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By James D. Hervol, Josepph W. III Leonard, Barbara J. Arnold

INTRODUCTION Dry concentration or pneumatic cleaning uses air currents as the separating medium. Pneumatic cleaning in the United States today is limited to three or four operations. In fact, the quantity of bituminous coal cleaned pneumatically has declined from its peak in 1965 of about 25.4 million tons (23.1 Mt) to its current level (1988) of less than 4 million tons (3.64 Mt) annually, as shown in Fig. 7-9 1. In 1940, pneumatic methods accounted for 15% of the bituminous coal and lignite cleaned in the United States as given in Table 7-32. Currently, pneumatic methods account for less than 2%. The largest air plant was built in Pennsylvania about 1968 with 14 units processing 1400 tph (1280 t/h) of 3/, in. X 0 (19 mm X 0) raw feed and two additional units retreating middlings. This same plant is one of only three or four plants still operating air-cleaning devices in the US. A wet plant was added in 1985 to increase the efficiency of cleaning the fines. The air plant is still used as a scalping operation for the coarse material. Five additional operations are either no longer in production or have discontinued the use of air tables within the last year. Wet cleaning methods are now being employed at the facilities that have discontinued air table operation. he advantage of dry cleaning techniques is the low moisture content of the clean coal product. However, to decrease respirable dust, mining operations must use water sprays on modem mining equipment. This water adds to the moisture content of the run-of-mine coal and decreases the effectiveness of dry cleaning processes that rely on low surface moistures in the feed to allow for good separations. Wet cleaning methods are preferred, especially since they also allow for much sharper separations. The advantage of dry cleaning techniques is the low moisture content of the clean coal product. However, to decrease respirable dust, mining operations must use water sprays on modem mining equipment. This water adds to the moisture content of the run-of-mine coal and decreases the effectiveness of dry cleaning processes that rely on low surface moistures in the feed to allow for good separations. Wet cleaning methods are preferred, especially since they also allow for much sharper separations. Despite dry concentration's drawbacks, research has continued into dry concentrating methods, including both electrostatic and magnetic separation techniques. The real application for future dry cleaning processes is probably at the power plant between the pulverizer and the boiler feed where particles are already dry and ground to fine sizes (- 200 mesh [ -0.075 mm]), increasing the effectiveness of the separation and also, possibly, increasing the liberation of clean coal from finely disseminated impurities. From a utility's viewpoint, pneumatic cleaning offers the best delivered Btu cost, based on the premise that a percent of moisture is just as detrimental as a percent of ash. Additionally, several other advantages are inherent in dry concentration techniques. The main advantage is that dry concentration re-

Jan 1, 1991

By J. Killian, S. Cox

Many questions exist within the geotechnical industry as to the applicability of Strength Reduction Factor (SRF) values to the design acceptance criteria (DAC) of mining operations. There are many factors that contribute to changes in the SRF estimate, some of which are numerical (e.g. dilation angle) and some of which are geometric (e.g. model grid size)[1]. Numerous studies have confirmed equivalent values of SRF and FOS are found for soil slopes of limited height as presented in, [2] and [3]. However, significant deviation of the SRF value was observed for highly cohesive materials (rock slopes) of heights typically encountered in mining, [4]. This paper builds on the work presented in [4], providing a more rigorous examination of the modeling parameters that control the SRF estimate. The failure mechanism assumed for this study was a simple circular shear surface in a uniform material. Both LEM and numerical methods are suited well for analysis of this failure mechanism. A parametric study was completed to evaluate and test the relationships of key numerical variables that affect the estimated value of the SRF using Itasca’s Fast Lagrangian Analysis of Continua (FLACTM) software. The effect of groundwater has been tested and may be significant but is not presented as part of this body of work for brevity and simplicity of presentation.

Mar 2, 2022

By G. E. Bandelis

In this paper, solid bed erosion processes and the mechanisms involved when a two-phase liquid-solid mixture flows over a bed of solids in a horizontal annulus are analyzed. There are few such studies available with some authors arguing that the flow rate for erosion is always less than the flow rate for the suspension of solid particles thus necessitating further study and analysis of such phenomena. Tests were run to determine the mechanism of erosion and the erosion velocities of the solid bed formed in a concentric annulus for three flu-ids, water, a low concentration water solution of Carboxyl-methyl-cellulose (CMC), which behaved as a Newtonian fluid and a higher concentration CMC water solution which behaved as a power law fluid. Glass spheres of 2 mm nominal diameter were used as solids. The flow rate at which erosion occurred together with the mechanism of bed erosion were determined visually, while the whole process was video recorded. Our tests indicated different erosion mechanisms and different erosion velocities for the three tested liquids, with the liquid viscosity playing a major role. The first mechanism, observed for water flow, gives erosion from the front of the bed while the second mechanism, observed for the flow of the two viscous fluids, gives erosion from the top of the bed. The role of each of the parameters to erosion is investigated, whether it is turbulent flow with the turbulent fluctuations being more capable to remove the solids from the formed bed or whether it is liquid viscosity, which results in the hydrodynamic shearing force. The results demonstrate that liquid viscosity plays a role in the erosion mechanism and yields different erosion velocities of the solid bed hence, it should be taken into account when modeling flow pattern transitions for solid-liquid flows. Prior studies have not identified this and it could be the reason of the disagreement between theoretical models and experimental and field data.

Jan 1, 2006

By T. B. Whitmoyer, J. E. Sneckenberger, M. A. Jerabek, C. F. Stanley

The measurement of acoustic impedance at frequencies up to 6,400 Hz has been proposed as a diagnostic tool to yield more detailed information about lung condition. A three generation branching system of plastic Y-tubes has been constructed. The impedance of the branched system is measured by an impedance tube using the two-microphone technique. The impedance spectrum is used to calculate the reflection coefficient spectrum, which is converted to an impulse response by an inverse FFT. The area function is extracted from the impulse response by the Ware-Aki algorithm. Measurements of acoustic impedance of the branched system were made for several different end conditions of the eight terminal outlets. The outlets were either blocked with modeling clay, plastic packing material, or aluminum blocks or were unblocked. Six measurements were made of each configuration. In general, the more outlets blocked, the larger the shift to the right of the impedance magnitude peaks for frequencies above 3000 Hz. The impedance magnitude peaks for the plastic packing material were shifted in frequency to the right relative to the peaks of the other system terminations. The impedance magnitude peaks for the clay and aluminum system terminations were significantly different in magnitude but not shifted in frequency. The area functions at the end of the branched system were significantly different (a > 99%) for most of the end conditions. Area functions for the modeled branched lung systems between the end of the impedance tube and the termination outlets of the branched system were similar but distinctively different for the various end conditions. These results indicate that the two-microphone technique can sense small changes in the terminating end conditions of a rigid system of branched tubes and thus may sense impedance changes due to lung disease in the conducting airways. Keywords: acoustic impedance, lung model, area function.

Jan 1, 1991

By William A. Dupree, Mark J. Schultz, William J. Francart

Maintaining intake escapeways clear of contaminants from mine fires has been an issue for many years. Removing fire sources from these entries reduces the risk of contamination greatly. In recent years, concerns have been raised related to contamination of the intake escapeways by leakage when fires occur in adjacent airways. This paper discusses measured leakage rates for multiple stoppings at various pressure drops across stopping lines in a number of actual coal mines. Also addressed are methods for reducing the contamination risks for the intake escapeways, considering the ventilation system employed. The Effects of Air Leakage The effects of air leakage through ventilation controls has been well documented. While normal leakage is expected through stoppings, overcasts and other structures, abnormally high leakage results in an inefficient ventilation system and can increase operational costs for a mine. In addition, excessive leakage results in less control over mine ventilation airsplits. This may compromise the isolation of mine escapeways, and allow adjacent entries to be rapidly contaminated by products of combustion in the event of a mine fire. A study was conducted at ten mines in which air pressure and quantity measurements were taken near the active sections. The method used to construct the ventilation controls to separate airways was also noted. Stopping Construction Table 1 shows the type of stopping construction commonly used in the 10 mines included in this study. Main entry stoppings were normally built with mortared joints (wet stacked) with sealant applied to the high pressure side of the stopping. In most cases the entire face of the stoppings were sealed. In one case only the joints between blocks were sealed but the perimeters were sealed in all cases. Some mines also sealed the low pressure side of the stopping. Solid blocks were normally used for main entry construction and hollow blocks were sometimes used for longwall panel development. [ ]

Jan 1, 1993

By Timothy Batchler, Khaled Mohamed, Morgan Sears

The National Institute for Occupational Safety and Health (NIOSH) initiated a testing program to assess the effects of lateral loading on various standing supports used in underground coal mines. This study is not intended to endorse a particular type of standing supports nor to provide a comparison between different types of standing supports. This paper aims to measure the lateral loading capacity of two types of standing supports and evaluate how it affects their vertical loading capacity and yielding mechanisms. The study utilizes the Mine Roof Simulator (MRS) at NIOSH Pittsburgh. Type-1 support is characterized by telescopic tubes made from ASTM A-513 steel with a specified locking mechanism, and Type-2 support comprises a single component constructed from ASTM A500 Welded Grade C pipe. Vertical load tests were carried out to assess the designed performance of Type-1 and Type-2 props. Type-1 props demonstrated a "constant yielding" behavior, achieving an ultimate capacity of up to 140 kips. Type-2 props showed a "load shedding" behavior, reaching peak verti- cal loads of up to 191 kips. Buckling analysis revealed that Type-1 props were not designed to buckle, and Type-2 props failed due to buck- ling rather than material yielding. Under vertical pre-loading, lateral loading tests showed that Type-1 props can withstand about 9 kips, and Type-2 props can withstand about 10.79 kips of lateral load. This represents roughly 6% of their vertical loading capacities. Testing props under lateral loads in addition to their intended vertical loads showed that the lateral loads could influence their loading mechanisms. Type-1 props did not exhibit the expected "constant yielding" behavior under lateral load but instead showed "load shedding" decreasing their rated vertical capacity by 70%. At high vertical pre-loading, Type-2 props demonstrated "constant yielding" behavior instead of the anticipated "load shedding," leading to a 60% reduction in their rated vertical capacity. These findings provide valuable insights into the behavior of these props under different loading conditions, aiding in their efficient use and design in underground mining operations.

Jun 25, 2024

By Richard L. Breithaupt

INTRODUCTION A shuttle car is a self-propelled rubber-tired haul¬age vehicle designed specifically for underground min¬ing, primarily in coal mines (see Figs. 1 and 2). The drive mechanism is located on the sides of the vehicle, and the load is carried in the center. A chain-and-flight conveyor distributes the load when the car is being loaded by a loading machine or continuous miner and also discharges the load onto a conveyor belt or into a mine car. As the name implies, the shuttle car shuttles back and forth between the working face and the un¬loading point and is not required to turn around. Shuttle cars have served as primary haulage in this country since 1940 and still perform this function in most coal mines today. TYPES OF SHUTTLE CARS Shuttle cars may be battery-powered, diesel-powered, or cable-reel electric powered. Battery-powered cars were used between 1940 and 1950, but are almost out of existence today because of their limited power. Diesel cars have been used in coal mines, but are used mostly in noncoal mines where rules and regulations are not as stringent. Also, objections by organized labor have limited the use of diesel-powered shuttle cars in coal mines. Today, the largest number of shuttle cars are cable-reel cars. Electric power is fed through a cable stored on a cable reel that may have a capacity from 152 to 244 m (500 to 800 ft), depending upon the size of the cable and the cable reel (see Table 1). This, of course, limits the travel of a shuttle car. However, time studies have shown that for the most efficient haulage cycle, the length of haulage should not exceed 91 m (300 ft). As haulage distance increases, the waiting time for the loader or miner becomes longer, and haul¬age efficiency suffers. Cable-reel cars are manufactured in standard hand and opposite standard hand models. A standard car has the operator's control on the left front side and the opposite standard car has the opera¬tor's controls on the right front. Generally a loader or miner operates with two shuttle cars, one of each type. This allows the trailing cable to lie on either side of the entry and prevents the cables from fouling each other. Shuttle cars are designed to operate with either dc or ac power. Cars using dc power are designed for 250-v supply and ac cars are designed for either 440, 550, or 950 v, three-phase voltage. Most coal mines that are relatively level can use dc cars where dynamic brak¬ing is not required. For mines having steep grades (in excess of 15%) ac cars are used because ac traction motors have inherent dynamic braking and are safer when tramming downgrade. A recent innovation in shuttle car design in the last five years has been the introduction of an ac/dc car. This design utilizes ac input power through the cable, and the power is rectified on the car to 250-v do for traction motors. The pump motor and conveyor motor are powered by ac voltage. The main advantage of this design is efficiency-a much smaller cable may be used and the smooth operating characteristics of the dc traction motor is obtained. This prevents shock loading on the traction drive train, which is an undesirable feature of ac traction motors. SHUTTLE CAR FEATURES Elevating Discharge Today, all shuttle cars are equipped with an elevating front conveyor that allows the shuttle car to discharge into a mine car or belt feeder without the use of ramps. Hydraulic Power System Today, all shuttle cars are equipped with an hy¬draulic power system. An hydraulic pump is driven by an electric motor which provides hydraulic power for steering, elevating the front conveyor, and powering the

Jan 1, 1982

By Henry R. Tiedemann, Eugene H. Skinner, George E. Wickham

The need for ground support research in underground excavation has been well identified in recent years by many studies in the subject area (1, 2, 3, 4, 5). A typical conclusion from these reports (1, 2) is that the drill-and-blast (D & B) mining or the mechanical tunnel boring machine (TBM) methods are expected to be continued in use for the needs of tunneling in the immediate future. The Organization for Economic Co-operation and Development (OECD) sponsored in June, 1970 advisory conferences on Hardrock and Soft Ground Tunneling (3,4). These two reports were based on the combined response received from approximately six hundred inquiries from 17 nations in the summer of 1969 and which concerned all aspects of present-day inadequacies of, and the needed improvements in the technology of tunneling. Some specific problem areas in ground support and tunnel linings were: 1) the ability to successfully predict rock loads that must be carried by the support system; and 2) inadequate methods for determining the load-carrying capacity of the various support systems. Therefore, the problem area is clearly defined by these two objectives: 1) To find means of predicting rock loads; and 2) To develop methodology for determining safe, efficient, and economical ground support design practices. Noting these research needs from a great number of sources, the United States Bureau of Mines, Spokane Mining Research Center (SMRC), acting as agent for the Advanced Research Projects Agency (ARPA), undertook contract research effort in these problem areas. Work by Jacobs Associates has studied a methodology for specification of temporary artificial support based on the dependent sequence of mining operations and prediction of various rock conditions in the pre-excavation stage. The major thesis is that rock behavior is the

Jan 1, 1997

By M. P. Evans, Rainer Kahrger

During the last decade or so great advances were made in the solvent extraction, electro winning technologies. They, combined with heap leaching, made it viable to exploit large, relatively low grade oxide copper deposits. The El Abra mine is a good example for such a deposit. Situated in Northern Chile at an elevation of about 4000m (14000 feet), 70 krn north of Calama, the El Abra deposit has been known for many years. In the early 1990's Codelco decided to sell a major holding in the deposit to international mining companies. After lengthy negotiations Cyprus Minerals finalized in June 1994 the purchase of 51% of the deposit with Codelco retaining the balance. lance. Already in 1993 Cyprus and the engineer on the project, began discussions with international materials handling companies with regards to the materials handling system required for the project. These early discussions were considered critical for this fast track project. Immediately after the notice to proceed with the project, several companies were invited to submit competitive conceptual proposals for that portion of the material handling system which was to transport crushed and treated ore from the plant to the leach pad and build the leaching heaps at a rate of up to 150 000 tons per day. A supplier was selected. Additional studies were made for the conveyor system from the mine to the fine crushing plant. Due to the extremely rugged terrain and high altitude at the mine, a decision was made to locate the plant and the leaching operations 15 km east of the mine at an elevation of 3400m. In July 1994 the supplier started the conceptual engineering with a design team at the engineer's San Francisco office. The undertaking was unique in the sense that the supplier's engineering team was integrated into the project for the duration of the conceptual engineering. The open and continuous communication greatly contributed to the success of the design work.

Jan 1, 1996

By B. P. Boisen

INTRODUCTION Application Measurement of rock deformation is essential to stability and behavioral monitoring of underground openings. The difficulties of obtaining adequate sample distribution, density, and orientation can frequently be overcome by the use of single and multiple position borehole extensometers. In its simplest form, a borehole extensometer consists of an instrument head, usually anchored in the collar of a borehole, and a number of in-hole anchors, each of which is fixed in place at a known distance from the collar. As rock adjacent to the borehole deforms, the component parallel to the borehole direction causes a distance change between adjacent in-hole anchors, as well as a distance change between individual anchors and the instrument head. These changes are measured by means of a mechanical link (usually wire or rod) extending to the instrument head from each anchor. The resulting data can be used to compute the distribution, magnitude, rate, and acceleration of deformation in the material intersected by the extensometer's borehole. Wire Vs. Rod Extensometers Since the early 1960s myriad varieties of borehole extensometers have become available through the combination of various special features of two basic instruments, the wire and rod versions of borehole extensometers. Recent publications have recommended shallow depth limitations for wire extensometers, and favor rod extensometers for a broad range of borehole inclinations and depths. There are three main reasons for using a deep wire system: its cost effectiveness; the fact that tremendous depth means tremendous rod weight which will cause the rod to buckle or rub against the borehole or casing wall; and users' attitude and proper understanding of how to install and apply a wire-type extensometer. Here experience plays a big role. Over the years, many have learned the hard way that untensioned rod extensometers are simple and inherently inexpensive, but limited to vertical up holes where sensitivity is not influenced by friction between the rod elements and the borehole wall or protective conduit, the inertia of the rod elements, or the direction of the movement to be measured. In deep, horizontal, or inclined holes, and in instances where compressive as well as tensile strains or displacements are anticipated, tensioned rod or wire systems are preferred. System Components Borehole extensometers are made up of at least the following five principal components: the instrument head; the collar anchor assembly; the downhole system of anchors and rods or wires; the signal cable if the head is equipped for electronic readout; and the readout apparatus, whether mechanical or electronic. These components are shown in Fig. 1. Borehole Requirements Most extensometers are designed for installation in nominal 75 to 57-mm (3.0 to 2.25-in.) diam drill holes. Either a diamond drill or a percussion drill hole is satisfactory, provided the variation in drill-hole diameter is within the tolerance of the standard anchor design. The basic system may be modified for other hole diameters by changing the dimensions of the collar anchor and downhole anchor assemblies. Some instrument heads and collar anchor assemblies require enlarged borehole collars, though usually not over 150 mm (6 in.) for a depth of 1 m (3 ft). It is desirable, although not essential, that holes be free-draining. Holes must be clean and free of debris before installation. Depth/Orientation of Extensometer Boreholes Extensometers are designed to measure axial strain components in the instrumented drill hole. As the drill hole is axially deformed, the distances between individual downhole anchors and the instrument head change, as do the distances between adjacent anchors. The distance changes measured by a multiple position borehole extensometer, for example, are reduced to functions of time, hole depth, anchor position, or other parameters of significance in terms of the structure under investigation.

Jan 1, 1982

By B. K. Cantrell, D. H. Carlson, W. F. Watts, K. L. Rubow

The use of diesel equipment in underground coal mines is controversial because the National Institute for Occupational Safety and Health (NIOSH) regards "whole diesel exhaust" as "a potential occupational carcinogen." Threshold limit values (TLVs) are recommended for the gaseous pollutants CO, CO2, NO, N02, SO2 and some hydrocarbons emitted in diesel exhaust. There is, however, no TLV recommended for diesel exhaust aerosols (DEA) nor is there a standard method for sampling these aerosols. The University of Minnesota and the USBM have collaborated to develop a personal diesel exhaust aerosol sampler (PDEAS) that utilizes size selective, inertial impaction and gravimetric analysis. During the field tests of this sampler, numerous air quality measurements were conducted in underground coal mines that used diesel equipment. The objective of this report is to present these data and to assess the impact of diesel face haulage equipment on mine air quality. Air quality data were collected in four underground coal mines. The mines used continuous miners and diesel haulage equipment. Daily samples were collected at the section intake, haulageway and on haulage vehicles. Both full-and partial-shift samples were collected using a variety of aerosol instruments. These included the PDEAS and the micro-orifice uniform deposit impactor (MOUDI) that sample the mine aerosol diesel exhaust fraction. Two of the mines were also monitored for the major gaseous constituents found in diesel exhaust CO, CO2, NO, NO2 and SO2. These measurements were made by either collecting bag samples for subsequent gas analysis or by Palmes dosimeters. Concentrations of CO, C02, NO, NO2 and SO2 were well below regulated levels. The mine average DEA concentration determined by the PDEAS in the haulageway for the three mines surveyed was 0.79 mg/m3, with a standard deviation of 0.19 mg/m3. DEA contributed 65% of the respirable aerosol, at this location. These results were the same as the mean DEA concentrations calculated from size distribution measurements obtained using the MOUDI. This was 0.78 mg/m3, with a standard deviation of 0.22 mg/m3.

Jan 1, 1992

By C. J. Webb

Column experiments were performed to determine the ability of Madison aquifer material to adsorb arsenic and other constituents in a heavily mined watershed of the Black Hills in western South Dakota. The Madison aquifer is a karstic dolomite and limestone reservoir that is exposed on the flanks of the Black Hills. Streams that cross the Madison in the eastern Black Hills typically lose all their flow to the aquifer at sinkholes. In this work, Madison aquifer material was obtained for loading of six-inch diameter columns, which were backfilled with Madison aquifer water. Stream water then was allowed to displace the aquifer water during column testing in order to simulate mixing of the two waters at sinkholes and to determined adsorption of arsenic and other dissolved constituents. Initial results show that some adsorption of arsenic occurred when compared to more conservative tracers such as nitrate.

Jan 1, 1999

By R. Timothy McCrum

It is not often that the Interior Department's Bureau of Land Management, Office of the Solicitor, and Interior Board of Land Appeals (IBLA) jointly act to do the mining industry a good turn. That happened, however, in the recent decision, Schlosser v. Pierce, 92 IBLA 109 (1986). It was ruled that the theorized "independent mine requirement" was not a part of the discovery standard for mining claim validity under the 1872 Mining Law. The phrase "independent mine requirement" would require the mineral reserves within the boundaries of each individual mining claim in a claim group to support an independent economic mining operation. This, without regard to the fact that reserves on adjacent claims could be mined in a group operation, resulting in greatly reduced mining and marketing costs per ton of ore. Such a rule would render invalid under the 1872 Mining Law mining claims for almost all low-grade, high-tonnage, commercially valuable mineral deposits. Until the Schlosser case, no judicial or administrative decision had addressed the issue of the independent mine requirement. Nonetheless, a series of Interior decisions in the 1970s implied such a rule. Even the Rocky Mountain Mineral Law Foundation reluctantly concluded in 1984 that, "each claim to be valid must be capable of being operated at a profit independently of any other claim," the Foundation did criticize the rule as "excessively stringent" and urged the adoption of a "more realistic rule." Fortunately, the IBLA did just that in Schlosser when it ruled that the "application of a requirement that each claim be independently capable of being mined and marketed at a profit is rejected," 92 IBLA at 134. Now, a year after the 1986 decision, there remains some unwarranted confusion over the issue resolved in, and the meaning of, the Schlosser decision. There is no need for this confusion. The Interior Department's decision is a clear and legally sound victory for the mining industry. Grazing rights vs. mining claims The Schlosser case was an unusual mining claim contest. A private contest was brought before the Interior Department. Schlosser was a federal grazing lessee and surface owner. He filed against Pierce, the locator of a group of mining claims for bentonite over the same lands. At a hearing before interior's administrative law judge (alj), Schlosser's counsel began laying the foundation for the independent mine requirement. Counsel asked the mining claimant's expert whether he would go through the permitting procedure, road building, and reclamation just to mine the amount of bentonite on any one claim in the group. Schlosser's counsel proceeded with this line of questioning for each of the 16 claims in the group. Pierce's mining expert testified that the tonnages of mineral on any individual claim could not be mined, processed, and marketed at a profit if only one individual claim was considered. However, if the claims in the group were operated jointly, all the mineral could be mined, processed, and marketed at a profit.

Jan 8, 1987

By Peter J. Rafferty, Friedrich Steinhäusler

INTRODUCTION The Nuclear Energy Agency plays an active role in promoting international cooperation among its member countries in the field of nuclear energy. In addition to various other functions, it plays a major role in encouraging harmonisation of government regulatory policies and practices, promoting exchange of information, and coordination of research and development in the field of radiological health and safety associated with nuclear fuel cycle activities. The work of the Agency is carried out through a number of specialised standing committees. In particular, the Committee on Radiation Protection and Public Health (CRPPH) is responsible for the Agency's activities concerned with radiological protection and related environmental problems. Its functions include review and discussion of national radiation protection policies and practices, review of developments in radiological protection, interpretation of ICRP recommendations and the study of the means of their translation into practical applications, including the establishment of radiological protection standards. Its functions also include the preparation of technical studies and reviews on specific problems requiring attention, and coordination of further research and development at the international level. Major attention is presently being given to the NEA programme of work on problems associated with radiation protection and environmental impact of nuclear fuel cycle activities, with particular attention to the front-end (uranium mining and milling) and the back-end (waste management) of the fuel cycle. In this context, the increasing attention that has been given in several countries to the problems associated with the exposure of man to radon, thoron and their daughters, and with their dosimetry and measurement, were readily appreciated by NEA, which began an active programme of work in this field in 1976. Because of the detrimental health effects,as demonstrated by epidemiological studies,caused by prolonged exposure to excessive levels of shortlived daughters of radon, particularly in poorly ventilated underground mines, the bulk of attention and needed effort has been focussed on radon and radon daughters in uranium mining. In certain countries some concern has been expressed also about the significant levels of exposure experienced by workers in non-uranium mines,and members of the public who live in particular areas or in dwellings built with particular materials which produce higher than average levels of radon and radon daughters. However, at the time of the first NEA involvement in this field it was considered that one of the most urgent problems to be solved was that of ensuring adequate personal dosimetry for uranium miners. Consequently, the NEA was urged to organise a specialist meeting on personal dosimetry and area monitoring for radon and radon daughters to provide an international forum for exchanging information and reviewing problems in this field. The meeting was held in Elliot Lake, Canada, in October 1976. A second specialist meeting, on the same subject, was held in Paris, in November 1978, to review further developments in this area. These meetings demonstrated that the overall problem associated with exposure to radon and radon daughters had many facets, each of which in recent years has been the subject of considerable attention in many countries for different reasons. It emerged from the meetings that further work was required on a variety of issues in two main areas: 1) dosimetry 2) metrology and monitoring. Conclusions and recommendations which emerged from the two NEA specialist meetings were discussed by the CRPPH. As a consequence the Committee approved, in September 1979, a detailed programme of work in the area of dosimetry and monitoring of radon, thoron and their daughters, and approved the setting up of an international Group of Experts on Radon Dosimetry and Monitoring to undertake the work. The work has been divided into two phases - phase I, on dosimetric aspects, and phase II, on metrology and monitoring aspects. The terms of reference of the work are given in Appendix 1. A list of national representatives on the Group of Experts on Radon Dosimetry and Monitoring is given in Appendix 2; many of these persons are participating here in the Conference. The Group of Experts met for the first time in April 1980 and met again in September 1981. The work of phase I is nearing completion and a report is expected to be submitted soon to the CRPPH for its consideration. Three technical papers providing interim information on the study appear elsewhere in the proceedings of this Conference. The papers cover the three principle areas examined in the study so far: 1) dosimetric aspects 2) review of the "working level" 3) review of objectives and requirements for measurement and monitoring of radon, thoron and daughters. The authors, their affiliations, and the titles of these papers are listed in Appendix 3. A brief overview follows giving the principal results.

Jan 1, 1981

By M. R. Paterson

Operational and economic factors tend to favor the application of continuous processes in preference to batch processes in the majority of mineral processing applications. In this paper, the requirement for a Continuous Elution Process is established and the development of a large-scale pilot plant is described. It is based on a Continuous Elution apparatus which has a number of special features. These are the subject of various patent applications. The apparatus is suitable for adaptation to the AARL, Zadra and other methods of elution. The performance of the Atmospheric Continuous Elution System (ACE) is concluded to have significant advantages over the batch process.

Jan 1, 1986

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By M. Karmis

The algorithms for the mathematical modeling to predict productivity of underground room-and-pillar mining systems are well known and documented. These algorithms consider the time-varying relationships between mining equipment for a given geometty of operations as well as other constraints. This paper presents WebConSim, a newly developed, Web-based, user-friendly computer simulation tool for the Windows environment. The simulator is easily customized for use by field engineers, planning engineers and management and will help mine operators plan the optimum mining sequence for different mine geometrie and equipment layouts. Program output includes production data and equipment utilization indices, as well as time driven event chains that can be used for virtual reality presentations.

Jan 1, 2002

By M. G. Gorman

The effects of particle size, particle shape and degree of aggregation on n-dodecylamine flotation of -20 µ m quartz and glass beads were studied. A microflotation cell suitable for use with very fine particles was used in the flotation experimentation. Degree of agglomeration of particles was estimated by comparing size distributions of the quartz and beads in a dispersed system to size distributions in systems in which the collector was present. Results obtained show that in such flotation systems the fine particles are not generally present as individual particles during flotation but rather as aggregations. However, the quartz floats more readily than the beads indicating that there are some shape effects in the systems, considering the large specific surface areas of the materials it also appears that substantial amine consumption in the systems may limit flotation or lead to a requirement of greater reagent dosage for equivalent flotation to that tor larger particles.

Jan 1, 1990

By A. C. Smith

Although it is only utilized in a few U.S. longwall mines, a bleederless ventilation system can be an effective spontaneous combustion control method in mines with a demonstrated history of spontaneous combustion. In order to provide insights for the optimization of bleederless ventilation systems for U. S. underground coal mines, a computational fluid dynamics (CFD) study was conducted to model the spontaneous heating in longwall gob areas with a bleederless ventilation system. A single longwall panel with a bleederless ventilation system was simulated, and typical ventilation data for a longwall mine was used in the simulations. The permeability and porosity profiles for the longwall gob were estimated using a geotechnical model and were used as inputs for the CFD modeling. The effects of gob permeability and resistance of the collapsed entries on the spontaneous heating were studied. The effectiveness of nitrogen injection to prevent spontaneous heating in the gob was also examined. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of the National Institute for Occupational Safety and Health.

Jan 1, 2008

By Gerry Norton

The overall economics of a coal production unit are governed by a number of interdependent parameters which individually and collectively affect the final result. A business judgement is measured through return on investment considerations and in this respect, a sensitivity analysis proves useful. Figures 1 and 2 show how the various parameters affect return on investment through unfavourable change from planned or expected values. In the case of metallurgical coals, the selling price negotiated is the primary and most sensitive variable, followed by transportation, clean coal yield and capital funds invested in mining. It is clear that coal preparation plant capital and operating costs are the least sensitive to R.O.I. For energy coals, the selling price is determined by heat energy content (x cents per million BTU's) and is therefore not con¬sidered as an interdependent variable. In this case, transportation cost is the major primary factor affecting the level of income. Again the amount of saleable clean coal or total yield has a major effect on return on investment. Most of the factors considered in Figures 1 and 2 are fixed and unfavorable change from planned values becomes critical but clean coal yield is to a great extent controllable during actual operations. Once the theoretical yield for a particular coal has been determined, the optimum return is achieved by approaching this theoretical yield as nearly as possible in practical commercial operation.

Jan 1, 1975