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By Chinmay Shingote, Robert Visintainer, Renjie Ke, Jaikrishnan R. Kadambi, John Furlan

Hydrocyclone separators are widely used in various industrial applications in the oil and mining industries to sort, classify, and separate solid particles or liquid droplets within liquid suspensions, which are considered to be multiphase systems. Numerous valuable studies have been conducted in recent years to investigate the flow fields inside hydrocyclones. However, much of the information regarding the performance of cyclones in the literature has limitations and much of it cannot be considered as completely applicable to most real-world applications; many of the studies investigated the flow fields within extremely simplified hydraulic designs that are not representative of the complex geometries or large sizes which are typical in industry. Therefore, in this study, the two-phase flow system inside the actual hydraulic geometry of a milling circuit hydrocyclone is explored with the aid of both computational and experimental techniques (particle image velocimetry (PIV)). More specifically, the flow field with an air core has been investigated. In addition, the air-liquid two phases flow in a hydrocyclone might cause some challenges on both the computational and experimental sides. Two turbulence models are utilized in the numerical calculations: the Reynolds stress model and large eddy simulation. The computational studies are mainly focused on the local flow behavior and the prediction of the dimensions and shape of the air core. Different section planes of hydrocyclone are selected as planes of interest and divided into several fields of view (FOV) for PIV measurements. Two-dimensional experimental velocity vector maps are obtained in each of the fields of view. The computational results are validated globally using pressure and flow rate readings at the boundaries and locally by comparison to the PIV velocity vector maps and profiles.

By Daniel H. Stevenson

Safety and health are fundamental values at Barrick Goldstrike. An effective safety and health program protects employees, controls costs and increases productivity. Barrick Goldstrike believes that all occupational injuries and illnesses can be prevented and that safety must be the responsibility of every employee on the property. Safety and health must be thoroughly integrated with the company's efforts to maximize production at the lowest reasonable cost. The goals of safety and production are not mutually exclusive. They must work together to achieve safe production. Barrick Goldstrike began to mine the Betze deposit in 1987. From 1987 to 1992, Barrick employed a traditional safety system using safety specialists on each crew. These specialists were responsible for ensuring that the work areas were safe and that employees were properly protected and trained in safe work procedures. As the work force increased, it became apparent that this system was not the most effective method of ensuring safe production. In 1992, Barrick Goldstrike implemented a loss control program. It was modeled after the International Loss Control Institutes's (ILCI) International Mine Safety Rating System. This program was based on the philosophy that safety is a management responsibility. It can be managed in the same manner that production and cost issues are managed. Under the loss control system, safety needs are identified and evaluated. Programs with performance standards are developed and implemented and the activities and results are regularly monitored. Between 1992 and 1996, Barrick Goldstrike worked at developing and improving its safety management systems. These efforts had a positive return. The incident rates began to decline as these improved systems were implemented. Studies were undertaken in early 1996 to evaluate the effectiveness of the safety and loss control programs in reducing accidents. The studies showed a strong correlation between the implementation of the loss control programs and the resulting reduction in all accidents. Studies in the process division revealed that Barrick had experienced reductions in personal injury accidents from the projected number that could have been expected had no changes been instituted. These reductions ranged from 39% in 1993 to 59% in 1995. These reductions from the projected rates held true for all types of accidents; personal injury, property damage and Mine Safety and Health Administration (MSHA) reportable.

Jan 1, 1997

By K. H. Boss

In modern iron pelletizing plants fuel consumption has been optimized to such an extent that only minor savings can be expected in the future. Inspite of this fact, different process parameters were studied with a mathematical model and firing tests. The influence on fuel consumption and productivity was established. Furthermore, the adoption of simplified firing patterns was investigated, in particular for coal firing. In view of high costs of oil and natural gas, only substitution of these fuels by coal can substantially reduce production costs. In existing plants partial replacement of oil and gas can be achieved by addition of coal to the iron ore feed and by firing the pellets with pulverized coal. Two systems have been developed for the Lurgi- Dravo travelling grate process which enables coal to completely replace oil or gas. This paper describes in the first part the potentials for fuel savings available in the Lurgi-Dravo process. In the second part, different techniques for the use of coal as fuel in iron or pelletizing plants and in particular the recently developed Lurgi Multistage Coal Injection and Combustion Process (LMC) are discussed. Combined with the LMC process, a wide variety of iron ores and coals - also with high ash contents - can be used in the Lurgi-Dravo travelling grate process.

Jan 1, 1983

By Jerzy Iskra, Wlodzimierz Sikora

INTRODUCTION The effect of stress concentration in coal seams leads to a sudden release in the form of rock bursts. This effect is controlled along many lines, but the most efficient methods consist in preventing the accumulation of energy in the seam or in the roof. The over- burden weight resulting from mining is the main cause of excessive stresses in a seam. An additional weighting of roof takes place due to formation of abutment loads. The accumulation of energy in the roof takes place where thick sandstone beds occur in overburden strata. An essential problem is to watch, with appropriate measurements, the stress of coal and roof rocks, and to conclude, on this basis, the degree of rock burst hazard. The methods for determining the stresses in a coal seam are quite well developed at present. Let me mention here the seismoacoustic methods, or borehole overcoring measurement methods. On the other hand, an examination of the state of stress in the roof rocks is almost impossible at the actual stage of development of the measuring technique. Hence, the model of stresses in roof was analyzed and determined theoretically. Based on this model, a technology for roof fracturing was developed which aims at destressing rock masses in a planned way. The essentials of the fracturing method consist of a controlled breaking of sandstone roof beds in order to: - prevent the excessive weight on the pillar - prevent a spontaneous breaking of roof beds at a limiting state followed by sudden energy release. High pressure fluid infusion is used to release buildup of stresses. THEORETICAL AND RESEARCH BASES According to the tests carried out by A. Kidybinski by the method of finite elements, isobars of stresses occurring in the roof over the working face are shown in [Fig. 1]. The results indicate: - the highest values of compressive stresses occur in the roof directly over the seam, at a distance of about 5 m ahead of the working face - at a distance of about 22 m above the seam there is a line of transition of compressive stresses into tensile - within the tested roof area, the highest tensile stresses occur at a vertical distance of 35 m, and at a horizontal distance of 20 m from the line of the working face. H. Gil (1) determined the pres- sure initiating the fracture of roof rocks by the formula: [ ] On the basis of results of analysis of isobars, and of the initiation of failure stress formula (1), isobars may be designed for a given roof class. Assuming for sandstones the Poisson ratio equal 0.3, a field of failure stresses presented in [Fig. 2] is obtained. In this figure it can be seen in which areas the effects of sandstone breakage may be obtained with the operation of a determined hydraulic pressure. To use formula (I), the knowledge of Rr is still necessary. This value may be determined in two ways: - taking a sample and testing it in the laboratory, if there is access to roof rocks in the existing openings, - drilling boreholes in the roof from surface and examination of the rock mass by using hydraulic breaking, if there is no access to roof rocks.

Jan 1, 1982

In this presentation, CiDRA?s patented technology platform and its applications will be described. CiDRA's non-invasive, passive sonar array-based flow meter technology provides the volumetric flow rate of single or multiphase fluids by measuring the speed at which naturally occurring structures such as turbulent eddies or density variations convect with the flow past an axial array of sensors. These sensors are incorporated in a band that is wrapped around the outside of the pipe, resulting in no process downtimes for installation and unprecedented reliability. This technology has resulted in a unique ability to measure the flow rate of most fluids ? clean liquids, high solids content slurries, pastes, and liquids and slurries with entrained air. Unique and difficult minerals processing flow measurement problems and their solutions will be described. Examples of these problems include flow measurements in froth lines and flotation feed lines with entrained air, slurry lines with magnetite and other magnetic ore, slurry lines with abrasive or corrosive materials, high pressure lines, and lines exhibiting scale buildup. Recent technology developments will be covered.

Jan 1, 2008

By Robert A. Jankowski, George E. Niewiadomski

INTRODUCTION The utilization of longwall mining technology in the United States has steadily increased since the first longwall plow system was installed in 1952. At the errs of FY 1992, some 76 coal mines were employing longwall mining systems for coal extraction, and approximately 78 shearer and 3 plow faces are operating nationwide at any given time. Although the number of operating longwalls has remained relatively constant over the last five years, longwall production levels have increased significantly during this period. During the same timeframe, the number of continuous miner sections has decreased by 13 pct., from approximately 1438 in FY 1988 to less than 1250 in FY 1992. According to Mine Safety and Health Administration (MSHA) data, the average production of longwall sections increased from approximately 2400 tons per shift (tps) in FY 1988 to more than 3100 tps FY 1992. In fact, the average production has more than tripled since 1978, when it was only 850 tps. Approximately 25 pct. of operating longwalls are capable of producing in excess of 4000 tps. Longwall production currently accounts for nearly 36 percent of U.S. underground coal production, compared to 4 pct. in 1978. These improvements are attributed to improved longwall equipment and increased availability, larger longwall panels and automation. While longwalls are highly productive and offer other advantages, operations employing this method of mining continue to experience dust compliance problems more frequently than other mining system. Although dust levels have remained relatively unchanged over the past five years, averaging 1.6 mg/m3 and 1.9 mg/m3 based on results of operator and inspector sampling, respectively, during FY 1992, approximately 25 pct. of the operating longwalls were in noncompliance during any given bimonthly sampling cycle. Although modest gains in longwall dust control technology have been made, these have been overshadowed by significant increases in production rates. As more coal is mined, more dust is generated (Figure 1). Therefore, the increase in longwall productivity has meant that far more dust is being produced which must be controlled. [ ] In view of the current compliance record, any efforts to further increase longwall output, as has been forecasted, must be accompanied by: (1) the reduction of dust levels at existing operations through the proper installation and maintenance of suitable engineering dust controls; and (2) the advancement of dust control technology for today's and future longwalls that will enable all face personnel o work in an environment that is free of excessive levels of respirable coal mine dust. MSHA'S RESPIRABLE DUST CONTROL PROGRAM Since 1972, federal regulations have required each coal mine operator to continuously maintain the average concentration of respirable dust in the mine atmosphere during each shift in the active workings at or below 2.0 mg/m3 through the application of environmental dust control methods. The regulations provide for further reduction in the dust standard whenever the

Jan 1, 1993

By T. E. Kobrick

When Ralph Banks called and asked if I would agree to present a paper on environmental health in the coal industry, I was reluctant because I have not been actively involved in coal mine dust control on a day-to-day basis for a number of years. Ralph reminded me that I am one of the few industry people with an industrial health background who remembers the pre-1969 Federal Coal Mine Health and Safety era. This marks the second time within the past year that I have had a similar request based on the same basic reasons. While I was pleased to have been remembered, I'm beginning to feel like the football player who has played one too many games. In responding to Ralph's request to present an overview of the past, present and future of environmental health in the coal industry, I have attempted to take the broad approach of someone who, having been intimately involved in an area of endeavor for a number of years and then leaves, has an opportunity to revisit that area at a later date to see what has changed and what remains the same, what have been the successes and the failures. Since none of us can foresee the future, trying to project too far ahead is fraught with frivolity, except that unless one is foolish enough to accept a future invitation to review those predictions compared to actual events, no one will recall how accurate or inaccurate those projections might have been. I can assure you that I will not ask to accept that challenge, so you can rest comfortably that five years from today you will not be asked to sit through a similar presentation by me - and that's a promise.

Jan 1, 1985

By Harlan L. Kebel

Gypsum, the dihydrate form of calcium sulfate, has a history of usefulness to man dating back several thousand years, and a worldwide industry has been built on the mining and processing of this versatile industrial mineral. However, anhydrite, the other prin¬cipal calcium sulfate mineral, has very limited use, so much so that separate statistics for it are not always kept, but are usually included with those for gypsum. Although often found in close association, the two minerals have important chemical differences in that anhydrite is the anhy¬drous form of calcium sulfate (CaSO4) whereas gypsum is the dihydrate (CaSo4.2H20). These two molecules of combined water result in significant differences in hardness, density, and solubility of the two minerals. Of most importance, however, is the difference in thermal properties wherein gypsum, upon the application of a modest amount of heat, can be converted to the hemihydrate of calcium sulfate (CaSO4 '21H20) or plaster of paris, an intermediate product upon which is based over 90% of the value of all calcium sulfate end products. Anhydrite, on the other hand, is inert at these temperatures and its uses have been limited to only those few cases where calcium sulfate in the anhydrous form may have an advan¬tage. The earliest known use of gypsum plaster dates back some 5000 years to the time of the Pharaohs of Egypt who had discovered that gypsum rock, when exposed to fire, broke down into a powder which, when mixed with water, would form a putty that could be plastered on a rough mud brick or stone wall to make a smooth finish. It was used down through the centuries in a limited way for ornamental purposes, but did not achieve wide acceptance because of the quick (25 to 30 min) setting or hardening time which made it difficult to use. The first real understanding of gypsum chemistry was developed in France about 1755, and the gradual growth of its present utilization dates from that time. In the United States, the first utilization of gypsum was to grind it to a fine material for use as a soil conditioner at about the time of the Revolutionary War, and the term land plaster for ground gypsum was coined and has remained in use. Another cementitious material-lime or calcium oxide-had become the usual plastering material, but by about 1870 a method for retarding the set of gypsum plaster was discovered; today it has almost completely replaced lime as a plastering material. In recent years, gypsum products sold in the United States have reached a total valuation of approximately $2 Gpy as illustrated in [Table 1].

Jan 1, 1994

By B. Pesic, J. E. Wey

Introduction One important class of manganiferrous ores is manganiferrous-silver ores. They occur in the oxidized portions of the silver deposits of the western United States. Silver is often associated with manganese oxides, which in some ore deposits are dispersed through calcite, causing the black color of the host rock, known as black calcite. Black calcites can occur in carbonate rocks (0.3 to 50 kg Ag/t) and in volcanic rocks (less than 30 g Ag/t). Silver mineralization associated with manganese oxides and black calcites can occur in veins (Hewett, 1968) or as disseminations (Graybeal, 1981). In the US, silver-bearing manganese veins are located in a broad arc that extends from southern Colorado to southwest New Mexico, southern Arizona, and southeast Califor¬nia (Hewett, 1964). Areas of disseminated silver-manganese mineralization include the Creede and Silver Cliff districts of Colorado and the Tombstone and Hardshell districts of Arizona (Graybeal, 1981). These deposits are also characterized by significant amounts of silver halides - cerargyrite (AgCl), bromargyrite (AgBr), and iodargyrite (AgI). Black calcites are also common in these deposits. Oxidized silver ores containing the higher oxides of manganese are generally refractory to hydrometallurgical methods of treatment. In the past when these ores could not be smelted, they were treated generally by cyanidation at very low efficiency. The US Bureau of Mines (USBM) studied the problem of treating silver-manganese ores in the early 1920s (Clevenger and Caron, 1925). It was found that the refractory silver in the original ore was insoluble in all the common solvents for metallic silver and its salts. The conclusion was that to obtain the highest recovery of silver, all the manganese had to be dissolved or the higher oxides of manganese reduced to manganous oxide. More recent results from treating refractory silver¬manganese ores were given by USBM (Scheiner et al., 1973). The silver ores from Candeleria District, NV, and Round Mountain District, CO, were investigated. Electrooxidation experiments recovered only 34% silver with minor amounts of manganese dissolved. Sulfurous acid pretreatment followed by cyanidation revealed that silver extracted was proportional to the amount of SO2 used. A similar approach, using sulfurous acid pretreatment followed by cyanidation, was also proposed by Rhoades et al. (1984). Hazen Research Inc. performed extensive testing on silver and manganese dissolutions from Ag-Mn ores obtained from the Candeleria and Round Mountain ore deposits. The methods examined were direct cyanidation, pretreatment with sulfurous acid followed by cyanidation, malonitrile leach, and leaching with ammonium hydroxide-sodium thiosulfate under slight oxygen pressure. The results were not satisfactory. The importance of the vast reserves of Ag-Mn ores and the lack of an efficient method to recover these metals warranted a search for new technologies. One novel method for extraction of these metals is treatment of Ag-Mn ores with acidified thiourea solutions. Preliminary results of the leaching of silver and manganese from Ag-Mn ores were very encouraging (Pesic, 1984a) ; thus, it was decided to examine the new lixiviant in more detail. Experimental Ore from the Round Mountain District was studied. It was crushed, ground, and dry sieved into monosized fractions in the range of 16 to +400 mesh. The ore composition was about 90 g Ag/t and 3% Mn, depending slightly on size. The leaching experiments were performed in apparatus described elsewhere (Pesic, 1982 and 1984b). All experiments were performed at constant temperature conditions. The leaching residues were analyzed for silver and manganese. Atomic adsorption was used to analyze the solution samples and fire assaying to analyze the leaching residues (for silver). The standard experimental conditions are given in

Jan 1, 1987

By G. J. Hartman

The Fort Cady in situ borate mining project involves recovery of boron from a mineral deposit located over 1400 feet below the surface. The resource recovery method is being developed using in situ mining technology employing weak acid solution injected directly into the ore body. As the acid contacts the calcium based ore (primarily natural colemanite), the contained borate values are digested as boric acid. This boric acid solution is then pumped to the surface by use of submersible pumps. At the surface the borate values are recovered as a finely divided solid white powder after a series of chemical reactions and precipitations resulting in the production of calcium borate. The refined product is then sold under the trade name CADYCAL 100.TM All environmental permits for up to a production rate of 90,000 tons per year have been obtained. These permits required an Environmental Impact study (EIS) for federal lands and an Environmental Impact Report (EIR) for the private portion of the property. Regulatory permitting required over 5 years to obtain all the various authorizations for this project. The ore body has an average thickness of 118 feet extending over 273 acres. At a production rate of 90,000 tons per year with only a 70 percent resource recovery allowance, in place reserves are sufficient to sustain production for over one hundred years.

Jan 1, 1997

By S. Becker, M. Gruner

"Microstructure stabilized crushed rock salt is a newly developed backfill material for infrastructural openings in HAW repositories. The motivation for this research is based on the fact that the microstructure stabilized crushed rock salt is considered to be superior to the currently used crushed rock salt. The avoiding of settlements with a simultaneous 100% backfilling of openings, early stabilization of the geological barrier and low permeability (< 10-14 m²) are the main improvements. With the newly developed backfill material, the gap between the crushed rock salt and the saliniferous composite backfill materials is closed. Due to the minimized amount of a binding agent, the material will still be easy to handle like a bulk material, yet guarantees close to 100 % backfill success without any settlements. This is accomplished with a long-term stable re-mineralization that is formed by the components of the binding agent. INTRODUCTION The permanent storage of high active waste in underground geological formations (HAW-repository) is the labored method in Germany as well as internationally. Experiences have been made over the last decades, e.g. in the Waste Isolation Pilot Plant (USA) and the mines Morsleben and Asse (Germany). For this purpose salt formations are preferred as host rock due to the adequate heat conductivity, creep capability, and low permeability. In order to maintain the low permeability of the host rock and to avoid any contact of the stored waste with the biosphere, the formation of flow paths has to be prevented. The cavities which have been mined for the purpose of waste storage have to be backfilled with material that is alike to the host rock. Therefore crushed rock salt is considered as an appropriate backfill material. Crushed rock salt is a bulk material that is produced by mining saliniferous formations. This material is mainly used in underground repositories as a dry backfill that is brought into the cavities by the use of mechanical (loading machine, slinger machine or scraper) or pneumatic energy (pneumatic stowing). In Germany it is intended to use dry crushed rock salt with a maximal grain diameter < 64 mm for backfill in a future repository in saliniferous formations. With this method applied, the porosity of the backfill is up to 35 % [1]."

Jan 1, 2015

By Richmond Bennett

Airborne geochemistry became a reality with the advent of a high-sensitivity airborne gamma ray spectrometer with large volume NaI (TI) crystal detectors. Small variations in the concentration of potassium (as K40), thorium (as T1208) and uranium (as B1214) along the flight lines can be determined by measuring the gamma radiation with a sensitive airborne spectrometer. Small volume crystal spectrometers are useful only for detection of major surface uranium (Bi214) variations, but the large crystal volume system is an airborne geochemical exploration method capable of locating hydrothermal mineral deposits. Radioactivity as a guide to ore has been discussed in the literature by Damon (1950), Gross (1952) and others. The theoretical basis is that potassium, thorium, and uranium are concentrated in late stage solutions during magmatic crystallization. Thorium and uranium are concentrated in late stage liquids because their large atomic radius and high ionic charge prevent them from entering into the crystal lattice of common rock-forming minerals (Rankama and Sahama, 1950). Potassium, which has a large ionic radius and small charge, also can be concentrated in the high water content late stage fluids (Heier and Adams, 1964). A large proportion of uranium and thorium in rocks occurs around grain boundaries of the essential rock-forming minerals (Hurley, 1950). Uranium and thorium substitute for zirconium in zircon, a comparatively early forming mineral. However, most of the thorium and uranium is concentrated in residuals where they can be fractionated and incorporated into allanite, monazite, xenotime, and other minerals.

Jan 1, 1970

By Dale D. Teeters

Today, I would like to talk about ways to maximize the efficiency of surface coal operations. I doubt I will present any startling new ideas, but my comments may, hopefully, stimulate some conversation today and later back at your own operations. Strip mining is a very old technique used by ancient societies. Only with the coming of the machine age in the early twentieth century did we attempt to increase the magnitude of our operations. The industry, as we know it today, really only got started in the early 1900s with the advent of steam shovels. In the first portion of this presentation, I will attempt to show that we have, in many cases, not changed the basics of strip mining, only the scale. In doing so, quite often we have complicated the operations to the point of being unmanageable and unprofitable.

Jan 1, 1981

By Hasan Sevim

This paper summarizes recent developments in modeling of underground coal slurry haulage networks for determination of optimum design and operating parameters. The model presented here is based on a stochastic dynamic programming formulation, applicable in any typical room-and-pillar and/or longwall mining configuration commonly practiced in the United States. The formulation and the structure of the model is general enough to handle design problems involving more than one face and mining method. Given the stochastic operating characteristics and a system layout, the model can be used to determine the equipment configuration and operating regimes that minimize the overall unit cost of coal haulage.

Jan 1, 1983

The National Institute for Occupational Safety and Health (NIOSH) conducratory and field tests to evaluate the effectiveness of dust collector bags for reducing dust liberation and operator exposure from a roof bolter?s dust collection system. The laboratory tests evaluated the bag?s effectiveness to contain dust and the effect on canister filter loading in both a bag and bagless condition. The dust emissions from the collector?s exhaust were also measured in each condition during laboratory testing and show that nearly 100% of the test dust fed into the collector was captured by the dust bags. Loading and pressure drop on the dust collector?s canister filter is greatly decreased when using the bags, enabling longer periods of drilling without filter removal/cleaning. The field results showed that, even though dust standards were met, respirable dust in exhausted emissions was reduced around the bolter. Laboratory and field results show that benefits from use of the bags are realized in all areas of operator exposure.

Jan 1, 2008

By Bobby P. Faulkner

Allis-Chalmers has combined a new heat transfer device, called the ROLLER GRATE, with a unique processing scheme, resulting in the ROLLER GRATE Oil Shale Processing System of Allis-Chalmers. In order to test both the process and the equipment, a batch Process Development Unit (PDU) ROLLER GRATE retort was designed and built. This unit vas designed with the flexibility to process shale over a wide range of test conditions in order to demonstrate the attributes of the oil-shale processing system. Instrumentation on the PDU and extensive sampling of the effluent streams permitted the PDU test data to be used in the design and scale-up of the process. Over 30 tests have been conducted with Colorado shale, and over 25 tests have been run with an Indiana shale. In addition, rests have been run with California and Utah tar sands.

Jan 1, 1983

By Robert Stefanko

The most perplexing problem that has confronted mining engineers since the dawn of mining has been strata control. For the purposes of this paper, strata control will be defined as control of ground in the immediate area surrounding the, mine opening. To prevent movement of the surrounding strata into underground mine openings, miners have traditionally relied upon some type of artificial supports placed within the openings and located in a manner determined by trial and error from previous experience. Without the means to measure changes of stress in the supports and within the mine structure itself, an evaluation of the efficiency of the supports and mining system was impossible. Strata control thus remained an art with very little application of science. During recent years a revolution has occurred in mining. The modern mining engineer is contented no longer-with trial and error methods but is utilizing all the technological information at his command to obtain a solution to this age old problem. However, serious problems have arisen due to the lack of certain basic knowledge in three important areas: 1, The original stress distribution within the earth strata prior to mining. 2, Redistribution and concentration of the original stresses with mining activity. 3. The properties of rock that are significant in failure and its prevention.

Jan 1, 1960

By Ronald P. Geitgey

Pumice and volcanic cinder are volcanic rocks characterized by a cellular structure. They are formed as gases (primarily water) dissolved in molten rock, or magma, exsolve to generate a froth that cools and solidifies into rigid foam. The cells or bubbles are referred to as vesicles and they range in size from a few thousandth of a millimeter to several centimeters. Due to their vesicular character, both pumice and cinder have lower density and higher porosity than most other rock types, and, as the vesicle walls are broken, sharp cutting edges are continually generated. These properties are the basis for their commercial value as lightweight aggregates, insulators, absorbents, and abrasives. There is a certain imprecision in the terminology applied to volcanic products, reflecting to varying degrees the diversity of volcanic rocks, the relative youth of volcanology as a distinct field of study, and customs and practices of commercial trade. In general terms, pumice is light colored (white, gray, pink, pale yellow, or brown) and highly vesicular, usually with vesicle walls that are visibly glassy. Many pumice fragments are light enough to float on water. Pumice is readily cut by steel tools and some can be crushed by hand. Individual pumice fragments may range from 1 or 2 m in one dimension to c1 mm. Commercially, finer fragments are called pumicite or volcanic ash. In volcanology the term ash refers to any fragment c2 mm in size regardless of its composition. Cinder, or scoria, is dark colored (black, red, brown) with thick vesicle walls that appear dull or stony and that may exhibit iridescence. Cinder typically is heavier than pumice and has a higher crushing strength. In volcanology the term scoria refers to dark vesicular material throughout a wide size range while it is common commercial practice to use scoria for larger fragments (usually greater than 2.5 cm) and cinder for smaller fragments. GEOLOGY Composition and Physical Characteristics Most pumice is silicic ranging from 60 to 70% SiO2, that is, dacitic to rhyolitic in composition. Pumice of less silicic composition, including basaltic, occurs but it is less common and does not have extensive commercial use. Scoria and cinder are typically basaltic to andesitic with approximately 50 to 60% SiO2. Pumice and cinder may contain phenocrysts of feldspars and various ferromagnesian minerals that crystallized in the magma prior to eruption. Fragments of rock through which the magma has passed may be entrained in the melt and wall rock may be fragmented and admixed during an explosive eruption. Some scoria particles have lithic fragment cores. Phenocrysts and lithic fragments are relatively unimportant in cinder end uses but their presence may be detrimental in certain pumice products. Pumice typically has vesicles <1 mm in size separated by thin walls. Vesicle shapes include irregular, spherical, elliptical, and elongate to the point of being tubular with a silky appearance (Fig. 1). In some pumices, the vesicles are interconnected making the fragments permeable and highly absorbent; in others the vesicles are isolated forming a highly porous but very impermeable pumice. Floating masses of pumice fragments from the eruption of Krakatoa in 1883 were reported in the Indian Ocean for up to two years after the eruption. Vesicles in cinder are larger and range from spherical to highly irregular in shape with much thicker walls and abundant interconnections. Scoria and cinder are typically heavier, porous, and do not float. Density and hardness of pumice must be expressed precisely to avoid confusion. Density may refer to the glass itself, the apparent density of the vesicular pumice particle, or the bulk density of pumice in a deposit or in a product. Pumice glass has a specific gravity of 2.5 or more depending on its composition. Pumice fragments typically have a specific gravity of 4.0 and so will float on water, at least until enough vesicles are saturated to cause them to sink. Expressed as density, pumice fragments typically weigh less than 1 g/cm3. Depending on moisture content, particle density, and particle size distribution, the bulk density of pumice typically ranges from 500 to 700 kg/m3. Typical bulk densities of cinder range from 700 to 900 kg/m3 for material used as lightweight aggregate while the bulk density of cinder used for highways and railroad ballast may be considerably higher. Hardness in pumice may refer to the glass vesicle walls, possibly including crystals or lithic fragments, or to the apparent hardness of the particle as a whole, which is more a measure of the strength of the vesicle structure. Pumice glass typically has a Mohs hardness of 5 to 5.5. A pumice particle may have a much lower apparent hardness and be easily cut with a knife or steel saw because the vesicle walls break readily. Pumice and cinder are primarily pyroclastic deposits formed as fragmental products of volcanic eruptions. Several classification - systems differing in various details have been proposed for pyroclasts and pyroclastic deposits based on particle size (Fisher, 1961, Schmid, 1981, Cas and Wright, 1988). Generally, particles >64 mm are referred to as blocks or bombs depending on their shape. Particles between 64 mm and 2 mm (4 mm in some schemes) are called lapilli and smaller particles are called ash. These names may be modified by compositional terms, for example pumice lapilli or scoria lapilli. Most commercial deposits are composed of fragments in the lapilli size range. In the United States, pumice fragments having one dimension of 5 cm or more are legally defined as block pumice. The procedures for acquiring pumice from federal land are determined in part by this definition and are described in a later section.

Jan 1, 1994

By R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981

By J. Harrison Daniel

More stringent dust standards, because of quartz in coal, have become a major compliance problem for the coal industry. Reducing quartz levels has become the major focus of the Bureau of Mines dust control research program. Results to date have identified hath the sources of quartz and prospective control technology. Control techniques are being developed for both room-and-pillar and longwall mining underground operations. Techniques include changes in face ventilation, mining cycle modifications, and better roof bolting dust control.

Jan 1, 1985