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By J. Guzman

Introduction The Carlin Trend is North America's premier gold producing district. It is located in northeastern Nevada's Elko and Eureka Counties along a northwest trending belt about 65 km (40 miles) long and 8 km (5 miles) wide (Thorstad, 1989; Jones, 1989). This trend is the worldwide reference site for epithermal, sedimentary rock-hosted microscopic gold deposits. At least 19 deposits have been discovered to date, varying in size from 933 kg to 1.08 kt (30,000 to 35 million contained oz) of gold (Fig. 1). Newmont Gold Co. and its parent, Newmont Mining Corp., jointly constitute the largest mineral right holders in the district. They own or control more than 1000 km' (386 sq miles) in and around the Carlin Trend and own all or part of I6out of the 19 mines and prospects identified to date. Since the initiation of Newmont's exploration activities in the Carlin Trend in 1961, 2.24 kt (72 million oz) of cumulative gold resources have been identified. Cumulative production from all mines since the start-up of Newmont's Carlin mine in 1965 to the end of 1989 was about 202 t (6.5 million oz) (Jones, 1989). The incentive of sustained high gold prices and innovation in processing technology resulted in a significant acceleration of gold output over the last few years. Newmont Gold alone produced more than 43.5 t (1.4 million oz) in 1989. That is equal to 22% of the cumulative 1965 to 1988 output, and an almost 200% increase over its 1986 output. The same incentives produced even more spectacular exploration results. In each of the last five years, net additions to reserves and resources outpaced current production by substantial margins. These facts demonstrate the spectacular past prospectiveness of the Carlin Trend and the success of focused, multi-disciplinary exploration methods that made it possible to more than offset the recent accelerated depletion of gold resources. However, is this situation sustainable? How long can the mining companies along the Carlin Trend keep on finding resources faster than they deplete them? These are some of the questions that motivated this study. The authors have not quantified the future potential for gold exploration in the Carlin Trend nor established a deposit discovery path. But strong indications were discovered that the [ ] Carlin Trend remains a relatively immature exploration district and that the potential for significant new discoveries is high. Methodology and data The approach chosen to address the above questions was simple. The authors compiled deposit size data, measured in contained ounces of gold resources, for all known deposits along the Carlin Trend (Table 1). The resource information was obtained by adding cumulative historical production (adjusted for mining losses and metallurgical recovery) to 1989 year-end published resource inventories. In a mature exploration area, where most deposits have been discovered, this distribution would be expected to approximate lognormality and would plot along a straight line on a lognormal probability scale. This result was found in previous work by Allais (1957) and recently confirmed by Cox and Singer (USGS, 1986) in regard to various types of mineral deposits in several regions of the world. It was also found to hold true for oil and gas pool size distributions (Arps and Roberts, 1958; Kaufman, 1962; McCrossan, 1969). [ ] The data used were compiled by Newmont Exploration geologists. The purpose of the study is to make inferences about the underlying geologic processes in the district and the maturity of the exploration effort. Therefore, deposits were not classified according to ownership but according to geologic occurrence as known from current information. Newmont's Post and Barrick's Goldstrike and Betze deposits, for example, are shown as a single occurrence to reflect the actual geologic setting. The cumulative frequency distribution of deposit sizes was plotted on lognormal probability paper (Fig. 2). The abscissa shows the cumulative fraction of deposits at or below a certain deposit size and the ordinate shows the deposit size in thousands of ounces of contained gold resources.

Jan 1, 1992

By Don Suttie, Frank Roybal, John Wright

The Roybal Raise incorporates the simplicity and low cost of a short bald-headed raise while overcoming the usual impediments to driving a high baldheaded raise. The concept provides for speedy excavation and maximizing the use of the excavated opening, using the rock for structural support. The design is simple: two baldheaded raises are driven parallel to one another, interconnected by dogholes about 8 m (25 ft) apart. In this fashion, the raise can be driven safely and efficiently to 60 m (200 ft) or more. When the raise is complete, the excavation requires no timber for support or partitioning. Rather, the rock itself supports the construction. As a test, a two-compartment Roybal Raise was attempted at the Independence mine in the Willow Creek District of Alaska. One side was driven 2 x 2 m (6 x 6 ft), intended for a manway and equipment skip. The other was driven 1.5 x 1.5 m (5 x 5 ft) for a muck compartment. Both sides were driven at a 65° angle from horizontal and 1 x 1 m (4 x 4 ft) dogholes spanned the 4.6-m (15-ft) pillar between the two compartments. After topping out at 50 m (160 ft), the raise was found to meet all expectations of efficiency, low cost, and practical simplicity. Technique Preparing the Roybal Raise for driving involves no elaborate or costly site prep except assembling materials and providing for muck removal. Normally, the raise muck falls directly onto the sill or track and is picked up either by an overshot mucker or a front-end loader. Raise mining proceeds initially by taking up both compartments simultaneously to the first doghole. It progresses by alternating one side as a clear manway for access with the other side, the active side, as a muck compartment (Fig. 1A). This stage proceeds quickly for no mucking or staging is required to cycle the first three rounds. On the fifth round, in addition to the normal raise round, rib holes are winged out toward the adjacent raise, thus starting the dogholes. Finally, as the sixth raise rounds are drilled and shot, the first doghole is also completed from both sides. The second stage commences by establishing one side as a clear manway for access while the other side, used as a muck compartment, is prepared for active mining. Figure 1B illustrates this step in the succession of raising with the active side advancing to the second doghole level. During this stage, services are maintained through the manway side and equipment is stowed in the doghole during blasting. As the active side advances, muck may be allowed to fill the bottom portion of that side since it is unnecessary to clear it for access. The next stage of raising involves cleaning out the former muck compartment and preparing it for use as a manway. At the same time, the former manway is prepared for advance and use as a muck compartment (Fig. 1C). As in the previous step, the same doghole is used for access and equipment storage. The active side advances until the second doghole is completed, at which point equipment and access are moved to the new doghole. From there, the active side is driven past the second doghole 8 m (25 ft) more to the third doghole level. Again, half of the doghole and the next raise round are mined. The fourth stage involves switch ing back the manway side for the muck side (Fig. ID). The new muck side is driven to the third doghole where equipment storage and access are reestablished. From there, the active side is advanced to the fourth doghole. In the same fashion, the raise is advanced to the top, alternating the manway side with the active side, maintaining access through the inactive manway, and storing equipment in the uppermost doghole. The cycle time for driving each compartment, regardless of the raise height, is greatly improved over that for normal bald-headed raises of comparable height. By storing equipment in the upper-

Jan 1, 1984

By R. H. Moore, B. D. Breitenstein

INTRODUCTION The United States Uranium Registry (USUR) tissue program was funded in 1978 by the United States Department of Energy (DOE) through its Human Health and Assessments Division, and registrant enrollment began in 1980. It is operated by Hanford Environmental Health Foundation (HERF), Richland, Washington, with support provided by Pacific Northwest Laboratory (PNL).** The USUR tissue program is the first systematic effort to contact and gain assistance of workers occupationally exposed to uranium and to seek and study uranium deposition in human tissue. This paper describes the program objectives, populations selected for study, procedures for enrolling workers in the program, types of tissue to be studied, and analytical procedures to be used. PROGRAM OBJECTIVES The objectives of the USUR tissue program are to: 1) Determine the distribution and concentration of uranium, if any, in the tissues of occupationally exposed workers. 2) Compare bioassay measurements of exposed individuals with the results of analyses of tissues obtained at autopsy. 3) Seek evidence of histopathologic changes related to any uranium deposition that may be present. 4) Conduct analyses of whole bodies, when available, to obtain more precise data on the uranium burdens, if any, in the body and organs, and especially the distribution in parts of the body, such as the skeleton, that are not usually accessible for sampling. 5) Develop data that will assist in evaluating a) the accuracy of current in-vivo measurement techniques, b) the propriety of existing regulations, and c) the adequacy of current protection programs. SELECTION OF POPULATIONS FOR STUDY Selection of the populations for the USUR tissue program was based on studies of the United States Uranium Registry described elsewhere (0c81). Potential populations for study were quite varied because of occupational exposure to different chemical forms of uranium, different levels and eras of exposure, varying ability to identify populations exposed in the past, and the general interest and cooperativeness of the populations. Nevertheless, certain identifiable populations emerged that were both willing to participate in the study and likely to provide useful information. Among the targeted populations are individuals who were exposed to uranium prior to the existence of modern exposure standards. These individuals are presently being enrolled by the USUR. The reliability of concentration and distribution data and the demonstration of effects are more likely in such individuals than in those who have received less exposure. However, the targeted groups also include individuals who Have been exposed to permissible levels of uranium, and reference employees in uranium plants who have not worked with uranium. The lesser exposures are of immediate interest because of the rarity of heavy exposures under current standards. The sampling of potentially exposed individuals with negative in-vivo measurement results may serve to check the adequacy of in-vivo measurements in the detection of deposition. [ENROLLMENT PROCEDURE] The enrollment of participants follows three patterns. 1) Presently employed workers are enrolled by the cooperative efforts of USUR staff with the medical, industrial relations, and/or health physics personnel of the uranium facility. Volunteers may be enrolled by medical personnel at the time of their physical examination or by health physics personnel at the time of a bioassay or lung counting procedure. Follow-up after initial contacts with currently employed uranium workers is carried out by on-site personnel, but subsequent follow-up is the responsibility of the Registry. 2) Retired workers are contacted by mail, with a covering letter from their former employer supporting the principles of the study but leaving participation in the program on an entirely voluntary basis. 3) An alternate method of enrollment is to obtain a USUR autopsy consent after death from the next of kin. As part of enrollment in the program, permission is requested for access to the individual's medical and exposure data, and a short occupational history is filled out. The autopsy permission agreement signed by the individual is for an initial period of 5 years; the USUR ordinarily seeks renewal of the agreement upon its expiration. While this agreement provides

Jan 1, 1981

By Dr. R. W. Mead

There are two basic problems when one attempts to separate a solution of metal ions utilizing a chelation-solvent extraction cascade (commonly called liquid ion exchange). First, it is impossible to produce any great degree of separation unless the separation factor (also called selectivity, which is the ratio of the distribution coefficients) is very large. Second, the richest possible extracted product, using an infinite number of equilibrium stages, is at best only in equilibrium with the feed solution. These factors combined with equipment design considerations have led to a situation where only the most selective chelating agents are applicable to economic industrial operations. The authors know that distillation techniques are used routinely to separate compounds with nearly identical relative volatilities and that there should be a way of operating solvent extraction cascades in an analogous manner to obtain separations of metal ions. The advantage of this approach would be that solutions of metal ions of any desired purity could be obtained even if the separation factor was close to unity. This paper presents this idea and suggests some possible applications.

Jan 1, 1977

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 Jack A. Simon

Coal is defined as a combustible rock that originated in the accumulation and physical and chemical alteration of vegetation. Coal can be ignited and burned like the wood that was man's earliest fuel, but coal gives more heat than the same amount of wood and has many industrial uses. Unlike wood, coal is not generally found on the surface of the ground but is mined from the solid rock of the earth. Before taking up the occurrence, formation, and the unique character of coal in the rock sequence, let us briefly consider the general character of all rocks in the crust of the earth. Relatively little is known about the very earliest history of the earth, although a number of logical theories have been developed. Manned exploration of the moon and the rocks returned to earth continue to provide some answers to questions concerning the earth's origin and history. The period of earth history with which we concern ourselves here begins after great areas of water had evolved on earth and substantial land areas of once molten (igneous) rock existed. Weathering and other processes that attacked the surface of the rocks reduced them to small particles. Water provided by rainfall transported this material, ultimately, to large bodies of water where it was deposited. Thus was initiated the geologic sequence of sedimentary rocks that lie above the once molten igneous rocks. The crust and upper mantle of the earth are made up of several large plates along which difEerentia1 movement (primarily horizontal) takes place. Such movements, although small, are occurring today. These horizontal movements are accompanied by vertical movements causing important changes in the relative position of the continents and oceans. Low-lying coastal swamps with favorable climates are ideal locations for coal accumulation and burial by river-borne sediments, and large areas are affected by relatively minor vertical movements of the shoreline. The large plates have continually or intermittently moved throughout the geologic past, resulting in various configurations of the continents and oceans, and coal-forming conditions have occurred at numerous times and places. Vast areas of our continents have been repeatedly inundated by shallow oceans and only relatively small parts of the present continents appear to have been above sea level throughout geologic time. Great forces acting principally along plate margins have folded and broken the crustal rocks, forming large belts of mountains.

Jan 1, 1981

By Juan H. Rojas

The Chuquicamata mining and metallurgical complex is the main division of CODELCO-CHILE. Its comparative advantage in the world of primary copper producers had deteriorated severely and, consequently, its top management launched an extensive modernization plan. This paper shows how Chuquicamata recovered its position as a low-cost copper producer and is now well positioned to face the end of this century, as far as competitiveness is concerned. It also shows the company's vision for the )(XI century. Given the volume of its ore reserves, one possibility is expanding the sulfide production line capacity, while in the case of the oxide production line, ore reserves are limited so that there is a need to maximize the effectiveness of both organization and technology use. A final consideration has to do with the balance of supply and demand. Most observers of the copper market agree that supply is growing faster than demand. On the other hand, copper is facing severe environmental problems, especially in the piping business, plus the usual threat of substitution. Nevertheless, copper producers are making very weak efforts to defend the market and increase copper consumption through the promotion of new uses of the metal. This issue is crucial and must be considered if copper producers want to maintain a profitable market.

Jan 1, 1998

By Duport. P. J.

INTRODUCTION In February 1981, the Canadian Institute for Radiation Safety (CAIRS) initiated a routine Personal Alpha Dosimetry service for personnel of the Canadian uranium mining industry. This service is based on the use of the [Personal Alpha Dosimeter] developed by the French Atomic Energy Commission (CEA). The origins of personal alpha dosimetry and its rational are briefly described. Technical and organizational aspects of a routine personal alpha dosimetry service are outlined in this paper. HISTORICAL BACKGROUND International recommendations (1) and Canadian regulations have established Maximum Permissible Exposures (MPE) for each source of radiation exposure. Uranium workers in mines and mills are exposed to external radiation ( [y] rays) and to internal radiations ( [B] and [a] particles) which are delivered to the respiratory track by airborne alpha emitters (Rn and Th daughters and Long Lived Dust). To date, dosimetry for uranium workers has been performed by area monitoring/collective dosimetry. In North America the concentration of radon daughters is routinely measured by grab samples taken at the work place and by on-site gross alpha counting. The concentration of potential alpha energy is then calculated (usually by Kusnetz method) and expressed in Working Levels (WL). The time spent by each worker at a given work place is determined from his time sheets and used to calculate the individual monthly exposures to airborne alpha emitters, which is then expressed in Working Level Months (WLM). The uncertainties attached to such a procedure are obvious even in the case of frequent grab samplings and can be expected to lead to an underestimation of individual doses. Among fifteen possible sources identified in a mine situation, (2) four may stretch the standard deviation of the measurements' distribution, nine may lead to an underestimation and two may lead to either an underestimation or to an overestimation. To improve this situation, in 1971 the Atomic Energy Commission began studying the use of personal alpha dosimeters to determine individual exposures from the airborne alpha emitters encountered in the uranium industry environments. Criteria for a Personal Alpha Dosimeter In order to minimize the difficulties encountered in determining exposures received by uranium workers, the CEA in co-operation with the Atomic Energy Control Board of Canada (AECB), has developed a set of criteria for personal alpha dosimeters. Exposures may be determined easily and accurately using this criteria. Autonomy The dosimeter must operate for at least 10 to 12 hours. Excess time spent in the mine or in the facility may possibly be related to an accidental situation causing unusual levels of radioactivity. Since the dosimeter may be needed in non-underground settings where a cap lamp is not used, full autonomy is desirable. Maintenance, Periodicity of Reading In order to complement other dosimetry systems, the personal alpha dosimeter should be read monthly when the filter should also be changed. Routine air flow checks can be made according to local conditions (e.g. diesel loading). Radioisotopes Identification Since the exposure unit (WLM) is based on the concentration of potential alpha energy in the air, the personal alpha dosimeter should be capable of identifying each short lived alpha emitter included in the calculation of the WI, and WLM. Permanent Exposure Record Three points may be considered here: 1. In many countries, lung cancer in uranium workers is a compensable occupational disease. In some instances, compensation is awarded when it can be proven that the worker has received an exposure above a certain limit. The present uncertainty of the individual exposure makes the compensation procedure difficult. 2. By design, a personal alpha dosimeter must representatively sample all airborne particles, ranging in size from the unattached fraction to the upper limit of respirable aerosols (0.001 to 5 µm). The dosimeter must offer minimal resistance to the penetration of these aerosols. While the mining/ milling environment presents harsh conditions which may accidentally contaminate the dosimeter, it is important to be able to distinguish these cases of contamination and still obtain accurate readings. 3. A dependable dose register is most valuable for further epidimiological studies. The dependability of such a data base increases with the possibility of a second assessment of the dosimeters' reading (filter, film).

Jan 1, 1981

By Earl M. Frizzell, Ronald W. Utley

Introduction For the past several years the Bureau of Mines has undertaken research to develop alternative mine material haulage systems that will improve mine productivity, reduce haulage costs, increase mineral availability, and conserve energy. One basic concept being pursued is a means for crushing primary blasted rock to a size that can be reliably and economically moved on conveyors. During 1980, mine requirements for this equipment and the capability of existing technology to satisfy these requirements were assessed. The study continued into 1981 on the basis that satisfactory equipment for this purpose needed to be developed. The objective of the 1981 program was to perform an engineering design of such a system. A mine survey and equipment development study was conducted in cooperation with the open-pit mining industry by GATX through its subsidiaries, Fuller Co. (a crusher manufacturer) and GARD Inc. (GATX's research and development center), under contract to the USBM. Results of a mine survey in the copper, iron, and molybdenum industries indicated that a universal crusher system could be developed even though details might vary from mine to mine. The crushing characteristics, including tonnages desired, were similar in all surveyed mines. Most mines had similar views about the desirability and frequency of crusher movement. The perceived need for in-nit crushing varied between copper and iron mining. The copper mines felt a greater sense of urgency for introduction of such a system, probably because typically greater vertical haulage is found in copper mines than in the iron mines. All mine personnel surveyed were interested in portable crushers and almost universally preferred gyratory-type primary crushers. They preferred large crushers, citing inherent reliability and durability, and the machine's relative freedom from clogging. Large size minimizes the number of crushers and conveyors needed. Analysis of desired crusher system characteristics (Table 1) fixed the crusher size at 1.5 x 2.3 m (60 x 89 in.). The 1.5-m (60-in.) feed-opening dimension could accept the largest material normally fed to the crusher, and the 2.3-m (894n.) head diameter could fulfill the 3.63 kt/h (4,000 stph) capacity requirement. Selected System Description The comments and analyses discussed in the contract report led to the design and refinement of a movable crusher system. This system represents the consensus of more than 100 mine workers and almost 50 construction and engineering company personnel. It is supported by more than 50 years of experience in crusher manufacturing. Although many alternatives are available for any given system, and selections are likely to be dictated by varying site conditions, the selected system is believed to be appropriate for most projected requirements. Final design would be dictated by mine operators during evaluation for approved projects. Detail design of the feed and discharge systems were not included in the study because they are mine specific. The basic system (Fig. 1) consists of a 1.5 x 2.3-m (60 x 89-in.) gyratory crusher on a structural steel frame. The frame is used as an operating platform for the crusher and as a carriage to move the crusher. The control tower includes a crusher lubrication room, a hydraulic equipment room, a switch gear room, a room for the motor control center, and a control room for operating the crusher and maintaining the equipment. The auxiliary system consists of a hydraulic rock hammer and pedestal crane with 74-t (80-st) and 14-t (15-st) hooks. Figure 1 also shows stairs, railings, and a work platform. Feed and discharge arrangements are shown on Fig. 2. The system shown is for mildclimate applications. Additional enclosure and heating would be needed for extremely cold weather conditions. Crusher A 1.5 x 2.3-m (60 x 89-in.) gyra¬tory crusher was selected for this application. This size can handle the typical throughput of 3-3.63 kt/h (3,500-4,000 stph) desired by the mines and the 1.3-m (50-in.) material to be crushed without difficulty. The crusher is the heaviest

Jan 4, 1983

By Syd S. Peng

Introduction Modern longwall mining, introduced to the US coal industry in the mid-1960s, is the latest coal mining technique. Today, longwall mining produces more than 15% of all underground coal production. The growth of longwall mining in the US is slow. However, nearly two decades of longwall mining have demonstrated that its benefits outweigh the disadvantages. Longwall Advantages Production and Economics - A typical longwall production ranges mostly from 0.9 to 1.8 kt (1000 to 2000 st) of clean coal per shift. This is about three to six times the production of comparable continuous mining units. Furthermore, there are much less rejects in longwall mining, typically 15% to 25%. It must be noted, however, that a new longwall with an inexperienced crew may produce less, depending on mining conditions. But as experience gains, production increases rapidly. Due to its high production, the cost per ton of coal mined over the mine's lifetime is cheaper with longwall mining than with continuous mining. A study on a conceptual mine of 25 year life using actual operational data (Dangerfield, 1981) concluded that the cost per ton of coal mined is 32% cheaper by longwall than by continuous mining. Another benefit from its high production potential is that it enables the production to be concentrated in a few longwall faces versus many units of continuous mining required to achieve a similar production. As a result, mine organization and management simplify considerably. Safety and Savings - Most US longwalls use shield supports that cover the roof with solid canopy and isolate a gob completely from the face area. Thus, if the supports are properly selected and operated, the potential for roof fall accident is almost zero. Furthermore, it is not unusual that a well-run longwall face is much more orderly and cleaner that most well-run machine shops. The ventilation system for longwall mining is the most ideal type. It is simple and unique. No auxiliary fans are required. The fresh air sweeps through the whole face. Its volume and velocity can be adjusted as demanded. Recovery - In spite of multiple entries layout for US longwall panels, the longwall method recovers up to 40% more coal than the continuous mining method (Dangerfield, 1981). The typical recovery for longwall mining is 70% to 85% versus 45% to 80% for continuous mining. Versatility - Longwall mining has been employed successfully in seam heights ranging from 0.8 to 3.7 m (2.7 to 12 ft); in seam inclinations from horizontal to 35°; in overburden depths from 46 to 914 m (150 to 3000 ft); in single or multiple seam mining; and in uniform or irregular seam characteristics. For seam height less than 1.2 m (4 ft), the plow is used instead of a shearer. However, in spite of recent developments, the plow still is unsuitable for hard coal seams. Slow Acceptance How come longwall mining, with so many advantages, has not spread more rapidly? Undoubtedly, the major reason is the large capital investment required. Under normal market conditions, the capital required for a longwall panel ranges from $7 to $11 million. Another factor is the uncertainty of its applicability to any specific coal seam of interest where longwall has not been used or operated successfully. Although longwall mining elsewhere has proven its applicability to wide-ranging seam conditions, there have been several failures in the past two decades. Analyses of those cases show that most failures can be attributed to inexperience, both operational and technical, including equipment selection. Therefore, personnel training on the longwall mining technique is absolutely necessary before its initial operation. Some earlier statements about adverse surface subsidence associated with longwall mining were misleading. Research has proven that surface subsidence under longwall mining is controllable and predictable. And, if longwall mining was mandatory since the onset of coal mining in the US, there would be no surface subsidence problems over abandoned mined land today. It is a blessing that surface subsidence occurs

Jan 3, 1985

By Henry O. Ash

On June 16, 1982, Rio Blanco submitted a request to the US Minerals Management Service for a suspension of operations under its federal oil shale lease. The company also wanted a five-year delay in the requirement for further minimum royalty payments and a five-year extension of the primary lease terms of 20 years. The request was granted in late July in the midst of yet another slump in oil shale activity. Rio Blanco holds one of four Federal oil shale leases that were granted under the Interior Department's Prototype Oil Shale Leasing Program in 1974. That year the DOI offered for lease six federal oil shale tracts, two each in Colorado, Utah, and Wyoming. The leases in Colorado and Utah were granted for substantial bo¬nus bids as shown in Table 1, but the Wyoming tracts had no takers. In the eight years hence, no oil shale has been produced commercially. Indeed, the prospects for successful oil shale development have fluctuated widely, continuing a 60-year- old cyclical pattern in the fledgling industry. The prototype program was developed and initiated while this infant industry struggled to begin anew, without the benefit of any federal oil shale leasing regulations, or any history of federal oil shale leasing at all. Permanent regulations were not required because the effort was limited and designed to provide the foundation for any future leasing program. Section 21 of the Mineral Leasing Act of 1920, a Nov. 30,1973 Federal Register notice, and the leases themselves constitute the legal and regulatory framework for leasing and developing prototype oil shale tracts. The prototype leases contain some unusual provisions including installment payments of the bonus bids, credit of early development costs against payments due the government, adjustment of royalty rates according to regional crude oil price fluctuations, and unusually extensive and detailed environmental stipulations. Several of the provisions were intended to provide incentives for early development and financial penalties for failure to develop. The requirement for minimum royalty payments in lieu of actual production royalties beginning in the sixth year of the lease falls in the latter category. After the sixth year, the required minimum payment is increased east year through the 15th year by an increment equal to that due in the sixth year. The minimum royalty then remains the same through the final five years of the primary lease term. Minimum royalty is calculates on a different production level for each prototype lease, based on the quantity of 125 L/t (30 gal/st) shale estimated to be recoverable from. each tract by room-and-pillar mining. That mining system was selected for reference as the only proven extraction method for oil shale at that time (early 1970s) The minimum royalty (production) rate was calculated for each tract by direct comparison of its estimated recoverable 125-L (30-gal) resource to a hypothetical tract containing 1.9 Gt (2.1 billion st) of recoverable 125-L (30-gal) shale. For the hypothetical lease tract, the minimum production for the sixth year was calculated to be 3.2 kt/d (3,500 stpd), increasing by a like amount each lease year through the 15th to a daily production rate of 31.7 kt (35,000 st). Under this formula, the hypothetical lease tract would yield $16 million in minimum royalty payments over the 20-year primary lease term if there were no production, no development credits taken, and no adjustment of the royalty rate due to fluctuation of regional crude oil prices. The production rate of 31.7 kt/d (35,000 stpd) was a somewhat arbitrary figure that was considered, in the early 1970s, to represent a minimum size for a commercial oil shale operation. A production rate of 7.9 dam3/d (50,000 bbls/day) soon supplanted that smaller mine production figure as a rule of thumb for the minimum size for a commercial facility. The proportional relation-

Jan 11, 1982

By Joseph R. Hatch

Previous studies of the Thompson and Temperly ore bodies show alteration aureoles in the Middle Ordovician host-rock limestones, dolomites and shales extending laterally as much as 60m from the ore. Thickness, texture, carbonate and clay mineralogy, major-, minor-, and trace-element contents; and carbonate carbon and carbonate oxygen Isotope compositions of the host rocks all have been significantly altered by hydrothermal processes. Organic geochemical analyses of samples from the basal 8-cm-thick carbonaceous shale bed and immediately overlying limestone bed of the Quimbys Mill Member of the Platteville Formation, collected from within the ore, the alteration aureole, and the unaltered country rock show: 1) solution of carbonates and residual concentration of organic matter, 2) decreases in the volatile (Si) and pyrolytic (S2) hydrocarbon yields In samples of the basal shale from within the ore, 3) decreases in the relative amounts of the lower molecular weight (<nC25) saturated hydrocarbons in samples from the ore, and 4) no changes in either the organic carbon isotope composition or the Rock-Eval or biomarker thermal maturation Indicators. In light of the 125-170°C depositional temperatures documented for the early marmatitic sphalerltes in these deposits by fluid Inclusion studies, the lack of thermal alteration of the organic matter suggests a relatively short mineralization episode, a conclusion supported by the relatively simple sphalerite band stratigraphy. The decreases in the volatile and pyrolytic hydrocarbon yields for shale samples from the ore supports a model for mineral deposition which obtains sulfide through reaction of brine sulfate and indigenous organic matter.

Jan 1, 2013

By B. K. Gupta

Arch Mineral&apos;s Captain Mine has used the Bucket Wheel Excavator System(BWE) for over 3 years in conjunction with a 180 yd3 (138M3) stripping shovel. The bucket wheels prestrip the soft overburden ahead of the shovel. This soft overburden is brought to the spoil side through a series of shiftable conveyors and placed by a spreader over the ungraded or partially graded shovel spoils. The placement of the soft overburden and top soil is accomplished in a manner that meets the reclamation mandates. The prestripping by the BWE has reduced the incidence of the highwall and spoil slides, depth the stripping shovel has to dig and the volume handled by the pullback draglines.

Jan 1, 1984

By John S. Cummings

Although there are many geologic situations which might be discussed at a session such as this, as I considered the problem, it became apparent to me that I should discuss a situation which has been of concern to me, and should be of concern to the geologic community. As a result, the questions which I will pose, and the criticisms, have not been arrived at quickly or easily, but have accumulated through time, to the point where I feel discussion is necessary. Very briefly let me say that my criticism of geologic mapping, of conclusions reached, and objectives or lack of such, results from my association with many areas where such studies have been completed, and resultingly my inability to comprehend the system of evidence or logic which supports extensive age-formational designations. The heart of the problem is that many conclusions resulting in geologic designations are based on weak or non-existent evidence, and such should be catagorized as hypotheses. Although the geologic community as a group does not approve or disapprove of such designations, it is evident that many or most of such designations attain a "status", and arc accepted as more or less valid statements. As will be shown, the amount of specific lithologic data provided in some quadrangle mapping projects is exceedingly slim. However, of concern herein is that the type of information presented at any stage is sufficient to warrant conclusions drawn. To see the magnitude of the problem, it is probably necessary to review some basic fundamentals upon which our geologic systems are established. The United States Geological Survey has defined a "formation" as the broad unit of geologic mapping consisting of a large and persistent stratum of one kind of rock. Obviously such a designation allows that a formation could be considered "time variant" in a horizontal as well as in a vertical context. However, other definitions enable a formation to mean units

Jan 1, 1972

By M. I. M. Darwish

In this paper, an effort was made to get more in-sight about the parameters that can affect air sparging technique optimization. Numerical simulations were used in predicting the effects of different parameters utilizing Shell?s in house MoRes simulator. The case of using air sparging to remediate source zone is considered in this work. Different air injection scenarios using continuous and pulsed air injection techniques, with different frequencies and durations, were compared. The amount of MTBE mass removed from the saturated zone and arrived to the ground surface, through the vadose zone, was the main criterion while comparing different air injection scenarios. The simulations results showed that pulsed air injection is better than continuous air injection in removing the MTBE and that the longer the OFF time is the better the MTBE removal.

Jan 1, 2008

By C. Cassios

EIA is a procedure aiming the elimination of adverse impacts to the environment from undertaken projects. Historically, it starts in 1969 when NEPA (Nat.Env.Prot.Agency) in the US initiated the EIA ? EIStatement. The EIA procedure was adapted by the European Union with the directive 337/85 and in Greece the Common Ministerial Decision 69269/90 enforced it to the Greek legislation. However, mining enterprises in Greece long before the implementation of the European Directive and in accordance to the legislation of the Public Law 998/79 for the protection of the forested lands were obliged to submit EIA?s and get an approval in order to operate in forested areas. The submitted and approved EIA?s and their implementation by the mining companies were among the positive aspects of the EIA procedures. This paper is examining the impacts of EIA?s in the Greek mining industry its positive aspects as well as their difficulties and their role for the reclamation procedure. The technical and procedural aspects which EIA presents for the mining industry are commented, while proposals, for a more flexible EIA implementation for the Greek Environment are also presented.

Jan 1, 2002

By T. E. Durney

In the Fall of 1987 The Wellmore Coal division of The United Company installed a fines separation circuit in its #20 plant near Big Rock, Va. The addition of the fines circuit included a bank of classifying cyclones and two flotation units; one a conventional 4 cell, 34 m3 Wemco cell, the other an 2.4 meter diameter by 7.6 meter tall, 28.3 m3 column flotation cell. The new fine coal separation circuit is depicted in Figure 1. The column flotation cell is manufactured by the Deister Concentrator Company, Inc., under the trade name Flotaire. The Deister Flotaire column cell utilizes a counter current flow arrangement and a different means of generating air bubbles than that of conventional mechanical flotation cells found in coal processing plants. Other features include diffusers for additional air supply and a froth washer. This cell is one of the few commercial scale column cells processing coal.

Jan 1, 1990

By C. O. Brawner

INTRODUCTION My first major involvement on a rock mine slope stability project was in 1966 after I had become a principal of Golder Brawner & Associates in Vancouver, Canada. I received a request from Kennecott to evaluate two serious rock slope failures adjacent to the concentrator on their British Columbia molybdenum property near Alice Arm, BC, which was under construction. A rock slope up to 200 ft* high had been excavated at a design angle of 700. The inspection indicated that several large blocks of about 1000 cu yd each had failed on steeply inclined joint surfaces. Stability was reduced by extensive seepage caused by heavy rainfall. In addition, considerable overbreak had occurred, indicating excessive blasting had been used to excavate the slope. A program was developed to rock bolt potentially unstable blocks that could slide on dipping discontinuities. Some rock was scaled back to the inclined joints, and drain holes were drilled to reduce water pressures. The project was completed successfully and within reasonable economics. As a result of this success, enquiries were received regarding other mining projects. This indicated the potential demand for rock slope engineering. As a result, I took one month off to meet and discuss rock mechanics and slope stability with various people internationally. These included D. Coates in Canada, E. Hoek in England, K. John and L. Mueller in Germany, M. Rocha in Portugal, and the US Bureau of Mines in Denver, CO. Early misconceptions Based on involvement in projects during the next few years, it was concluded that there were a considerable number of misconceptions that existed among some geologists and mining engineers dealing with surface rock mechanics. Some of these included: • Water is a lubricant. This is incorrect. Water reduces the shear strength along discontinuities due to buoyancy, which reduces frictional strength. It also reduces the cohesion of clayey-type rocks. • The optimum pit slope angle in rock is 45°. The optimum slope angle for safety and economics is variable and is influenced particularly by structural geology, water pressures, mine blasting practices, and the depth and degree of weathering. • The flattest angle for slope stability is dictated by the angle of repose of broken rock, about 370. Where clayey-type rocks or clayey infill in discontinuities exists, failure can occur at much flatter angles, requiring flatter slopes or artificial stabilization. • Uniaxial and triaxial tests provide the necessary shear strength parameters to design rock slopes. Most failures in rock slopes are controlled by directional discontinuities. Therefore, the direct shear test is the correct test to evaluate strength parameters for most problems.

Jan 1, 1987

By Lawrence E. Mannion

In southwest Wyoming lies buried the world&apos;s largest source of natural sodium carbonate, or soda ash. Perhaps a hundred billion tons of trona (sodium sesquicarbonate) are contained in ancient lake beds in the Green River basin&apos;). There the Green River Formation encloses more than 40 extensive beds of trona. At least 25 of these exceed 3 feet (1 meter) in thickness and most of them are more than 300 square miles in extent. Five have been exposed in mines and lie between 800 and 1500 feet below the surface. Others lie at depths of from 400 to 3500 feet. Three companies have underground mines that supply trona to adjacent soda ash plants. These are the FMC Corporation, Allied Chemical Corporation, and Stauffer Chemical Company. A fourth, Texasgulf Corporation, has a mine under development. In 1974, annual soda ash capacity in the Green River area was approximately 4,300,000 tons, which requires the mining of about 7,500,000 tons of trona.

Jan 1, 1975

By Marcel Michaud

Over the past five years, the two-fold growth in Canada&apos;s coal industry "Figure, 1" results primarily from the coking coal market to Japan and, to a lesser degree, by the increasing use of sub-bituminous and lignite coal for power generation in the Prairie Provinces. This expansion will continue, but production of coal for domestic thermal generation of electricity will grow at a faster rate than coking coal exports "Figure, 2". Half of the 20 million tons of thermal coal used for electrical power generation in Canada during 1974 were imported by Ontario Hydro from U.S. coal companies in West Virginia and Pennsylvania. Ontario Hydro&apos;s Nanticoke station on Lake Erie which will consume 7.5 million tons of coal in 1977 when fully operative, as well as other future planned projects, could cause Ontario Hydro&apos;s coal consumption to rise from 10 million tons today to 20 to 25 million tons by 1985

Jan 1, 1975