refinery operation to another. Typical laboratory procedure is to crush and size the slag followed by two-stage lab-scale electro¬static separation. The conductor fractions from the first and second pass are then combined as a prill concentrate; middlings fraction and clean slag tailings from the second pass are held separate. Selected results from laboratory testing are shown in Table 1. It is important to note in Table 1 that the assays represent overall silver and gold, not metallic values. From experience, the laboratory results are typically lower in grade and recovery than the industrial installations. This is largely due to the hydroscopic nature of precious metal slags coupled with the high local humidity in Carpco's Jacksonville, Florida, location. Relative humidity in Jacksonville can range form 60-95%, while most mining locations in the

Jan 1, 1987

By James M. Barker, Richard O. Y. Breese

The recognition of perlite as a distinct volcanic glass, and of the arcuate fractures that often characterize these natural glasses, dates back at least to the late 19th century and perhaps into the 18th century (Howell, 1974). Some authors suggest that recognition of perlite dates considerably further into antiquity, perhaps to the 3rd century BC (Caley and Richards, 1956). Most recent definitions identify perlite by the presence of vit¬reous, pearly luster and by the presence of characteristic concentric or arcuate perlitic fractures. As discussed more fully below, arcuate fractures in perlite need not be megascopic. The textures of perlite which commonly occur in deposits range from dense to highly vesiculated to pumice-like. Thus, the classical definition of perlite is descriptively restrictive compared to the natural range of textures and neglects to address the issues of genetic origin and geologic occurrence. In spite of the petrological usage of the term perlite, the in¬troduction of expanded perlite aggregate into industrial markets caused the term to be applied as well to the lightweight, cellular aggregate that is produced through the rapid thermal expansion of milled perlite ore. Perlite is distinguished petrologically from other natural glasses by a silicic or rhyolitic composition by the presence of 2 to 5 wt% total chemical water held within the glass structure and often by the presence of pearly luster and onionskin-like perlitic fractures. Worldwide, occurrences are associated with Tertiary through early to middle Quaternary volcanics and with the glassy portions of silicic domes and flows, with vitric tephra, with the glassy chill margins of dikes and sills, and with the glassy portions of welded ash flow tuffs. Less commonly, perlite is reported in association with the glassy portions of volcanic plugs and lacco¬liths. The perlite carapaces that partially or fully comprise extru¬sive domes and flows are often thick and areally extensive, and thus provide commercially attractive targets for low cost open pit mining. Perlite tuffs and tephras are also important commercial sources. Mining costs are minimized by open pit quarrying and by either bulldozer ripping or blasting. Crushing and sizing facilities are generally close to the pits. Due to the low weight and large volume of expanded perlite, unexpanded perlite that has been crushed and sized is usually shipped nationwide directly to local markets where it is expanded and processed for distribution to end users. The vast majority of international trade is in unexpanded perlite. In the United States, the evolution of the expanded perlite industry can be traced to the late 1930s and early 1940s when expanded perlite aggregate was introduced into gypsum plaster and concrete markets as a substitute for vermiculite (Schundler, Jr., oral communication, 1991) and was promoted for use in a much broader spectrum of products (Howell, 1974). By the early 1950s, end user markets were well established and small expanders were numerous east of the Mississippi River. Today, mining and expan¬sion is undertaken throughout the world. Worldwide production of perlite is dominated by the former Soviet republics, the United States, and Greece. Total worldwide production of processed perlite during 1990 is estimated at roughly 1 778 000 tons (Davis, 1991). The physical properties of expanded perlite that give this prod¬uct unique commercial value include: low bulk density, chemical inertness, high insulating ability, nonflammability and fire resis¬tance, and its ability to retain water. These and other physical properties give expanded perlite value in such uses as: lightweight acoustical and thermal insulating aggregate; component in plaster, insulating cements and lightweight concrete; loose fill insulation; as aggregate in horticultural applications; and in milled form as filter aid. In the United States and Europe, the largest end use is within building construction products.

Jan 1, 1994

By Gary Beerkirkcher

INTRODUCTION The mine hoisting equipment that is selected and installed at a mine is for the lifetime of the mine and, therefore, it is extremely important that a proper choice be made. Several types of mine hoist designs, including double drum hoists, single drum hoists, friction hoists, and various special designs of these, are available to¬day. Each has advantages and disadvantages and they should only be used in suitable applications. Drive systems for mine hoists must be evaluated together with the mechanical systems for each hoist application. It does not make sense to properly select the hoist mechanical system and then choose a drive system that turns out to be a bottleneck. MINE HOIST TYPES There are two basic types of mine hoists available in the world today: the drum hoist, on which the hoist rope is actually stored during the hoisting cycle, and the friction (Koepe) hoist, which merely passes the rope over the wheel during the hoisting process. Drum Hoists Single Drum Hoist: A single drum hoist (shown diagrammatically in Fig. 1) can be used as a service or production hoist with the cage or skip in balance with a counterweight (Figs. 2 and 3). It can efficiently service one or more levels, since the location of the counterweight at any time is not important. As a pro¬duction hoist with skips in balance, the single drum hoist is best used for single level hoisting. All rope adjustments for proper spotting must be done manually for the single drum hoist. This has to be done periodi¬cally to compensate for rope stretch. A variation of the single drum hoist is the divided drum hoist. If multilayer winding is necessary, the single drum hoist must have a divider to allow a sep¬arate compartment for each rope. If a counterweight is used with a divided drum application, the counter¬ weight rope can be wound on a smaller diameter (Fig. 4). Consequently, it moves a lesser distance than the main conveyance and rope adjustment problems are re¬duced. Double Drum Hoist: A double drum hoist in oper¬ation is shown in Fig. 5. Double Drum Hoist With One Drum Clutched: As a service hoist with cage and counterweight, a double drum hoist with one drum clutched (Fig. 6) can serve several levels efficiently. The advantage of having the clutched drum is that it allows rapid adjustment of ropes due to initial stretch. Some users may prefer the single drum hoist with manual rope adjustment, while others consider the added expense of the second drum and clutch is justified to make rope adjustments quickly. The value of this feature for a particular application should be discussed with the hoist manufacturer. As a production hoist with skips in balance for a multi¬level operation, the clutch can be adjusted for efficient hoisting from any level. Double Drum Hoists with Both Drums Clutched: There are also double drum hoists that have both drums clutched (Fig. 7). The main advantage claimed for this type of hoist is that if something happens in one of the two compartments, the hoist can operate in the other compartment to raise and lower men and sup¬plies. This feature is particularly favorable if there is only one shaft entrance to the mine. Multiple Drum Hoist: The multiple drum hoist, a South African innovation, is a type of drum hoist about which everyone should be aware. Basically, it uses two hoist ropes per conveyance which at great depths per¬mits use of smaller ropes and smaller drums. There are none in use in the United States today. However, they should be considered in very deep mines. Friction (Koepe) Hoists This hoist (Fig. 8) was introduced to the mining world by Frederick Koepe in 1877. The basic prin¬ciple of the friction hoist is that the hoist rope passing

Jan 1, 1982

By P. Stanley Jacobsen, Hemant B. Gala, Richard E. Hucko

Coal is the nation's most abundant fossil energy resource. However, its use is limited because of the nature and amount of associated impurities, such as mineral matter, sulfur, trace elements, and moisture. The moderate-to high- sulfur content of many U.S. coals is of particular concern as current and newer, potentially more stringent, regulations require reduction in sulfur oxide emissions. Release of particulates, NOx, and other pollutants during the coal-use cycle is also of concern. The current practice of controlling sulfur oxide emissions in the major steps of the coal- use cycle is limited mainly to coal preparation by physical beneficiation and, to a smaller extent, to post-utilization cleanup (flue gas desulfurization). Approximately two-thirds of the coal from states east of the Mississippi river is physically cleaned, whereas only 15% to 20% of the operating boilers in the utility industry are equipped with scrubbers to desulfurize the flue gas (Henzel, 1988). Physical beneficiation methods can be very effective in removing mineral matter, including pyritic sulfur, from coal. These processes take advantage of the differences in physical proper- ties between coal and mineral matter to achieve the separation. In a typical physical cleaning operation, raw coal is crushed to about a 50-mm (2-inch) top size. It is then screened into coarse, intermediate, and fine sizes. Crushing the coal liberates some of the ash-forming minerals, including a portion of the sulfur that is not chemically bound to the coal. The per- centage of impurities freed from the coal matrix generally increases as the coal particle size decreases. The coarse- and intermediate-sized coals are typically subjected to a cleaning pro- cess based on differences in specific gravity between coal particles and the associated impurities. Such cleaning processes used include jigs, dense-medium baths, cyclone systems, and concentrating tables. The fine-sized (typically less than 0.5-mm) coal, if beneficiated, is typically cleaned using a froth flotation technique based on the differences in surface proper- ties between coal and its associated impurities. At fine sizes, greater cleaning potential exists because a greater portion of the impurities is released from the coal matrix. However, this potential is not fully realized owing to inefficiencies in separating fine particles using state- of-the-art flotation technology. Overall, current commercial coal preparation processes are estimated to remove up to 60% of the ash and 10% to 30% of the sulfur, depending on the coal characteristics, with product weight recoveries rang- ing from about 60% to 90% and thermal recoveries of about 85% to 98% (Versar, 1985; Science Applications International Corporation, 1987; Hoffman- Muntner Corporation, 1978). For many U.S. coals, grinding to sizes finer than those used in current coal preparation plants is required to liberate most of the physically bound impurities from the coal matrix. As noted, conventional coal-cleaning methods are not very effective at these fine sizes. Accordingly, the expanded use of domestic coal requires the development of advanced coal cleaning technologies that can provide economic performance while reducing environmental impacts. Program Goals and Objectives The overall goal of the U.S. Department of Energy (D0E)'s Coal Preparation Program is to provide the scientific and engineering knowledge base that industry can utilize to produce economically competitive and environmentally acceptable clean-coal products for introduction as market conditions warrant. The coal preparation process(es) should remove undesirable components from coal and improve its handling characteristics, thus making the product coal technically, as well as economically, viable as a competitor with petroleum-derived fuels and natural gas, and with alternative environmental control technologies. The coal preparation process must be applicable to a wide variety of bituminous and low-rank coals. The cleaning process should reduce the variability among clean-coal products from various raw coals, thus significantly increasing the range of potential end uses for each raw coal. To the extent that deeply cleaned coal products have greatly reduced levels of

Jan 1, 1988

By J. B. Gwiazda

In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the .u factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the .u of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves.

Jan 1, 1987

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

GENERAL STATEMENT The primary objectives of any field data gath¬ering effort should be to (1) identify and gather the data necessary for the project and (2) obtain the data in a state-of-the-art manner. All too often the initial field data are collected both areally and tem¬porally in an illogical manner without the guidance of a conceptual model of the ground water flow systems involved or even a review of existing geo¬logic literature on the area of interest. The initial data collected frequently are of limited value while necessary basic reconnaissance information is miss¬ing. Initial field data should be collected with the intent of developing a hydrologic overview of the potential mine site and surrounding area. Ob¬viously, one of the initial objectives is to define the area requiring a hydrologic investigation. The data requirements should be identified by the time frame in which collection should be made and by the corresponding increase in sophistication of the data requirements with development and operation of the mine. The data requirements are summarized in Table 1. INITIAL LEVEL SITE INVESTIGATION Area Determination The initial task of any hydrogeologic investi¬gation is to determine the boundaries of the area requiring study. Obviously, the site of the proposed mine is included in the study area. The areal extent beyond the site may be determined from an eval¬uation of existing geologic and topographic maps. Those formations that overlie the ore body, the formations containing the ore body, and the formation(s) that lies immediately beneath the ore body are of direct concern for proper site recon¬naissance. Additional formations below the ore body may require study depending upon their thick¬ness, hydraulic conductivity, and degree of inter¬connection with the mine workings. This initial viewpoint identifies hydrostratigraphic units based strictly on geologic concepts such as mineralogy and structure. Formation outcrops, synclines, an¬ticlines, faults, and fracture and joint patterns are used to delineate the area of the site reconnaissance. The simplistic hydrogeologic environment (il¬lustrated in Fig. 3, chapter 2) requires that field data be collected via test wells and/or geophysical techniques. This approach is necessitated by the lack of surface features such as formation outcrops, streams, and springs. Fig. 5 (chapter 2) illustrates a slightly more complex hydrogeologic regime. The potential mine sites at locations A, B, C, D, and E each intercept a different ground water flow sys¬tem or combination of flow systems. Therefore, each mine location requires that a different area and size of area be investigated. A more complex geologic setting as illustrated in Figs. 6 and 7 (chapter 2) may be approached differently. The area included for the site recon¬naissance should encompass sufficient surrounding area to include the outcrops of those formations suspected of being influenced by the future mine. Even adjacent areas not suspected of being influ¬enced may be investigated if the formations of in¬terest crop out in those areas. Such an extension of the area of investigation would provide a greater regional understanding of the hydrogeologic properties of the formations (hydrostratigraphic units) of interest. Geologic Investigation The initial step before conducting the site re¬connaissance is to review all existing literature on the geology of the area. Existing information should be augmented with new exploration data on the dip, strike, thickness, and lateral extent of the for¬mations in the area. Exploration hole logs should be reviewed for indications of lost circulation, rub¬ble zones, and water producing zones. Existing aer¬ial photos such as those available from the US Department of the Interior, EROS Data Center,

Jan 1, 1986

By James R. Arnold

Gold Fields Mining Corporation started processing ore from it's Chimney Creek operation in northern Nevada in November, 1987. Since that time production has averaged about 200,000 ounces per year from the milling and dump leaching circuits combined. The events that led up to commissioning of the plant started in January of 1985 when the Gold Fields exploration team drilled their first hole in an area now known as Discovery Hill. The major orebody was located later in the year in what is currently the major focus of activity, the North Pit. Subsequent drilling since plant startup has found another 14 million tons of ore two miles south of the original orebody. This South Pit orebody is currently being stripped and metallurgical testing is underway. Engineering Engineering on the Chimney Creek plant started in earnest in May, 1986. Davy McKee Corporation from San Ramon, California, did both the engineering and the construction management. Before engineering began the entire senior staff of the Plant Department was hired and assigned to the project. Senior staff included the Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist. Gold Fields does not employ an internal central engineering staff to design their plants. We feel that the people who design the plant should be the ones who will be required to run it once it has been commissioned. This philosophy affords a continuous accountability from start of design through commissioning and operations. All four of the senior staff spent the most critical phase of the engineering stage at Davy McKee, so the plant is truly a reflection of all four disciplines. In bringing on the Maintenance General Foreman from the start, there was a focus placed on ease of maintaining the plant. Another philosophy that was adopted from the start was that the plant be easy to work in and maintain for both operators and maintenance personnel. Numerous design criteria were given to Davy McKee to ensure this. For example, we required that every piece of equipment could be reached easily with a crane or fork lift. All conveyors were required to have walkways on both sides for ease of maintenance and a minimum of 30 inches of clearance under them for ease of cleanup. Visibility for operators was considered very important and the control rooms and general plant layout were designed with this in mind. A list of about 100 general rules were given to Davy McKee during the early layout meetings and these were held as the general design criteria throughout engineering and construction. While engineering was progressing in San Ramon, California, dump leach test work was being done at the site. Three 6 foot diameter fiberglass columns were constructed and filled to a 20 foot height with low grade ore. Leaching characteristics were determined and the dump leach design was based on information collected from these columns. Construction The major contractors on the project were The Industrial Company for the bulk of the piping and structural, EMKO on the electrical/ instrumentation and Hunewill for most of the earth work and leach pad construction. All plant operating and maintenance personnel were employed eight weeks before startup for intensive training. Instructors were Gold Fields personnel, including foremen, and Davy McKee engineers. On occasion, selected suppliers representatives were brought in to discuss major pieces of equipment. Project completion was targeted for January, 1988 at a projected cost of 78.9 million dollars. Actual startup was about 1 1/2 months ahead of schedule and slightly under budget. Full design capacity was achieved the first full month of operation. Crushing Run of mine ore is delivered to the crusher and dumped into a 90 ton hopper which is topped with a 30" X 30" opening stationary grizzly. Ore is fed to a 50" X 60" Telsmith jaw crusher with a 5' wide apron feeder. The closed size setting of the

Jan 1, 1989

By James N. Grassby

Metal production has traditionally been a cyclical industry. 1977 in the base metals industry has highlighted that fact with a steady decline in the market place for base metals. Corporate aggregate planning is based on good forecasting and when changing conditions make that forecasting inaccurate, alternate analysis of schedules must be accomplished quickly. The computer is the only medium powerful enough to answer management's questions within an acceptable time frame. Worldwide communications networks now make it possible for the many multinational mining companies to arrive at coordinated corporate production plans based on similar basic data from several producing divisions. The Ontario Div. of Inco Metals Co., a unit of Inco Ltd., encompassed during 1977 14 producing mines, 2 nonproducing mines, 1 nonproducing open pit, 5 flotation mills, I smelter, 2 nickel refineries, 1 copper refinery, and an iron ore recovery plant. Several new mines are under development. Mine production approximates 63 500 tpd (70,000 stpd) from some 1000 workplaces. The need for longrange mine planning and scheduling is obvious and until recently has been satisfied by manual methods. However, while manual methods are often quite practical for a small number of mines, or for mines with relatively short lives, or for mines with only one mining method, the situation in the Ontario Div. is too complex and long-term to allow effective, timely planning by such methods. In 1967, a year after the first computer was installed in the division, work was begun on the computerization of the more than 30,000 borehole logs kept in the mines exploration files with a view to speeding up the assessment of mineral resources. Also in 1967 one of the first substantial systems developed on the main business systems computer was the monthly scheduling of development and production for all mines. This system encompassed tons productions per workplace, feet development per workplace, auxiliary and service activities and the requisite labor for each of these. These two dates are mentioned to show that concurrently with the acquisition of computers in the division there was an early understanding of the importance of automating mine scheduling and planning and of the potential for effective and useful computerization of such activities. In 1970, a pilot program was developed in one of the mines containing a large number of workplaces (350) to schedule the production on a network basis. Varying degrees of interest were shown by the individual mine engineers based on the complexity of the production planning problem at each mine. In an evolutionary process, mines were added to the system as its value became obvious in terms of time saving, flexibility, and accuracy. In 1975, the central mine engineering department decided that the system should be extended to all mines so that a complete division-wide long-range scheduling system would be available upon which to base development and production planning. This system would use basic ore data, allow the addition of scheduling commands, and calculate a schedule for the life of the mine. Many reports could be produced for each mine and all-mine summaries would also be available. The significance of the system can be stated in one phrase-speed, accuracy, and flexibility. Whereas in the past a long-range all mines schedule might have been done once a year, taking from two to three months, now it can be done in a day. The chief mines engineer can incorporate changes in input data and external constraints and call for a "new look" at the long-range schedule to assist him and mines production management in their decision-making processes. Using previous manual methods this could not be done quickly enough to be useful.

Jan 1, 1979

By Sidney A. Hiscock

INTRODUCTION Lead and zinc are today rightly regarded as sister metals. Historically, however, they differ markedly with lead being known and in general use as the metal since 3000 BC in several countries which at that time might have been described as mature economies. Zinc was not isolated and recognised as an industrial metal with distinct and valuable properties until about 1700-1800 AD (although it had, of course, been utilised as a constituent of brass for some 2000 years previously). c were well established as industrial metals by the beginning of the 20th century. The present paper briefly reviews the overall growth in world consumption* and changes in the three main areas which today represent the mature economies or industrialised areas, i. e. ,Europe, the United States and Japan. Trends in overall consumption are determined by the demands for the individual uses for lead and zinc. The changes in patterns of use in the industrialised areas - individually and collectively - and reasons for the changes are also considered. LONG TERM TRENDS - 1900-1984 The World Picture Since the beginning of the 20th century the world consumption of lead (3.9 million tonnes in 1984) has grown almost fourfold and that of zinc (4.7 million tonnes in 1984) is almost eleven times greater than it was in 1900. The long-term trend has been one of continuing expansion for both. metals although the pattern has been severely disturbed at certain periods, for example in the years immediately following the two world wars and at times of marked recessions in industrialised areas. Apart from such major setbacks, new levels of consumption have been established regularly every few years. Recently, however, consumption has not really grown and with only a slow and partial recovery in world industrial activity at present, there is no indication that the peak consumptions of 4.2 million tonnes of lead, set in 1979, and 4.8 million tonnes of zinc (1973) will be exceeded in the near future. The growth in world lead and zinc consumption since 1900 is shown in Figure 1 (which includes changes in copper and aluminium for comparison). Overall growth over the period 1900-1984 has been about two percent a year for lead and about three percent for zinc (annual growth for copper has also been about three percent). Increases in tonnage consumption and growth rates by decade for lead and zinc are summarised in Table I. However, the growth rates by decade sometimes conceal very large individual annual increases and decreases. For example, in 1975 lead consumption fell by 13 percent and zinc consumption by 22 percent compared to 1974. In 1976 consumption of lead and zinc rose by 12 percent and 18 percent respectively. The years when consumption 'landmarks' (ie one, two, three and four million tonnes) were first reached are shown on Table 2. Clearly, the pattern of shorter periods being required to attain each extra million tonnes of consumption was broken in the 1970's. The consumption of zinc overtook that of lead for the first time in 1940, and since 1946 has always been at a higher level. Trends in industrialised areas The two key industrialised areas at the beginning of the 20th century were Europe and the United States which between them accounted for some 95 percent of world lead consumption and 98 percent of zinc consumption. Figures 2 and 3 show the growth in lead and zinc consumption, and Figures 4 and 5 the percentage shares for each area since 1900. Europe has usually been the major user of lead except for some years in the 1920fs, most of the 40's and early 50's when consumption in the United States was larger. In 1900, Europe accounted for about 64 percent of world lead consumption and the United States just over 30 percent. Currently, Europe takes about 40 percent of world consumption and the

Jan 1, 1986

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By R. Kent Comann

Introduction This article describes ways to maximize employment potential for those looking for a job in the minerals industry. Advice on the right way to prepare a resume is followed by tips on where and how to look for a job and ways to make the best of an interview. Information on current salaries is also included. RESUME NAME: Give home and college addresses and telephone numbers, ADDRESS: if still in school. Otherwise, give home address and TELEPHONE: telephone plus business telephone number. JOB OBJECTIVE: What do you want to do? EDUCATION: List degree(s), name of college(s), location(s), and date(s) degree(s) received or to be awarded. Do not list your high school but do include postgraduate education, training, or seminars. WORK EXPERIENCE: List your most recent experience first, and the rest in reverse chronological order. Include dates (month and year) and company names. First show the entire time spent with each employer, then list the various jobs (with dates and duties) you may have had with this employer. MILITARY SERVICE: List dates, branch of service, and duties. If you did not serve, delete this section. COLLEGE ACTIVITIES: Show organizations, sports, and offices held. Only list these if you are a graduating senior, otherwise delete this section. PERSONAL DATA: List age, marital status, height, weight (optional), foreign languages spoken, professional memberships, and health (if good or excellent, otherwise do not mention it in the resume). If you prefer a certain location and would not accept a job anywhere else, list the location desired. Otherwise, do not restrict yourself. Also list special skills, hobbies, or interests that will help sell you to the company. Be brief in this section and use good judgement as to what you say. REFERENCES: Furnished on request. Resume Guidelines There are some things to remember when preparing a resume. These include: • Limit your resume to one page, unless you have a lot of experience. Then it should be a maximum of two pages. • Use 8.5 x 11 plain bond paper, white or off-white. Stay away from colored paper, odd size paper, and cute resumes. They are counter-productive. • Make certain your resume is typed professionally. Be sure to check for spelling, grammar, and typographical errors. • Have your resume reproduced by copier or printing shop. Do not use carbon copies. • Keep your sentences and remarks short and punchy. Use action verbs. • Have a good layout that is easy to read, brief, and uncluttered. Allow some white space in your resume. There are also some things you should not do when preparing a resume. These include the following: • Do not include a photograph or mention your religion, race, or sex. • Do not leave gaps of time unaccounted for in your resume. • Do not include your college transcripts with your resume or list all the courses you took in college. • Do not list all the articles you have written. Only list one or two, if appropriate for, say, a research and development job. • Do not include reference letters with your resume. Any good, prospective employer will make telephone reference checks. • Do not use "Curriculum Vitae," use "Resume." • Do not include unnecessary data such as your social security number, passport number, or PE registration number. If you are a PE or EIT, be sure to mention that in your resume. Do not mention US citizenship or US work permit unless you have a foreign education or are not a native US citizen. • Do not include present salary or salary requirements. This will come out in the interview. • Do not use the word "I" in your resume or refer to yourself in the third person. • Do not lie, distort, exaggerate, or editorialize in your resume. • Do not have your resume prepared by an outside resume service. This is unnecessary, costly, and generally not effective.

Jan 2, 1985

By Rajendra K. Devala, Sukumar Bandopadhyay

In order to be able to design an adequate mine ventilation system it is necessary to understand how pollutant masses disperse into turbulent ventilation air streams. The mixing of gases and dusts with ventilating air, and dispersion of these pollutants in mine airways, are important flow or transfer problems. For the solution of such problems, the finite difference technique is often used. In this paper, numerical approximation schemes for the solution of the mass transfer problem have been developed using several explicit finite difference schemes. Results from these schemes have been compared with the analytical solution for spatial and time increments using the concept of Root Mean Square error. The comparison of various schemes reveals that the explicit finite difference schemes analyzed here give more accurate results at lower velocities, though the results are dependent on the spatial and time steps considered. At higher velocities the schemes tends to oscillate about the analytical solution. INTRODUCTION In mining, most of the mine air pollutants, such as respirable dust, methane, blasting fumes and diesel exhaust, are produced at the mine face and in haulageways. Injection of any of these pollutants in the mine ventilating air causes disturbances of the equilibrium state of the pollutant concentration of the ventilating air. The mixing of these gases and dusts with the ventilating air are typical examples of mass transfer processes. Pollutants are diluted by the stream of ventilating air flow along airways and discharged into the outside atmosphere of the mine through a return shaft. The dispersions of these pollutants in mine airways are important flow or transfer problems. Knowledge about pollutant sources, emission and transport of these pollutants are essential for ventilation planning and design to ensure good engineering control over air quality, quantity, and environmental conditions. Modeling of mass transfer phenomena in mine ventilation is largely based on extensive studies of turbulent dispersion in wind tunnels and the lower atmosphere of the earth. The general approach in the formulation of models lies in the recognition that these mass transfer phenomena can be conceptualized as a flow problem that can be solved by the general application of Taylor's theory (1954) of turbulent flow and diffusion with modification appropriate static and dynamic parameters. If the face zone is initially contaminated by a pollutant over a long section (e.g. instantaneous exposure of large gas accumulation, rapid methane release etc.), the process of decontaminating the mine workings can be regarded as displacement and diffusion of the tailing edge of the cloud of contaminates in the form of a semi-infinite gas body, for which a one-dimensional mass transfer equation is represented by: [at + u • ax = Ex ex2 + f (x,t)(1.1)] In Eq. (1.1), the term [~~] is the rate of growth of the concentration in the differential element, while a [(5xx )] is the net gain of material due to convective transfer. The two terms balance the total loss of material due to turbulent dispersion that is represented by [Ex . 21] and f(x,t), the source term for the pollutant in the roadway. Several cases of mass transfer that occur in mining can be identified and modelled using the conservative volume element method. The governing equation, along with the associated boundary and initial conditions, source terms and their solutions are detailed elsewhere (Bandopadhyay and Ramani, 1988, Devala, 1989). Only one mass transfer situation is presented (Figure 1) here for the present analysis. Mixed Evolution of Pollutant (multiple pulses) Typical examples of these situations are emission of pollutants from multiple diesel vehicles in the face area, and gas emission from standing pillars as well as from broken coal seam during coal cutting with a continuous miner. The mathematical statement of the problem along with the initial and boundary conditions are given below, Governing Equation: [do+u- =E d2c(1.2) ataxx dx2] Initial conditions:

Jan 1, 1993

By Sue Sawyer

In October, November and December, teachers at the National Science Teachers' Association (NSTA) area conventions were lured to visit the Minerals Education Coalition (MEC) booth by the promise of free rocks, video tapes, and literature on earth science topics. Teachers were content to stand in lines that at times took as much as an hour and a half to progress through the exhibit. Adding to their interest was the opportunity to converse with people from the minerals industry who were staffing the booth. The MEC was formed by a group of associations and companies in the minerals industry. They felt the need to educate the public about the current-day minerals industry and how it affects everyone's life. The American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc. (AIME) underwrote the major expenses of exhibit space and exhibitor services at all three conventions in 1997. SME members have been active in staffing and chairing the NSTA exhibit booth. But they represented only about one-third of the volunteers for the booth. Additional volunteers were drawn from local minerals companies, associations, interested teachers and other support groups. (See donors list on page 64.) In 1997, the conventions were held in Pittsburgh, PA, Denver, CO and Nashville, TN. The local SME GEM Committees chaired Pittsburgh and Denver and The Rogers Group Inc. supplied the chairperson in Nashville. Each exhibit relies on a local committee to plan and run the booth. Support is available from the GEM Projects Coordinator at SME headquarters. Usually the chair of the committee is drawn from the local SME membership. It is easy to see why the MEC booth is so popular. Free literature, videos, posters, workshop packets and rock specimens were handed out to about 5,500 science teachers at the three sites. Feedback from the attending teachers was overwhelmingly positive: "Who do we thank for these great materials?" Some of those great materials included Caterpillar's outstanding donation of Common Ground videos and teacher packets for Pittsburgh, Denver and Nashville for every teacher expected to attend. Placer Dome sent enough What's a Mine coloring books and It Comes from a Mine activity books for teachers at all three sites. Newmont supplied all teachers with an excellent poster describing gold processing. Women in Mining Education Foundation had plenty of its classroom activity handouts. The American Coal Foundation provided a list of its educational materials that are given out free to teachers who request them on school letterhead. The National Mining Association had its Facts About Minerals and Facts About Coal available. US Borax provided the funding to reproduce the NMA black and white posters that were published as inserts in Mining Voice magazine. In addition, each site had other materials chosen by the local committees to round out their selections. At each exhibit site, committees raise between $8,000 and $12,000 in cash donations from industry to support the high overhead of producing this exhibit. Usually, there are between 12 and 18 different rock specimens representing a range of mineral ores from aggregates to the metal ores. The committee arranged to have these donated (along with the freight charges to ship them to the site). Literature is also donated by the industry. It covers a diverse spectrum from basic information on rocks and minerals to posters, flow charts, videos and coloring books. Other donations include storage space before the event, haulage of items to the hall and door prizes. These exhibits could not have been staged without the network of SME Local Sections and the able chairs and committees at each site. At this time, credit should be given to their extensive donation of time and effort.

Jan 1, 1998

By W. C. Wheeler, G. Suranyi, J. F. Gentleman, J. Muller, R. Kusiak

INTRODUCTION In 1974 two of the present authors reported the results of a pilot study indicating an increase of lung cancer risk in Ontario uranium miners. (Muller, Wheeler, 1973, 1974) The study was based on data contained in a computerized Mining Master File maintained by the Ontario Workmen's Compensation Board that contained information on miners examined in Ontario who had either 60 months of dust exposure in mines or had signs of pneumoconiosis or tuberculosis. Including the above conditions the definition of uranium miners added the condition of one month or more of uranium mining experience in Ontario. This list of Ontario uranium miners contained 8,649 names. Following the results of this first pilot study, we embarked on creating a file of uranium miners containing information on men with one month or more of uranium mining experience in Ontario without any further conditions. This file was used by the Royal Commission on the Health and Safety of Workers in Mines in their study of risk in Ontario uranium miners. (Hewitt 1976) This file contained 15,094 names. In this report we give an outline and progress report on a study of Ontario miners that we are conducting at present. It was felt that the male population of Ontario is not necessarily an adequate control population for uranium miners. A preliminary examination of the work history of uranium miners indicated that the majority of them (about 90 percent) had other mining experience in addition to their exposure in uranium mines. We therefore considered it useful to evaluate the possible effects of non-uranium mining on risk, and for this reason decided to make the Uranium Miners Study part of a study dealing with the mortality of Ontario miners in general. Aims of the Study The aims of the Study include the evaluation of: 1) the risk of dying by cause in non-uranium miners as compared to the male population of Ontario and Northern Ontario. 2) any differences that might exist in the death experience of non-uranium miners by cause according to ore mined. 3) the effect of length of exposure in non-uranium mines on age-specific risk by cause. 4) the dose-response function for primary cancer of the trachea, bronchus and lung from exposure to radon and its short-lived daughters. 5) the possible effect of the mining environment on deaths from causes other than cancer of the trachea, bronchus and lung. The study will address itself to a number of other factors that might well affect the dose-response function. These include: a) factors in the mine environment - other than radon daughters - that might affect lung cancer mortality. b) the effect of non-uranium mining on lung cancer risk in uranium miners. c) the effect of age as well as age at time of exposure on lung cancer risk. d) questions of latency and the possible dependence of latency on age at time of exposure. e) smoking as an important factor in lung cancer risk. f) Histological type of cancer in relation to the various parameters of exposure and age. MATERIALS AND METHODS The Study is making use of existing computerized data files and has set up certain new files. These include the Mining Master File and the Model Development File. The Mining Master File This file is a computerized record of data on individual miners obtained at yearly miners' examinations that have been carried out since the mid 1920's. The conditions for inclusion in the Mining Master File have been indicated above. Information contained in the file includes: (1) Identifying information: a) Surname and given names b) Date and place of birth c) Miners Certificate Number d) Social Insurance Number if available. (2) Updated Employment data obtained at each miner's examination: a) Year of first dust exposure in Ontario b) Year of first dust exposure outside Ontario c) Number of months worked in mining d) Ores mined e) Mining areas and mines f) Occupations

Jan 1, 1981