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

The literature on four different topics is re¬viewed herein for purposes of mathematical mod¬eling of mine water inflow. The first of the topics reviewed is modeling of fissured or fractured ma¬terials; the second is general Darcy based flow models; the third is field methods for determining fracture flow properties; and the fourth is methods used presently for determining mine water inflows. The four topics are considered in the foregoing order. FRACTURE FLOW MODEL Two basic types of fracture flow models have been developed. The first type uses various tech¬niques for space averaging to obtain the equations. Systems are considered in which the porous media properties vary smoothly enough and the fracturing is sufficiently profuse and well distributed that the locally averaged properties are meaningful when viewed on the macroscopic scale. The final result is that a coefficient is used which is similar to saturated hydraulic conductivity and in many cases, simply is hydraulic conductivity. In other cases, the coefficient is apparent hydraulic conductivity be¬cause the response from flow in the fractures is more rapid than flow in the pore spaces. Faust and Mercer (1980) refers to this method as the contin¬uum approach. Some of the papers on this type of problem are by Duguid (1973), Verner et al. (1974), Owili-Eger (1975), Duguid and Abel (1974), duPrey and Weill (1974), and O'Neill (1978). Duguid and Lee (1977) also considered heat flow in the fluid and the porous medium. These papers all result in various types of computer programs. Some papers on this topic that do not result in computer programs are by Snow (1968, 1969) and Streltsova (1976). Some of the foregoing articles consider con¬solidation of the porous media and the resulting effect on hydraulic conductivity. Barenblatt et al. (1960) developed an analytical solution for fracture flow and discussed the time that is involved in the drainage process and the effect that fracture flow has on the drainage time. The second type of fracture flow model con¬siders the width of the fractures as a function of the pressure in the field. Faust and Mercer (1980) refers to this method as the discontinuum approach. This approach often requires two finite element programs that consider the stress in the material and the flow in the fractures separately. The pro¬grams are then linked to arrive at the final solution. Some of the articles that consider the interrela¬tionship between pressure and the size of the frac¬ture but do not use computer solutions are Gringarten et al. (1975), Narasimhan and Palen (1979), and Witherspoon et al. (1979). Papers that use a linked computer program between stress and flow are by Ayatollahi (1978), Gale et al. (1974), and Noorishad et al. (1971). Ayatollahi's program is primarily for the petroleum industry because it does not consider the effect of gravity. Ayatollahi assumes that pressure effects are very large in com¬parison to gravity effects. This assumption would apply only to very deep aquifers. Articles that consider fracture flow on a more applied basis are by Wittke et al. (1972), who con¬siders that the flow is entirely through fissures and negligible through the pores of the rock. Gringarten and Witherspoon (1972) developed type curves that may be used for analysis of fracture flow systems; the curves are based on the assumption that it is possible to distinguish between aquifers with hor¬izontal and vertical fractures and that it is possible to analyze the system as an equivalent anisotropic homogeneous porous medium with a single fracture with much higher permeability. DARCY FLOW MODELS The Darcy flow models of greatest interest to mine water inflow prediction are those which allow

Jan 1, 1986

By A. P. Grobbelaar

Andalusite is a member of the sillimanite group of minerals consisting of andalusite, kyanite, and sillimanite. These minerals are the structural polymorphs of Al,Si05 (the anhydrous silicate of aluminium) and although chemically similar, they differ in physical properties due to their different crystal structures. For further information on the sillimanite group, see the chapter on Kyanite and Related Minerals that follows. The minerals are generally the products of the metamorphism of alumina-rich sediments and are found in schists, gneisses, or hornfelses, which have undergone either regional or thermal metamorphism from their original form. Sedimentary deposits may also form since the minerals are chemically and physically resistant to breakdown. When heated to a sufficiently high temperature the sillimanite group minerals convert to mullite with a resultant volume expansion and the liberation of silica. Due to the stability of the mullite formed, its high hot strength and resistance to chemical and physical erosion, they are widely used in the manufacture of refractories and special ceramics (Table 1). The reaction can be represented as follows: [ ] The worldwide use of these minerals depends to a large degree on their availability (Fig. 1). Andalusite, with its lower conversion temperature and the lowest volume expansion on being heated is the natural choice as a raw material to produce energy-saving refractories without prior calcination. The Republic of South Africa holds by far the largest minable andalusite deposits of the world, estimated at some 70 Mt of known in situ reserves. The only other large commercial production of andalusite comes from France, with known reserves of about 5 Mt. Potential new deposits are known from China, Ireland, Australia, and South America. World consumption of andalusite in special performance refractories has risen from less than 10 ktpy in 1960 to nearly 300 ktpy in 1990, which is due to cost effectiveness in many industrial applications compared to other conventional refractory products. Andalusite-based refractories have traditionally been used in Europe, South Africa, and the Far East; usage in America was limited in the past due to relatively high delivery costs and availability of other local refractory minerals. However, there is a significant trend towards greater use of andalusite in North America, thus opening new markets and replacing other traditional refractory raw materials. GEOLOGY Andalusite occurs in metamorphosed rocks of clay-like com- position, as in the andalusite-hornfelses in thermal aureoles (contact metamorphism), formed under conditions of high temperatures and low stress, and in regional metamorphic rocks, such as the andalusite-schists. These schists are mostly unconnected with definite igneous intrusions but were presumably also formed at high temperatures and under low stress. Andalusite can also appear in alluvial deposits associated with the primary deposits. At depths of 10 to 20 km, temperatures of up to 450°C might be expected. Another result of increasing depth is an increase of pressure due to the load of the rock cover. The load stresses involved in metamorphism is generally believed to lie in the range 0 to 10 000 bars. During contact metamorphism the increase in temperature is brought about by the injection of magma. The temperature reached in the country rock at some distance from the igneous body will always be less than the initial temperature of the intrusion, which is presumably less than 1 200°C, and will vary with time and distance from the contact. A simple metamorphic reaction involving dehydration of an aluminous sediment can be: [ ] At higher temperatures sillimanite would form at the expense of andalusite, while kyanite would be the mineral formed at higher pressures, as can be seen in the phase diagram of A12Si05 (Fig. 2).

Jan 1, 1994

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By Richard R. Klimpel

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

Jan 1, 1989

By Agne Rustan

INTRODUCTION This work is a continuation of the preparatory study by Rustan and Stöckel 1979 (3) and (4) of the possibilities to introduce automatic mine ventilation in Swedish mines and construction works. The preparatory study has been financed by the Swedish Board of Technical Development. In the study it was recommended that an automatic step by step control of the fans near the working place due to diesel exhaust emission would be the best solution. A more advanced alternative would be to regulate the fresh air continuously in proportion to the contamination. The latter alternative has been studied by Boliden Mineral Co in the Udden mine. In this study it was found that the air supply could be regulated continuously but it was very important where the sensors feeling the contamination were placed. An intense study of the environment underground is now being planned by our division together with specialists from the National Board of Health and Safety, The Swedish Work Environment Fund, LKAB and the local Governmental Inspection of the Workers Safety. Different gases, radon, dust, airspeed, moisture temperature and pressure is planned to be measured continuously and correlation between the parameters is going to be established to decide the best measurement parameters for regulating the system. The best situation of sensors is also going to be studied. Despite this work hasn't been done yet this paper covers some of the problems which could be encounted with continuous regulation. In this report the environment has been studied during the test with step by step regulation of fans by radio. Diesel vehicles are the largest polluters under ground in Swedish mines and we have therefore suggested that the large diesel vehicles should regulate the fans. The goal of this study has therefore been to find how active diesel vehicles should be able to start and stop gallery fans at both intake and return air. In this paper different methods for activating fans are discussed. Then the best methods are chosen and tested in the mine. In this study we have found that changes of temperature is a cheap and effective method to start and stop cross cut fans in sublevel caving and radio signals seem to be an attractive system for start and stop of gallery fans. This work has been done in close collaboration between LKAB in Kiruna and the University of Luleå - Division of Electronics and Division of Mining and Rock Excavation. A reference group for the project has been established with ventilation experts from the mines, representatives for the local mine worker's unions and authorities from the local Governmental Inspection of the Workers Safety. VENTILATION SYSTEM STUDIED In the project the ventilation system at LKAB in Kiruna has been studied. The mining method used is sublevel caving with 9-13 m slice height. The air is heated winter time over ground by hot water heat exchanger. The main fans are pushing air down to a main distribution level along the 4000 m long and about 100 m wide orebody. From the distribution level, air is taken through shafts down to the sublevels, where it is pushed through the levels by local gallery fans on both intake and return side, see fig 1. The air ( ~ 8 m3/s) is moving freely through the gallery head. The cross cuts are ventilated with smaller fans, diameter 500 mm, and flexible duct, see fig 2. The aim of this study has been to study how to start and stop gallery- and cross fans and examine the environment when the fans are stopped . In Sweden there are no regulations telling that a minimum of air amount must be supplied. Instead the hygienic limits must be fullfilled.

Jan 1, 1981

By Charles A. Beasley, Vince Joyce, David E. Beasley, Orlando A. Gallegos, David A. Shuman, Mehmet H. Erten

COAL CHARACARISTICS AND PREPREPARATION REQUIREMENTS Preparation and Market Specifications The national ambient air quality standards (NAAQS) established by the Environmental Protection Agency in 1977 require that when coal is burned, the environmental degradation should be a minimum. For this reason, even coal containing low ash and sulfur must be used in compliance with the NAAQS. The most important pollutants resulting from the combustion of coal are ash, SO2, and NOx gases, and the particulate matter escaping with the combustion gases. The ash is the residue left after complete combustion of coal. It results from the inorganic mineral matter existing in the coal before it is burned. The SO2 results from the combustion of inorganic and organic sulfur associated with the coal. The inorganic sulfur may occur as pyrite, marcasite, or as sulfate sulfur of mostly iron and calcium. The authorities disagree in the definition of organic sulfur in the coal. However, according to ASTM Standard D2492-77,1 the organic sulfur is that sulfur occurring in the coal after subtracting the sum of pyrite and sulfatic sulfurs from the total sulfur in the coal. The NOx gases are produced from high temperature oxidation of nitrogen in the coal that is released into the atmosphere. The particulate matter may consist of fly ash and unburned coal. Selective Mining The cleaning of coal from included ash and sulfur-bearing materials should start at the working faces of surface or underground mines. This can be done by (1) mining only those seams which meet certain quality standards, (2) not mining sections of the seams which may cause deterioration in overall quality of the mined coal, and (3) using a mining method that allows selective mining of coal areas or horizons either underground or on the surface. Since these steps improve the coal quality before the coal reaches the preparation plant, they can be considered to constitute precleaning of coal. During the exploration and development stage of a mine, core and channel samples taken from the coal seam are used to determine (1) petrographic analysis and physical properties,2 (2) character of roof and floor of the seam, 3,4 (3) character and quality of impurities,5 (4) proximate and ultimate analysis, and (5) washability characteristics. However, all of these parameters are subject to considerable variation even within the same seam during the advance of the coal face. This creates a problem in meeting the market demand for coal quality if the coal is to be shipped directly from the mine. If the mining operation is large enough, special blending silos or mechanical blending systems may have to be used to provide uniform quality to the market. However, even if the coal is shipped from a coal preparation plant, prepreparation

Jan 1, 1991

By Rudolf E. Greuer

Introduction The Institution of Mining and Metallurgy and the Institution of Mining Engineers organized the Third International Mine Ventilation Congress held in Harrogate, England. Sixty-one papers were presented and about 300 participants registered. About 30% of the participants came from Great Britain, 20% from South Africa, and about 5% each came from Australia, Canada, China, France, US, and West Germany. The remainder of the participants came from 18 other countries. Eleven of the papers presented dealt with methane, three with diesel exhausts, four with dust, four with radioactivity, 18 with heat, and 15 dealt with main and auxiliary ventilation. Six of the papers dealt with mine fires, which is a boundary region between main ventilation and gas concerns. Ventilation Congress proceedings are available. See ME, December 1984, page 1687, New Books page. With CH4 and radioactivity topics, the Australians dominated since they are becoming large coal and uranium producers. With diesel exhausts, the most important problem in highly mechanized mines, the North Americans were prominent. The South Africans, working in 90% SiO2 in their gold deposits, were preeminent in presentations on dust. They also led in topics related to heat, joined by the West Germans whose mines are getting deeper. In main and auxiliary ventilation, Great Britain and West Germany provided the majority of contributions. Methane The large number of methods for the precalculation of CH4 liberation in longwall mining all contain three elements: gas content of the coal, gas emitting zone (or influence zone of face), and degree of gas emission. Determination of the gas content, commonly accomplished by taking coal samples, does not pose much difficulty. But this is not the case with the other two factors. Most existing approaches rely on rock mechanics observations. Some only rely on intuition or speculation. The West German coal mines conducted a large research project between 1977 and 1982, in which gas pressures around longwall faces were measured. Gas pressures and gas contents can be related. Therefore, influence zone and degree of gas emission can be determined simultaneously. The precalculation of CH4 liberation in room-and-pillar mining is simpler. Since foot and hanging walls remain essentially intact, gas pressure distribution and gas flow can be calculated using hydraulics equations. A key to these calculations is permeability. A group of Australian researchers reported that the permeabilities of rock under three dimensional stress differs from rock under destressed conditions. This fact was known, but no systematic observations existed. These were provided. Methane drainage has been practiced for more than 40 years. It is used with great success in longwall operations. Some of the West Germans think that methane drainage from the footwall is neglected. Under certain geological conditions, they claim this is as important as drainage from the hanging wall. A research project was presented in support of this claim. Considerable efforts have existed for more than 10 years to use methane drainage in room-and-pillar mining. A paper described the accomplishments of Consolidation Coal Co. Since coal and rock are not fractured as much as in longwall mining, gas transport takes place by the slow diffusion from micropores into the cleats. Then, it is transported by laminar flow along the cleats until a drainage borehole or the mine workings are reached. Boreholes about 300 m (1000 ft) long can reduce the gas content of a band of coal 100-m (330-ft) wide by 50% in one year. A paper from India described methane drainage from gob areas. A French paper reported that especially high methane concentrations can be drained from gob areas if the face ventilation is descensional. Methane buoyancy and ventilating pressures compensate each other, and little air dilution by leakage takes place. Another French paper described a newly developed gas and air velocity monitoring system making use of a microcomputer. Four papers on methane dealt with gas outbursts. Two Australian contributions described the problems encountered with CH,-CO2 mixtures and the attempts to solve them with drainage boreholes having 10 to 20-kPa (1.5 to 3-psi) suction pressures. A Hungarian paper described problems and attempts to reduce the risks of gas outbursts in Hungary. Prediction of gas outbursts in coal seams was the focus of an English paper. Observations show that they are frequent when coal, through its geological past, has become soft and brittle with a resulting increased desorption rate. An instrument for gravimetric de-

Jan 1, 1985

By B. A. Hancock

It is not at all surprising that causticized potato and potato derived amylopectin starch solutions performed much better than their parent starches. Some preparation is required to rupture the starch granules to effect the polymeric adsorption and interparticle bridging necessary for selective flocculation. In laboratory work comparing the deslime performance of causticized and autoclave cooked laboratory corn starch solution preparations, it was found that higher deslime weight rejections, with attendant proportionally greater iron unit losses, occurred with the causticized starch. These results may be specific to the ore involved but they do suggest that cooking and causticizing cause different starch granule rup- ture and/or starch breakdown, which have an effect on desliming response. I calculated from the data in the article that the slimes product grades were high - 24.3% and 20.3% Fe when 53.7% and 54.5% Fe concentrate products were obtained, respectively, in Table 4, and 21% Fe with a 62.6% Fe concentrate in Table 5 - using the natural tailings sample, which had a head of 34.3% Fe. It may be advisable for the authors to consider different starch preparations in future investigations. The combination of upgrading and selectivity results presented in Table 4 are not as good as the authors suggest. The authors' claim that a system has been developed to produce saleable concentrates from the Kudremukh tailings is quite disconcerting. There are many hurdles yet to be crossed before commercial application of selective flocculation becomes possible because differ- ences between the very small-scale laboratory tests conducted and commercial application are rather large. Among the many differences are varying circuit feed grades that will occur from use of tailings, the apparent face that much lower tailings grades will be encountered in practice (it is much easier to achieve a high concentrate grade with reasonable recoveries using 34.3% Fe tailings as in the study rather than 25.3% Fe tailings grades that the plant apparently averages), the hydraulic nature of the thickeners used in operations compared to the static system used in laboratory tests, the different size distributions that will be obtained from a plant closed grinding- classification circuit, and differences in water used in a plant operation and the laboratory. The authors wrote that it was necessary to overgrind to be sure that the coarse gangue would not settle with the iron oxide floccules. This situation is likely to be exaggerated in commercial operations where it is assumed cyclones would be used for classification. Because cyclone classification is greatly influenced by particle densities, there will probably be an even greater difference in size between the iron and gangue particles in the plant, which would make the gangue slightly coarser still in relation to the iron. This would make the selective flocculation-desliming separations using the procedure employed by the authors even more difficult and, using the dispersant system the authors employed, greater overgrinding would be required. To grind finer to minimize the coarse gangue in the flocculated iron oxides is quite inefficient and appears not to broach the problem. The actual problem appears to be insufficient dispersion of the ground pulp. In this situation, addition of a dispersant would likely be required to attain a sufficiently high pulp dispersion level to efficiently effect a selective flocculation-desliming separation. Although the very coarse particles would still have a tendency to settle with the floccules, it probably would be found unnecessary to overgrind as much as indicated. Use of an optimum combination of dispersant and pH modifying reagents may also significantly improve the selectivity of desliming. Additionally, although it is possible that sufficient dispersion may be obtained by pH control alone in some situations, it is quite probable that added dispersity was obtained in the reported work from using distilled water. It is research experience that distilled water enhances dispersion. In commercial operations it may not be expected that sufficient dispersion will be obtained by pH control alone, unless the water used in the process is by nature quite dispersive. Overall, a change in the Kudremukh tailings dispersant scheme appears necessary where a dispersant is used in conjunction with a pH modifying reagent. With this change, different dispersion-flocculation responses will result that would have to be further evaluated. Therefore, it is still an open question whether an efficient and effective selective floccula- tion separation using Kudremukh tailings may be obtained that will produce saleable concentrates.

Jan 1, 1987

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By J. A. Herbst

Introduction The ability to evaluate the improvement resulting from the implementation of an automatic control system is essential to the Justification of its cost. This evaluation may occur after anew control system has been commissioned. In other cases an existing plant which is operated manually may be automated only In part in order to determine In ether the met of it should be. Finally. in some instances dynamic isolation my be used prior to installation of the control system to determine the strategy to be used or to evaluate the improvements In process performance expected from automatic control. This paper describes the development of a methodology for making these evaluations. In order to obtain an accurate assessment of Improvements In a short time, statistical methods can be employed. Mineral processing plant, are subjected to many types of both random and deterministic disturbances of various duration. The statistical method, available for process control at analysis are designed to counter the obscuring effects that random disturbances have on such as evaluation. The basis for evaluation and comparison is some freely chosen measures of process performance termed the performance index or objective function. The experimental data used to calculate the performance index should, if possible, be collected by alternately running a circuit Or plant with the two or more type, or control being tested ("serial testing”) or more preferably, by running two circuits with the two type, of control side by side ("parallel testing") . The performance index is a random variable so that when comparing two types of control, statistical hypothesis testing must be used. By testing s hypothesis statistically one accept, or reject, a (null) hypothesis [ ] concerning the mean or variance of one or more performance indices. For example, a hypothesis Ho might be that the mean of a performance index with stabilizing control is larger than with manual control. Defining a Performance Index The performance index, denoted by [ ] , is a measure by which alternative strategies or situations my be cam peered on either s technical or an economic basis (see Table 1). Technical measures are concerned with the effects of control on process variables or outputs or on the deviations of controlled variable, from their setpoints. For commotion circuit, important technical measures of performance are throughput, product fineness and the consumption of liners, media and electrical energy. For flotation circuits important technical .satires are recovery, grade and reagent consumption. Economic measure such a, cost and profit n usually determined from a combination of technical measures and are usually the final factor upon which a decision may be made, although a control system my achieve higher values of technical measure , the control effort expended my be of such magnitude that the economic measures, such as cost, are poor. Running Comparisons Comparison, between automatic and manual control are normally done with data obtained before and after control system Installation, unless ore characteristics remain the same during the period of data collection the results will be biased. By running the test circuit with the control system alternately turned on and curled off until all possible ore type, have been run through the circuit, unbiased result, can be obtained. This mode of testing is termed “serial testing.” An alternative and potentially more powerful and quicker made of testing is "parallel testing" in which two circuits, one with the type of control being tested and one without, arm run aide by side. In this type of testing both circuits process the same ore so that valid results occur sooner than with aerial testing. The data may be averaged before testing or the differences in the performance indices may first be calculated before averaging. However, the two circuits must be nearly identical in their mechanical characteristics otherwise differences other than the degree of control will bias the results. It mechanical differences exist than it my be necessary to adjust one or the circuits prior to testing. Sometimes it is possible to correct far mechanical differences based on Inherent efficiency calculations made with a model. Hypothesis Testing Once the data is collected, hypothesis fencing can be carried out to determine if the automatic control system has resulted in a statistically significant improvement. One can construct a histogram of the distribution or performance indices with am without control such as Figure 1, or the

Jan 1, 1991

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By H. H. Richards, Ray L. Tobie, L. A. Thomas

GENERAL DESCRIPTION Since the beginning of operations, with the exception of a small tonnage mined by slushing, ore extraction has been by full gravity caving. Formerly, a checkerboard sequence of block undercutting was followed with the even-numbered blocks in one panel and the odd¬numbered blocks in the adjacent panel being mined. As these blocks were depleted, the intermediate or pillar blocks were mined (Fig. 1). Following this checker¬board, the mining sequence went through a number of changes, finally evolving into diagonal retreat panel cav¬ing by blocks (Fig. 2). The numbers in Fig. 2 indicate the sequence in which blocks were undercut. Gaps in the numbering sequence indicate undercutting on the level outside the illustrated area. Geology The ore body is a low-grade deposit of chalcopyrite mineralization disseminated throughout structurally weak, highly fractured, strongly altered granitic host rocks. It takes the shape of a gently dipping elliptical cylinder consisting of an ore shell of variable thickness surrounding an interior waste core. Major and minor axes of the mineralized cylinder are 1524 m (5000 ft) and 762 m (2500 ft), respectively, and length approximates 2438 m (8000 ft). Ore is sufficiently fractured to break readily into medium-coarse size. The igneous rock complex containing the ore body is covered by a wedge-shaped blanket of Tertiary con¬glomerate which was brought into place by faulting along the major regional structure of the San Manuel fault. Thickness of the conglomerate cap varies from only 9 m (30 ft) at the east end of the ore body to more than 610 m (2000 ft) at the west. Structurally, the con¬glomerate is much more competent than the igneous host rocks and, when caving, it breaks into massive chunks. Conglomerate boulders seen in drawpoints underground are very coarse. The total rock column over the initial mining area of the 1415 grizzly level was 354 m (1160 ft) of which 122 m (400 ft) was ore, 79 m (260 ft) was leached igneous capping, and 152 m (500 ft) was conglomerate above the San Manuel fault. Diamond Drilling: From 4572 to 7620 m (15,000 to 25,000 ft) are drilled annually from underground workings to delineate the ore body. MINE DEVELOPMENT Haulage Level In the south or main ore body (see Figs. 3-6), with the exception of the draw and transfer raises, all the extraction openings are concreted (Seaney and Tobie, 1965). The haulage panel drifts, which are 18 m (60 ft) below the grizzly drifts, are first driven with pre¬concrete ground support. The drift, which has an arched section, then is concreted using mobile collapsible steel tunnel forms. The haulage drifts leading from the pan¬els to the hoisting shafts are not concreted. After the panel drifts have been concreted, the raise stations from which transfer raises will be driven are constructed and the raise-station ore-drawing chute is installed. The chute is prefabricated of A-36 steel with undercut guillo¬tine gates made of abrasion-resistant 2.5-cm (1-in.) steel plate powered by 20-cm (8-in.) air cylinder installed on each side of the raise station. Transfer Raises The transfer raises are lined with 15 x 20-cm (6 x 8-in.) cribbing and are 1.22 m (4 ft) in the clear. Each cribbing is protected from wear by a high carbon steel angle which is nailed onto the cribbing. The transfer raises are driven from each side of the raise station on an angle of 1.1 rad (63°). Each raise con¬sists of a main and a backover branch. The transfer¬raise driving crew consists of two men working one shift only. Grizzly Drifts After the transfer raise reaches the grizzly level, the grizzly drift can be driven. The grizzly drifts are spaced at 10.6-m (35-ft) centers and are driven parallel to the long axis of the ore body (see Fig. 2). This drift is driven by a two-man crew working on one or more drifts at a time using feed-leg machines. The eight grizzlies in the 42.7-m (140-ft) long drift are spaced at 5.3-m

Jan 1, 1982

By Thomas B. Borak, Keith J. Schiager, Janet A. Johnson

OBJECTIVES OF RADIATION MONITORING Monitoring is a tool used in the practice of radiation protection. The primary reasons for monitoring are to reduce radiation exposures to levels that are as low as reasonably achievable and to assure that no individual receives a dose exceeding the maximum individual dose limit. The documentation of radiation doses for legal, medical, or epidemiological reasons is a subordinate function of any monitoring program. The investment in radiation monitoring programs should be guided by four criteria: (1) the detection and avoidance of unnecessary exposures, (2) the magnitude of potential health risks, (3) the determination of combined doses and risks with adequate confidence, and (4) the verification of compliance with established limits. FIRST CRITERION: DOSE REDUCTION - DETECT AND CORRECT UNNECESSARY EXPOSURES The system of dose limitation advocated by the International Commission on Radiological Protection (ICRP, 1977), and subscribed to in a broad sense by various regulatory agencies, is comprised of three essential ingredients: (1) [justification] of any practice that produces radiation doses by some commensurate net benefit, (2) [optimization] of radiation control measures by reducing doses to levels that are as low as reasonably achievable, and (3) [limitation] of individual doses to preclude inequities and [moldistribution] of risks. All too often, only the third criterion - the limitation of individual doses to prescribed regulatory limits - is explicitly addressed in everyday radiation protection programs. The emphasis of most exposure control and monitoring efforts appears to be directed toward limiting and documenting individual doses that might approach the legal limit. The first two criteria, i.e. justification and optimization, should contribute to a rationale for allocating monitoring efforts. When applied to individuals, these criteria mean the detection and elimination of unnecessary exposures. This should be a high priority of any monitoring program. Measurements should be directed toward detecting inoperative or ineffective control measures, whether or not there is a risk of exceeding the individual dose limits. The ICRP recommends a procedure that can be used effectively to reduce unnecessary exposures. A n investigation level should be established at an exposure rate substantially lower than the regulatory limit, e.g. 30% of the limit (ICRP 26, 1977, p.33). Measurements obtained during routine monitoring that exceed the investigation level are evaluated with respect to cause and potential reduction. To be effective, the evaluation results should be formally recorded and conveyed both to management and to the workers involved. Although the investigation level recommended by the ICRP is based on a fixed exposure rate or derived air concentration, an equally effective evaluation program may be based on the investigation of a percentage of all measurements. For example, one might investigate the highest 5% or a random selection of all measurements. In any case, the objective is to detect and correct situations that are producing unnecessary exposures. SECOND CRITERION: MONITOR IN PROPORTION TO THE MAGNITUDE OF RISK The ICRP criteria apply generally to all radiation exposures. However, a second priority for monitoring programs should be established on the basis of the nature of the exposures and the magnitude of the health risks involved. Current practices in radiation protection are based on dose limits to specific organs and assessment of individual exposure pathways, with little consideration of the combined doses from various pathways. Although the intent of the ICRP recommendations was to limit the total dose to each critical organ, combined doses from external and internal sources are rarely determined. The recent recommendations of the ICRP (Publ. 26, 1977) are based on the limitation of total health risk from occupational radiation exposures. Implementation of this concept would necessarily require the measurement, calculation and summation of doses and concomitant risks to all organs of the body from all exposure pathways. The U.S. Environmental Protection Agency has taken the first step toward translating the ICRP recommendations into regulations. The proposed recommendations for Federal Radiation Protection Guidance for Occupational Exposures (USEPA, 1981) include provisions for summing the risk-weighted organ doses to determine compliance with an effective whole-body dose equivalent limit. Whether or not the proposed guidance is modified before final adoption, it seems clear that some version of dose summation and combined risk limitation will be included in future regulations. With this

Jan 1, 1981

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

By D. Akers, R. Dospoy

Introduction The geological processes that form coal can also concentrate trace elements in the coal. For example, the average concentration of arsenic in bituminous coal (20 ppm) is ten times the average concentration found in all the other rocks that make up the earth's crust (2 ppm). Similarly, other elements, such as antimony, cadmium, mercury and selenium, are more concentrated in coal than in the earth's crust. When coal is burned, trace elements can be further concentrated. Although no new constraints on trace clement emissions were placed on the power generation industry under the 1990 Clean Air Act Amendments, the act does mandate a three-year study of air toxics. Any new regulations aimed at electric utilities as a result of this Federally-mandated study will almost assuredly be very costly-an estimated $1 billion per year. Many trace elements in coal are associated with mineral matter. For example, arsenic is commonly associated with pyrite, cadmium with sphalerite, chromium with clay minerals, mercury with pyrite and cinnabar, nickel with millerite, pyrite and other sulfides, and selenium with lead selenide, pyrite and other sulfides (Finkelman, 1980). There are also cases in which some of these elements are organically bound. Just as both organic and pyritic sulfur can be found in the same coal, the same trace element may be both organically bound and present as part of a mineral in the sane coal. Physical coal cleaning techniques are effective in removing mineral matter from coal and can potentially remove at least some of the trace elements associated with specific minerals, thereby reducing the release of these elements into the atmosphere. Conventional coal cleaning to remove trace elements As part of a project funded by the Electric Power Research Institute, CQ Inc., a wholly-owned EPRI subsidiary located in western Pennsylvania, has demonstrated that large reductions in the concentration of many trace elements are possible if conventional coal cleaning techniques are properly applied. Four examples are given in Tables 1 to 4. In each example, the results shown were generated by cleaning the coal at CQ Inc.'s commercial-scale cleaning test facility. Cleaning results for Upper Freeport Seam coal from Northern Appalachia are provided in Table 1. Data are presented in the table in two ways: as a weight-based concentration (parts per million) and as a concentration per heat unit (grams per billion Btu). Grams per billion Btu is analogous to pounds per million Btu, but avoids the use of numbers with many decimal places. The heat-based concentration provides a better measure of boiler impacts, because the increased heating value obtained through coal cleaning reduces the number of tons that must be burned to produce a given thermal output. Reducing the quantity of coal burned reduces the quantity of trace elements entering the boiler. This raw coal is relatively high in several trace elements of environmental concern, including arsenic, cadmium and chromium. Cleaning provided large reductions in the quantity of arsenic, barium, cadmium, chromium, fluorine, lead, mercury, nickel, silver and zinc. The results for tests with a Powder River Basin coal, Rosebud/McKay, are presented in Table 2. Large reductions in arsenic, barium, cadmium, fluorine, mercury, nickel, selenium and zinc were observed with cleaning. The concentration of chromium increased with cleaning, while lead concentration increased in one test and decreased in another. Table 3 presents test results for Croweburg Seam coal from Oklahoma. Large reductions in arsenic, barium, cadmium, chromium and zinc were obtained with cleaning. Smaller reductions were obtained with lead and nickel, while chromium, fluorine and mercury increased in at least one of the tests. Table 4 presents cleaning test data for Kentucky No.11 Seam coal. In this case, large reductions were obtained with all elements measured. In general, these data indicate that physical coal cleaning is effective in reducing the concentration of many trace elements, especially if they are present in the coal at relatively high concentrations. The degree of reduction achieved is coal-specific, relating in part to the degree of mineral association of the specific trace element and the degree of liberation of the trace element-bearing mineral. The extent of trace element removal also depends on the method of cleaning the coal. Figure 1 is a washability plot, by size fraction, of arsenic vs. ash content for Upper Freeport

Jan 1, 1994

By Bernard L. Cohen

INTRODUCTION The most serious occupational radiation episode in U.S. history was the exposure of uranium miners to radon daughters in the 1945-1968 time period. Although the maximum permissible dose at that time was 12 WLM (working level months) per year, average exposures were several times that figure, and ten times higher exposures were not uncommon. Starting in the early 1960s, mounting epidemiological evidence appeared for an excess incidence of lung cancer among these miners; up to 1974, there was an excess of 134 cases (159 observed vs 25 expected) among the 4000 in the group under study (NAS-1980). When this excess became apparent, the Federal Government took jurisdiction, greatly tightened enforcement, and in 1969 lowered the maximum permissible exposure to 4 WLM per year. These standards were met largely by a great improvement in ventilation. As a result of these measures, average miner exposures were reduced 5-fold from 1965 to 1968 and by another factor of 3 by 1970. In 1978, the 4 WLM/year maximum permissible exposure was exceeded by only 37 of 7500 underground miners, for only 16 wars it exceeded by more than 25%, and for only 8 was it exceeded by 502 (AIF-1980). The average exposures to the various groups are listed in Table 1. We see that the average exposure for all miners was 1.03 WLM, and for those who worked essentially full time underground it was 1.45 WLM. For purpose of later discussion, it will be convenient to choose a single value for average exposures. In view of the fact that employment situations and job categories are bound to vary over a lifetime of work, we take this to be 1.3 WLM/year. The purpose of this paper is not to dwell on history, but rather to address the question of whether or not this present situation is satisfactory. This question was considered recently by a study group under the auspices of National Institue of Occupational Safety and Health (NIOSH-1980) and it answered the question in the negative, concluding that the maximum permissible exposure should be substantially reduced. However, their arguments were incomplete and were lacking in perspective. The principal thrust of this paper is to provide some of the material overlooked. RISK TO URANIUM MINERS FROM RADON EXPOSURE In order to quantify the risk to uranium miners under present working conditions, it is first necessary to estimate the risk of lung cancer per WLM of exposure. The most straightforward way of doing this is to use the data on the group of uranium miners under study. These are listed in Table 2 for 8 exposure ranges and for the total group (NAS-1980). If all data are given equal weight, the risk is seen to be 3.5 x 10-6 per year per WLM. Much can be said in favor of using this value for the risk as dose independent; it is within one standard deviation (SD) of the observations in 5 of the 8 dose categories (as expected from the definition of SD), and within 2 SD for all 8, and for the three cases differing from the mean by more than one SD, two are above and one is below. However, the uranium miner exposures in the present situation will be far below those covered in Table 2, so it is worth considering the possibility that the risk differs from this at low doses. From the last column of Table 2 we see that the experience from all exposures below 600 WLM indicates a risk higher than the mean for the entire group by more than two standard deviations (the value 600 WLM in this comparison was deliberately chosen to maximize this deviation; for exposures below 360 WLM and 840 WLM, the excess is only 1.0 and 1.6 standard deviations respectively). In view of this tendency for the risk to be higher at lower dose, we will take the risk to be 5.0 x 10-6 per year per WLM in what we will refer to as model A. This risk estimate is based entirely on exposures much [higher] than those that will be experienced by

Jan 1, 1981

By William D. Cassity, Batric Pesic

Introduction Gold mining on an industrial scale is becoming increasingly dependent on low-grade refractory sulfide ores that contain large amounts of pyrite (FeS2) and arsenopyrite (FeAsS). Gold is usually present as sub-microscopic particles dispersed evenly throughout the sulfide matrix. Without some form of preoxidation, efficient recovery of this finely disseminated gold through cyanide leaching is hampered because the surface of the sulfide particle becomes passivated and impervious to penetration of the cyanide. Traditional methods for recovery of refractory gold from arsenopyrite ores include roasting and autoclave leaching. A modem alternative method is to preoxidize the mineral using the bacterium Thiobacillus ferrooxidans prior to attempting to re- cover the gold through cyanidation. The bioleaching of arsenopyrite ores presents a problem in the mobilization of large quantities of arsenic, present both in the +5 and +3 ionic states. Dissolved arsenic species pose a dual problem: they are -1 to the bioleaching process itself through inhibition of bacterial activity, and they also pose an environmental threat. The arsenate anion (AsO2) has been shown to be more toxic to living organisms, including Thiobacillus ferrooxidans, than the arsenate anion (AsO43) (Collinet and Morin, 1989). There is some disagreement as to the fate of arsenite in bioleaching solution. Mandl, Matulova, and Docekalova (1992) reported that dissolved arsenite concentration was constant during a long-term bioleaching study of chalcopyrite using Thiobacillus ferrooxidans. Torma and Oolman (1992) report that Thiobacillus ferrooxidans has been found capable of oxidizing As3+ to As5+. The solid products of arsenopyrite/pyrite bioleaching include jarosite (MFe3(SO4)2(OH)6 where M can be K+, Na+, NH4+ or H3O+), scorodite (FeAsO4.2H2O), ferric hydroxides, and ferric hydroxysufites (Carlson et. al., 1992; Van Breemen, 1982; Lazaroff et. al., 1982). The particular species formed is a function mainly of pH. Little study has been performed on the effect of solid products on bioleaching. Pesic and Kim (1993) showed that Thiobacillus ferrooridans cells served as a nucleation sites for jarosite particles, which grew rapidly and eventually killed the cell. Most bioleaching studies are performed at low pH&apos;s (<2.0) under uncontrolled conditions. By studying bioleaching of arsenopyrite under controlled pH conditions and at higher pH&apos;s (2.0 to 3.0), the distribution of iron and arsenic between the solid and liquid phases can be controlled, and their effects studied. The objective is to develop a bioleaching process that operates at high efficiency while keeping amounts of dissolved heavy metals low. Pesic, Stohok, and Torma (1993) demonstrated the usefulness of this approach in the leaching of cobaltite (CoAsS) concentrates. Materials and Methods The focus of this study was the bioleaching of arsenopyrite ore by Thiobacillus ferrooxidans. The parameters studied included the effects of pH, and the effects of arsenate (AsO43 and arsenite (AsO2). pH studies included pH set initially but not controlled, and pH controlled continuously with NaOH or LiOH. Bioleaching results from two strains of Thiobacillus ferrooxidans with different adaptation histories were compared. Bioleaching experiments were conducted in 150 ml stirred glass reactors using a strain of Thiobacillus ferrooxidans adapted to arsenopyrite ore, supplied by the U.S. Bureau of Mines. These bioreactors were innoculated with stock cells grown on arsenopyrite ore for 14-16 days. Typical elemental breakdown for this ore was: As, 9.90%; Fe, 18.60%, Stot, 9.26%; Suifide, 7.50%; SiO2, 25.90%. Mineralogically, the ore contained both arsenopyrite and pyrite. Bioleaching efficiency was measured by withdrawing bioleach solution samples at selected intervals, filtering, and measuring the amount of dissolved cobalt, iron, and arsenic by atomic absorption spectroscopy. Because the precipitation of iron and arsenic at higher pH&apos;s would lead to unreliable results, a 100 mg tracer of cobaltite concentrate from the Blackbird mine in southern Idaho was added to solution. Dissolution of the cobaltite tracer was used to indicate bioleaching of the bulk arsenopyrite ore. Cobalt has been shown to be stable in solution even when iron and arsenic were precipitated (Pesic, Storhok, and Tonna, 1993). The pa-

Jan 1, 1995

By Gunter Zimmermeyer, Hartmut Eicker

Radon in the Environment Radon, as a natural nobel gas, can be detected nearly everywhere in the environment as a decay product of ubiquitous uranium. As it is emanated from soil and rocks measurable concentrations have been found at the surface of soils and in even higher concentrations in enclosed spaces like, for example, mines and buildings. While above soil surface activities caused by radon have been found in an order of magnitude of up to 1 pCi/l (Weigel, F. 1978), concentrations in enclosed spaces and mines are higher because of the lack of atmospheric circulation. Beside air circulation the relevant figure depends on the Ra226-concentration in the surrounding rocks or building material, as well as on emanation coefficient and the diffusion coefficient. While representative Rn222concentrations in well ventilated buildings are reported to be in an order of magnitude of 1 pCi/l maximum values up to one order of magnitude higher have been found in badly ventilated brick buildings (Ettenhuber, E., Lehmann, R., Clajus, P., 1978) (Aitken, J.H., et al., 1977). Just now it was stated that the reduced air circulation due to German legal regulations on energy conservation will increase radon exposure of the public considerably (Jacobi, W., 1979). Radon in Mines Radon exposure of workers is, of course, a matter of concern in uranium ore mines where relatively high concentrations of the uranium to be mined are present. Measures to protect workers&apos; health have been implemented, based on experience on dose-effect relationship. They serve to meet exposure standards by limiting inhalation of radioactive particles, in reducing radon concentrations or in limiting working hours. Both improved measuring devices and capacity as well as the lower discrimination threshold enable to measure radon concentrations in other mines, e.g. in coal mines. It is known that radioactivity in coal is small compared with that in other minerals and even soil, rocks. Nevertheless, radioactive elements were identified in coal and so the question was whether the concentrations of radon in coal mines might be a subject of concern. The problems encountered when measuring radon in coal mines are described below, as the measuring device has to be flame proofed which is an important additional requirement. Measured radon concentrations in British coal mines have already been published (Duggan, M.J., Howell, D.M., Soilleux, P.J., 1968 (Dungey, C.J., Hore, J., Walter, M.D., 1978). The authors found concentrations of up to 14 pCi/l in Cornish mines. In most cases the values were in the order of 2 pCi/l. These results were consistent with measurements reported from U.S.-coal mines (Lucas, H.F., Gabrush, A.F., 1966). Such concentrations of radon were not considered to represent a hazard for British miners (Ogden, T.L., 1974). In Germany, too, first measurements have been carried out in five coal mines in the Saarland in the 60&apos;s. Air samples were taken at different places in the coal mines, dried, fed to an ionisation cell and measured by a device including reference cells. Samples taken at ventilated places showed radon concentrations consistent with the lower British results. They all kept within the standards of the first German regulation on protection against radiation. Measuring the radon daughters was renounced because of the relatively low radon concentrations and the requirements for flame proofness in coal mines. Moreover, it can be ascertained that because of the effective ventilation the disequilibrium factor between the decay products and the radon concentration remains far below the value of one (Muth, H., 1978) (Keller, G.). In 1979 the committee on mine safety and health protection in coal and other mines of the EEC proposed to have measured and evaluated radon concentrations in European coal mines to find out whether they complied with international standards. Great Britain and Germany agreed to this proposal and by commissioning such measurements to scientific institutes complied with the request to harmonize the methods used. In the Federal Republic of Germany, e.g. Westfälische Berggewerkschaftskasse (WBK) and Staatliches Materialprüfungsamt; Dortmund (MPA) were requested to carry out the measurements in coal mines of the Ruhr coalfield whereas Saarberg Interplan was responsible for the Saar coalfield. The WBK measurements are reported in later paragraphs.

Jan 1, 1981

By Harold Stocker

INTRODUCTION This presentation is concerned with the administrative and technical capability of the Atomic Energy Control Board (AECB) to assure compliance with the individual exposure limit for radon daughters (currently 4 WLM per year). It is not concerned with the epidemiological bases for setting the exposure limit. Moreover, the intent is to show how sophisticated methodologies and advanced technologies, applied to radon daughter concentration measurements in uranium mines, convey the spirit of compliance by providing better estimates than do the historical methods. These better estimates mean that more accurate and more precise estimates of each worker&apos;s exposure are determined using these more modern methods and devices. The estimates so derived should provide more convincing evidence to an individual worker that his assigned exposure is a valid indicator of his true exposure. In addition, a perspective on the exposure estimate in relation to the exposure limit is given as further evidence that an exposure limit is not the dividing line between "safe" and "unsafe" exposures. A brief description is given of the compliance aspects of the Atomic Energy Control Regulations and of the limitations of purely statistical non-compliance procedures. Most of the emphasis of the paper will be placed on the uncertainties associated with conventional radon daughter exposure determination and the means being employed (and anticipated) to reduce these uncertainties. NON-COMPLIANCE Under current Atomic Energy Control Regulations (1978), the annual individual exposure limit for radon daughters is given without reference to the possible methods of sampling and calculation of radon daughter exposure and without any reference to possible uncertainties or their magnitudes. This is common in such statutes, the details of sampling, calculation, error analysis, and so on, being left for licence conditions or provided as a specific guideline to the licensee on the matter of compliance with the Regulation. Since the exposure limit is contained in the Regulations, compliance with it is absolute, as with any other law. In Canada, a state of non-compliance with the radon daughter exposure limit exists when an exposure (attributed to an employee) is reported by the licensee to exceed the limit. No uncertainty in the measurements or in the overall determination of exposure is reported nor is any requested. Removal of the worker and the loss of his services are the immediate and direct penalty suffered by the licensee for failure to maintain the exposure at, or below, the limit. A worker may be re-instated to employment for the balance of the reporting period only if the licensee can assure the AECB that further significant exposure to the worker will not ensue. In other jurisdictions, such as the United States, non-compliance is defined on a statistical basis. For example, NIOSH, the National Institute for Occupational Safety and Health presents procedures for calculating the 95% Lower Confidence Limit (LCL) in order to "compare the results of occupational environmental sampling to an occupational health standard and make a decision with a known chance of making an incorrect decision that a state of non-compliance exists" (Leidel and Busch, 1975). (In the nomenclature of this presentation, exposure limit would be used in place of "standard", in the NIOSH sense). Furthermore, it is emphasized in the NIOSH document that such numerical comparisons "are necessary only if the sample mean is greater than the standard". The NIOSH document points out, quite correctly, that the "statistical procedures presented below will not detect and do not allow for analysis of highly inaccurate results, i.e., systematic (non-random) errors or mistakes ... If a systematic error is known to exist in an instrument or analytical procedure then correct the sample mean of the data before analyzing for non-compliance". It is certainly not the purpose of this paper to criticize the sophisticated statistical approach to non-compliance as given in the NIOSH document or in similar approaches used in other jurisdictions. Rather, the purpose is to approach, with some introspection, the question of the determination of exposure by the employer for his employee and especially the employee&apos;s understanding of, and confidence in, the accuracy of the exposure determination and its relation to the exposure limit. DETERMINATION OF EXPOSURE Historically, in uranium mines, exposure to radon daughters for an individual miner

Jan 1, 1981

By R. M. Bricka, T. J. Olin

Heavy metals contamination is an environmental problem at Army installations engaged in firearms training and munitions production. At these facilities, weathering and corrosion of expended munitions and leaching from wastewater lagoons, landfills and burn pits have resulted in heavy metals contamination of the soil. The principal metals encountered in firing-range soils are lead, copper and zinc. Cadmium and other metals, such as antimony, that are often incorporated in the munitions are sometimes seen in lesser concentrations. Mercury is associated with various propellants and, while present in much smaller concentrations, is of concern because of its acute toxicity. Chromium is primarily associated with plating operations. The transport of metals into groundwater has been confirmed at some locations, which has required treatment of the soil and groundwater at these sites. Certain treatment processes for contaminated soil produce metals-laden extracts, which also require treatment before reuse or disposal. Ion exchange is generally quite effective for removing metals from aqueous streams. However, resins are expensive and must be regenerated, and activated carbon is generally less effective for most metals and also requires regeneration. Therefore, alternative effective and economic sorbents are needed. Twelve sorbents were screened in initial batch testing. These included activated carbon, bark, chitosan, crown ether, corn cob, xanthate, clay (kaolinite and montmorillonite), peat moss, seaweed and reagent-grade zeolite (aluminosilicate, Sigma Product No. Z3125). Of these, zeolite demonstrated the highest capacity for Pb, Cr and Cd. For this reason, zeolite was selected for further testing in batch, kinetic and column studies. Materials and methods Zeolite. The zeolite used in the second-phase batch and column studies was obtained from a natural deposit of clinoptilolite-rich rock located in South Dakota (Rocky Ford SDH) (Desborough, 1996). Large blocks of the material were crushed and sieved into the following three particle size ranges: 0.5 to 1.0, 1.0 to 4.0 and 2.0 to 4.7 mm. This material demonstrates high structural stability in acidic solutions (pH 2.5) (Desborough 1996) and has a measured surface area of 30 m2/g. The measured total cation-exchange capacity (TCEC) was approximately 10 meq/100 g. This is well below what has been indicated for commercially available zeolite, which has been reported to be about 180 to 220 meq/ 100 g. The TCEC test was repeated (Method 9081, SW 846) using sodium acetate. The test resulted in a TCEC of 54.5 meq/100 g. The difference between values obtained for this material and published values for zeolites may be attributable to the greater heterogeneity in the material used in this study, compared to commercially available materials, or to the effect of the relatively large particle sizes utilized. Batch studies. Seven batch studies were conducted using synthetic metal solutions and soil extracts (Table 1). Extract composition: The P-extract was prepared by sequential surfactant extraction of organics from a burnpit soil followed by acid extraction of metals. The pH of this solution is approximately 1.1. A number of metals and organic compounds were present in the soil. Analysis of the extract was restricted to Pb, Zn and Cu concentrations for this study. The FBH extract was produced from a firing-range soil that was oxidized with a 0.01 M CaO solution and then extracted with 0.1 M acetic acid. This was filtered through 0.5-µm Whatman No. 5 filter paper and stored at room temperature. The pH of the FBH extract was approximately 4.5. pH Control: Calcium carbonate (CaCO3) may be present in the zeolite horizon or bed. Calcium ion (Ca") is released from the exchange sites when in contact with solutions containing ions for which it is more selective, such as lead. This results in a rise in solution pH over time. Acid washing removes most of the carbonates, eliminating the need for a buffer. Batch studies were conducted using both acid-washed zeolite (AW) and unwashed zeolite (UW) for performance comparison. The UW zeolite was rinsed with distilled deionized (DDI) water to remove fine soil particles. Both materials were dried at 105°C (220°F) overnight, so that the dry mass could be determined. Column studies. Ten column studies were conducted. Because it was expected to have the best hydraulic properties, the largest particle

Jan 1, 1999