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By Rong Guo, Qi He, Jinsuo Zhang

"Coal mine ranks No.1 among the 6 major high-risk industries, so it is no doubt that the senior managers in the coal mine bear more stress than other industry executives. The massive working stress results in serious influences on the physical and mental health of the senior managers of coal mine, which might also induce the safety risk in the coal mine. Therefore, to evaluate the senior managers’ working stress and put forward strategies of mitigating the stress is of great theoretical and practical significance in improving the physical and mental health of the senior managers and reducing the safety risk in the coal mine. The research on the working stress has been made in the western countries for almost 50 years, China began the research in the mid and late 1990s, the research is spread across multiple industries including education, medical treatment, computer and aviation. However, there is little research on the senior managers’ working stress of the coal mine, so the available literatures are few. In the paper, we, based on the working stress model of Robbins and through questionnaire survey, analyze the element composition of the senior managers’ working stress of the coal mine; then, we construct the evaluation indicator system of the senior managers’ working stress of the coal mine based on the individual factor, organization factor and environment factor; after that, we determine weights of each evaluation indicators by using the improved triangular fuzzy number weight method, and set up the evaluation model of the senior managers’ working stress of the coal mine on the basis of the fuzzy comprehensive assessment method. At last, in accordance with the model, we evaluate and analyze the working stress for the senior managers of some typical coal mines, and diagnoses the major sources and intensities of the senior managers’ working stress, and then propose corresponding strategies of mitigating the stress."

Jan 1, 2017

By Alexander N. Winchell

"The deposits of the east end of the Mesabi Range cover very large areas, and consist of the same taconite that produces the hematite of the central and western Mesabi. It is not deeply covered with drift, but outcrops boldly in large masses. It is not concentrated locally into enriched bodies separated by leaner areas, but is all a fairly uniform, hard, banded, unaltered chert, carrying about 25 to 30 per cent iron as magnetite.The Mesabi Iron Company, organized by Hayden, Stone & Co., and under the direction of D. C. Jackling, has taken over these deposits and is building a Concentrating plant at the new town of Babbitt, about fourteen miles east of Mesaba, Minn. This plant is to have a capacity of 2,000 to 3,000 tons per day, and is based on the application to this low grade material of the methods which have made the so-called porphyry copper properties so successful.In 1916, an experimental mill was built in Duluth; containing full sized machinery and capable of handling from 100 to 200 tons daily. A considerable tonnage of material was concentrated here during 1916, 1917 and 1918. The work showed that the resulting product could be varied, almost at will, between 60 per cent and 70 per cent iron and .006 per cent to .030 per' cent phosphorus, and that this could be made on a large scale well within the limits of cost set by iron trade conditions. A small special cargo was sent down the lakes, late in 1918, assaying 63.27 per cent iron (natural and dry) and .008 per cent phosphorus; but it is assumed that large scale shipments will normally be of a Bessemer' grade carrying about 63 per cent iron and .020 per cent to .030 per cent phosphorus. This material is a sinter, carrying only traces of sulphur and titanium, with silica between 6 per cent and 10 , per cent, and alumina, lime, magnesia, and manganese, each below 1 per cent. Some of the iron in this sinter is hematite and some is magnetite, the percentages varying according to the treatment given on the sintering machine. Like the analyses, this hematite percentage is within reasonable control. Sinter made from finely ground magnetite has much the structure of coke, being firm yet porous, and gives good account of itself in the blast furnace."

Jan 1, 1920

By B. M. Sadashiva Murthy

Increasing quantities of nutrients and pesticides in agricultural runoff in recent years has caused the more developed western nations to become concerned about diffuse agricultural pollution. Thus, the problem of diffuse pollution is of greater importance and concern for any integrated water quality management programme. The above mentioned sources and activities causing pollution to groundwater cannot be completely avoided. The best solution is to entail a firm understanding of the principles of solute transport through porous media. The complex factors that control the movement of contaminants in porous media and the resulting behavior of contaminant plume are usually intricate to assess because of interaction of many factors such as dispersion, advection, sorption, degradation, volatilization, etc. As accurate quantification of solute under field conditions is inherently difficult, the fate of nitrate and phosphate have been analyzed in laboratory column and channel studies. From the analysis of the breakthrough curves, the dispersion co-efficient for nitrates and phosphates through sandy soil was found to be 0.262 m2/d and 0.114 m2/d for column studies, 0.321 m2/d and 0.231 m2/d for channel studies respectively. The column studies revealed that the concentration of nitrates and phosphates increased steadily in the first few days (around 5 days) and then it increased rapidly after 8 to 10 days. However, it attained equilibrium after 20 days. Similarly, the channel studies revealed a steady increase in concentration of nitrates and phosphates in the first few days (1 to 2 days) and rapid increase up to 15 days and there after it attained equilibrium. In the study, one-dimensional analytical model has been developed which include the advection, dispersion, and a characteristic term called ?total elimination rate?. The comparative study of column and channel studies, and model output (in which the term total elimination rate ?K? is considered as zero) exhibited a variation of 30 to 50 % in leaching characteristics of pollutants (nitrates and phosphates).

Jan 1, 2008

By Donald A. Dahlstrom

INTRODUCTION Solid-liquid separation is an important and many times a critical step (or steps) in a mineral or coal processing plant. Some examples of typical critical separations are as follows: 1. Operating a seven to eight stage CCD thickening circuit to re¬cover almost all of the caustic and alumina trihydrate values in an alumina plant. This must be accomplished with highly scaling slurries, starting temperatures of over 100°C and operate on a closed circuit system. 2. Making a filter step operate to produce a proper moisture for the balling step so that the pellet furnace can produce a proper product. Bentonite addition must be minimized and a good quality filtrate must be maintained to prevent abrasion of the internal parts of the filter. 3. Operating a closed water system in a coal washing plant so that reclaimed water containing less than one percent suspended solids is produced for reuse to maintain beneficiation ef¬ficiency. Refuse solids must also be thickened so that they can be effectively dewatered to meet state disposal requirements. 4. Recovering phosphoric acid values from a slurry containing gypsum and other gangue solids at an elevated temperature. This must be done while minimizing wash water and maximizing P2O5 re¬covery while simultaneously minimizing scale problems. While solid-liquid separations are important to overall performance in the processing plant flowsheet, it is also a significant factor in plant costs. Capital and operating costs for all the solid-liquid separation steps (includ¬ing water reclamation for reuse) in such plants are probably between 10 and 20%. Energy consumption can also be important in these separations. For example, in the copper concentrator, it is the third largest category in energy consumption, after comminution and flotation. Obviously, solid-liquid separation steps in a plant demand proper atten¬tion in design in order to achieve both optimum performance and cost. Simultaneously, knowledge of funda¬mentals are important to improve results in existing plants.

Jan 1, 1986

By Michael Drozd

Introduction Flotation is a recovery process that utilizes natural or chemically-introduced hydrophobicity of minerals to separate ore from gangue by attaching the ore particles to bubbles. Flotation has been known since 1906, when it was patented by E. L. Sulman, H. G. K. Pickard and John Ballot. Flotation's first documented plant use was in 1911, when it was used on gravity tailings from a zinc plant. Cyanide leaching was first patented in 1887. Most early precious metals flotation was linked to lead or copper concentrates. Most precious metal flotation is linked to sulfide recovery and utilizes such reagents as xanthates, dithiophosphates and thiocarbanilides. But the involvement of flotation in precious metals recovery does not have to be limited to active flotation of the gold and silver bearing minerals. Flotation can be used to remove interfering minerals, such as carbon, or cyanicides, such as copper, arsenic bearing minerals or other cyanide complexing minerals. Cyanide Leaching The use of cyanide to recover precious metals is the backbone of the industry. The various recovery processes that use cyanide are: 1. Leaching-Counter Current Decantation(CCD) Merrill-Crowe 2. Carbon-in-pulp (CIP) 3. Carbon-in-leach (CIL) 4. Vat leaching 5. Heap leaching These processes are selected depending on ore grade, ore leaching characteristics, reserve size and capital limitations. Ore grade is used as the first criteria to determine the type of cyanide plant required. However, the reserve size can determine whether a ground ore alternative is used. The ore leaching characteristics may indicate that the higher recovery, associated with a process that employs grinding, is the only economic recovery alternative. There are some extremely large leaching plants with grades that would be non-economic for some existing heap leach plants. Many factors, such as strip ratios, metal recovery and reserve size, determine what the ultimate plant design will be. The exact processing scheme that is used, is a multifaceted decision matrix, which gives the process metallurgist some job security. Precious Metal Recovery Using Flotation Flotation recovery can conceivably be used on any ore that has been ground. Many operations utilize gravity separation on the ore to remove native gold from the ore prior to sending it to cyanide leaching. On certain ores, where the gold is in visible nuggets or particles, this processing alternative is advantageous because the native gold cyanide recovery is

Jan 1, 1993

By William D. Meakins

INTRODUCTION Today, auxiliary fans, or boosters as they are sometimes called, are commonly used underground to provide ventilation for safe working conditions for personnel. Fans are installed in many ways and for many purposes; however, their use falls into three main categories depending upon the application. Auxiliary Face Ventilation For Coal Mines Usually the application of auxiliary fans underground in coal mines presents problems not normally encountered in other types of mining. Two of the major problems deal with liberation of methane at the coal face and the production of coal dust in the mining process. For these conditions it is necessary to use permissible (explosion-proof) type of equipment. As a result, these fans are discussed in a category by themselves. Auxiliary Fans For Tunnel and Drift Ventilation Tunnels and drifts present unique ventilation requirements. For instance, tunnels and drifts can be developed for considerable distances into virgin territory. Since they are generally dead-ended, the only way to obtain ventilation at the face is to use ventilation pipe to establish one leg of a ventilation circuit. The ventilation of tunnels and drifts therefore falls into another natural category. Other Uses of Auxiliary Fans For Mining Purposes Fans can be applied to mining applications in many ways other than those discussed under the other two categories. For example, many times auxiliary fans can do a much more practical job of proper air control and distribution that can be achieved with any other mechanical means; therefore, these other uses are discussed as a special category. AUXILIARY FACE VENTILATION FOR COAL MINES After 1936, federal and state coal mining laws prohibited the use of auxiliary underground fans in coal mines since such equipment had been widely misused. In certain cases, recirculation of face air had been allowed to occur thereby creating a serious potential explosive hazard by the possible buildup of excessive methane concentrations. By the late 1950s, mining machines with rapid face penetration were commonly being used in coal seams which liberated higher quantities of methane than in earlier years. As a consequence, in the late 1950s and early 1960s, the use of auxiliary face fans was again allowed under strict experimental conditions because conventional methods of face ventilation control without fans were becoming inadequate. Today, auxiliary face fans are widely used under strict federal and state control. Such auxiliary ventilation does not reduce the need for the amount of primary air from the main mine fans; however, it does help to distribute primary air more efficiently at the face. The present trend is toward an exhaust system of ventilation utilizing 406- to 610-mm (16- to 24-in.) diam tubing extending from the face to an outby auxiliary fan. Since the tubing is normally applied with negative pressure, spiral wire reinforced tubing has been replaced largely with metal and, more commonly, fiberglass tub¬ing. Additional auxiliary blower units are often used at the mouth of the room being developed, or installed on the miner or loader near the face, depending upon the amount of gas liberation present. Often one (or more) of the fan units is equipped with an integrally mounted hopper and feeder which automatically adds a continuously metered amount of rock dust to the exhaust air. The combination unit tends to make the mixture of discharge particles noncombustible and reduces to some extent the need for additional rock dusting of return airways. In this part of the chapter, typical auxiliary ventilation systems are described including curves and photograph of typical/applicable equipment. Reference will be made to "blowers" and "ex¬hausters." Both are fans and are basically the same units. They take on their name through the manner in which they are used. A blower usually has an inlet bell on the inlet side to help to obtain good air flow into the blades. An exhauster often has a discharge cone added to reduce the discharge loss. Any basic fan unit can be used either as a blower or an exhauster. Description of Auxiliary Ventilation Systems Several auxiliary ventilation systems and modifications of basic systems have evolved over recent years. Some of these systems are described and illustrated. System No. 1, an exhaust system for faces with comparatively low gas liberation, is shown in Fig. 1. Here

Jan 1, 1982

By Arthur S. Radtke

The structural setting of disseminated-replacement gold deposits of the Carlin-type and other aspects of their genesis are closely related to events which took place during the evolution of the Basin and Range province. The deposits formed as a result. of chemical and physical processes which took place at shallow depths in the roots of hot spring systems at moderate temperatures (about 200°C) and low pressures (about 25 bars) (Radtke and Dickson, 1974). Locations of Carlin-type gold deposits in the Basin and Range province are shown in Figure 1. Most of the block-faulted mountain ranges in the Great Basin formed by tectonic activity that happened since Miocene time, although faulting in some places probably began in Eocene to Oligocene time (Hamilton and Myers, 1966). The Carlin-type deposits are spatially related to Basin and Range-type faults: Many deposits, such as Getchell, Pinson, Preble, and Cortez, occur along the flanks of ranges, while other deposits, including Carlin (Radtke, 1974) occur within the ranges in areas characterized by crosscutting and intersecting high-angle or Basin and Range-type faults. The association of the ore deposits with Basin and Range-type faults is evidence that ore deposition is young, Miocene or later. The steep, high-angle faults played three roles: first, they provided deep channels for upward-flowing solutions; second, in preparing the rocks for fluid penetration by brecciation and general rock dislocation; and third, they moved sedimentary carbonate rocks, favorable for ore deposition, from deep levels to near-surface positions where the physico-chemical conditions were more suitable for the deposition of ore.

Jan 1, 1976

By Larry J. Hughes

Ground-based electrical geophysical techniques have be valuable for three indirect exploration strategies: "tracer" sulfide mineralization, 2) mapping structure with or controlling mineralization, and 3) mapping proven to 1) mapping? associated alteration zones. When sulfide mineralization is associated with gold, it often can be mapped using induced polarization (IP). This technique can detect pyrite in amounts as small as 0.2% by volume, and in some cases can distinguish pyrite from anomalies due to graphite and clays. A case history illustrates mapping sulfides in a shear zone. Structure and alteration mapping are often the domain of controlled source audio-frequency magnetotellurics (CSAMT) due to the superior lateral resolution and depth penetration of this technique. Typical CSAMT targets include silicified sediments and volcanics, faults, brecciated zones, argillic alteration, clay and graphite horizons, and certain lithologies. Examples of mapping silicification and mapping epithermal alteration are presented in CSAMT case histories over Cannon mine in Washington and Hishikari mine in southern Japan. In our experience, four things are needed to use geophysics effectively in precious metals exploration. First, the explorationist must correctly identify the exploration stage at which the geophysics is most useful. Second, the explorationist must clearly define the types of "tracer" mineralization, host geology, or alteration which need to be defined by the geophysics. Third, the right electrical technique needs to be applied to the problem to provide the best chance of success. Fourth, the explorationist must be willing to commit adequate finances to do the geophysics properly. With careful adherence to these principles, geophysics can greatly enhance exploration success while reducing the total cost of exploration.

Jan 1, 1989

Production In 1993, it was estimated that United States mine production of barite increased 24%. Barite consumption, either sold or used by grinding plants, increased by 39% and imports increased 4%, according to the US Bureau of Mines (USBM). Nevada continued to be the leading producer of barite with 80% of the total, followed by Georgia. It is estimated that 83% of the domestically mined barite was used as a weighting agent in drilling fluids. The remaining 17% was used in barium chemicals, glass or as a filler. Most of the production from Missouri, Georgia and Tennessee was used in the non-oilfield sector. Prices The value of domestically produced barite, f.o.b. mine, decreased 1%, according to the USBM. The declared value c.i.f. US port of all imported crude barite during 1993 decreased from $40.09/t ($36.37/st) in 1992 to $37.72/t ($34.22/st). Uses Of the total barite used by grinding plants and chemical manufacturers, it is estimated that 90% was used in drilling fluids and 10% in the non-oilfield sector. This included barium chemicals, fillers in paper, plastics, brake shoes and glass. The tonnage consumed in the drilling fluids market increased due to the higher average rig count during 1993. The average number of rotary rigs operating in the United States was 754, 5% greater than the 1992 average of 717. Imports Crude barite imports in 1993 were 443 kt (488,000 st) greater than 1992, representing an increase of 137%. Imports accounted for 67% of the barite consumed (apparent) in the United States.

Jan 1, 1994

By Wayne F. Crockett, Martin H. Col

Gold, as a precious metal, has been sought for several thousand years. During this time economics have generally determined whether or not a deposit is minable. Today, few would spend $500.00 or $1,000.00 to produce one ounce of gold, yet most gold deposits currently being mined would have such costs if they were mined using the methods of the 18th century. Innovation has been the key to success in the world of mining. We are constantly having to look for new ways to mine deposits of lower grade, at greater depths, or with higher stripping ratios. The mining of placer gold deposits has always been subject to the volatile swings of the world price and the constant rise in production costs. The rise in costs is compounded by the drop in overall ore grades. Occasionally, a major technological innovation will dramatically reduce the costs of placer mining to allow previously marginal or unminable deposits to be mined at a profit. The continuous bucketline dredge was the last such major innovation in the gold placer mining. The continuous bucketline dredge has allowed large volumes of placer deposits to be picked up and processed on a continuous basis. Unfortunately, as the overall grade of the deposit declines, it becomes impossible to process enough ore to pay for the operating costs. There are many gold deposits around the world waiting for the next technological innovation that will allow them to be mined. One such deposit has been the Grey River deposit near Greymouth in New Zealand. The deposit contains gold reserves of 8992 kg, yet the deposit has been sitting idle for many years because the average value

Jan 1, 1988

By Alan G. Fricker

Cyprus Gold NZ Ltd will be recovering residual free cyanide in barren gold pulp by air stripping. Laboratory investigations by DSIR have indicated additional advantages by recovering the cyanide from pregnant pulp. These are a lower leach pH, a greater cyanide recovery and less carbon consumption. They have a net economic advantage of $1.20/tonne over recovery from barren pulp. The realisation of these advantages is dependent on three crucial technical aspects -optimum pH for dissolution, the absolute loss of cyanide and the stability of the aurocyanide ion. The laboratory investigations described herein address these aspects. Earlier work had indicated that the rate of dissolution of freely revolving gold discs at constant cyanide ion concentration increased with decreasing pH from 11 to 8. Attempts to repeat the work in a more controlled manner by measuring the corrosion potential and current density have, to date, not been reproducible, but do not contradict the earlier observations. Cyanide recoveries and balances were obtained on leaches at pH 10.5 and 9 at the same cyanide ion concentration -the lower pH requiring three times as much total cyanide. Unaccounted-for losses, unfortunately, exceed the accounted-for losses, making interpretation of the balances difficult. Nevertheless, it appears that the additional amount of cyanide at the lower pH of 9 may not be fully recoverable. It may not, however, prove necessary in practice to use more cyanide at a lower pH. Both the gold and silver cyanide complexes remain available for carbon adsorption, as do mercury, copper and iron. Zinc, and probably cadmium, nickel and lead, may precipitate during stripping, which is an environmental advantage.

Jan 1, 1991

By M. C. Williams

The dynamic behavior of batch mill, CSTR mill, and a closed grinding network consisting of a mill, sump, and cyclone was analyzed using the dynamic population balance model (PBM). The dynamic solution of the PBM of a batch, CSTR and a closed grinding network consisting of a mill, sump, and cyclone forms the basis of the dynamic analysis presented here. Two numerical dynamic solution approaches were used. These are: 1) providing additional constraints on breakage selection functions or 2) performing the Arbiter-Bhrany (or other) normalization of the selection functions. Actual experimental anthracite batch grinding data was used to obtain the functionality of the batch dynamic mill selection and breakage functions for a real physical system. The Levenberg-Marquardt algorithm for systems of constrained non-linear equations is used to solve the batch dynamic PBM grinding equations to obtain the grinding selection and breakage rate functions. The mill, sump and hydrocyclone were modeled as a CSTR operating at various retention times. Batch dynamic PBM data was used to provide the mill kinetic and breakage selection function data. Different dynamic solutions were obtained depending on the numerical approach used. Each solution approach to a dynamic PBM with transport, while giving the same prediction for a single batch grinding time, gives different solutions or predictions for mill composition for other grinding times. This fact makes dynamic nodal analysis and control problematic. The fact that the constraint solution approach gives a solution may suggest that normalization for closed networks is not necessary. Differences in solutions to the PBM cannot be excused away by inaccuracies in the data used to model the grinding phenomenon.

Jan 1, 1995

By A. J. Kroha, N. Wesis

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

Jan 1, 1985

By D. P. Plute

The mechanization of the materials management function must contribute to the overall goal of optimization of profits. A computerized system contributes to the profitability of the company by saving more than it costs to operate. Design and implementation of the sys- tem must respond to the needs of users, the operating environment of the property, and must support the objectives of materials management. Three basic functions interact within the scope of materials management. They are inventory control, purchasing, and warehousing which have the following objectives: 1) Inventory control-to request materials in a timely manner and in quantities that provide the best level of support to users at the lowest cost. 2) Purchasing-to purchase requested materials and services at the required time and at the lowest cost. 3) Warehousing-to efficiently receive, store, and issue materials and to prepare accurate transaction documents for the activities. The development of the mechanized materials management system will require a plan and objective analysis by management and system users of system requirements. A typical plan would include: 1) System Design Objectives-the development of objectives that support mechanization of the materials management function. 2) System Characteristics and Concepts- the development of key system concepts and characteristics, which identify the needs of materials management. 3) System Inventory Item Classification- development and assignment of inventory category codes to inventory items. 4) System Order Decision Rules-the development of decision rules and forecasting techniques to minimize the inventory investment and to provide acceptable support levels. 5) System Report Structure-the development of a report structure to support the goals and objectives of materials management. 6) System Report Design-the development and design of reports for effective utilization of the mechanized system.

Jan 1, 1979

By Frank W. Bowdish

During the early 1960's, a method was developed for refining sand from the Arkansas River to produce a feldspathic product that was proposed for use in making colorless glass. Because of the need for chemical uniformity, a cooperating glass manufacturer specified certain ranges of percentages for each chemical constituent of the sand. These specifications were met by fulfilling the single requirement that alumina vary by no more than 20.1 percent from the specified value. This translated to a total variation of only about 1 percent in the feldspar content of the mixed quartz-feldspar products. Because the feldspar was lighter than the quartz, the average density of refined sand varied depending on composition. Densities, determined to five significant figures, of several samples were correlated with their chemical analyses, and the correlation could then be used with density determinations to more reliably indicate compositions than by chemical analyses only. The effects of several variables on the composition of refined sands were studied. There was a regular variation in the proportion of feldspar with screen size. A series of samples with their natural size distributions were more nearly uniform in composition than the same sands after screening and recombination. The screening and recombination process was done so that the sands had identical size distributions. Variations in the refining showed little effect on composition. A pile of flow-back sand from a construction sand washer was found to be severely segregated. However, samples taken from many points in the river near the sand plant and from the flow-back discharge pipe over several weeks very nearly met the uniformity specification. If care was taken to avoid segregation either before or during processing, it should be possible to meet the uniformity specifications for a feldspathic glass sand.

Jan 1, 1978

By R. Gong, R. Bhaskar

[Full-scale laboratory experiments were performed to evaluate the effect of ventilation and water spray pressure on the potential dust capture efficiency of a ventilation system comprised of blowing ventilation and an underboon scrubber. The tests were performed for a 1.52-m (5-ft) roof height and 2.36 m-3/s (5000 cfm) blowing ventilation volume, and a 2.44-m (8-ft) roof height and 4.72 m3/s (10.000 cfm) blowing ventilation volume at three face positions of the miner. The scrubber suction rate was varied for the tests to study the effect of scrubber flow rate on capture efficiency. The airflow patterns and scrubber efficiency were determined by performing tracer gas tests. Tracer gas was released through a manifold placed in front of the rotating miner drum. The gas was sampled in a grid fashion using a real-time multi-point gas sampling system. The capture efficiency of the scrubber was evaluated based an the peak gas concentration of the gas laden air drawn through the scrubber and the amount of gas passing, through the scrubber as a fraction of the gas released at the face. The data were analyzed to yield scrubber peak concentrations, potential exposures at various operator locations and the gas removal percentage (GRP). The underboom scrubber appears to work best in the end sump position, relative to the start sump and start slab positions. The capture efficiency scrubber was higher for 2.44-m (8-ft) roof heights compared to 1.52-m (5-ft) roof heights. The data also showed that the scrubber air discharged behind the mining machine may be drawn into the face area by the action of the blowing ventilation system, resulting in recirculation.]

Jan 1, 1994

By Walter E. Jaworski, William R. Beloff, John Dunnicliff

As a step in the endeavor to reduce future tunneling costs in New York City, the Environmental Protection Administration, Dept. of Water Resources, instituted a program in 1972 to monitor the performance of the lining of one of its tunnels. A monitoring program was included as part of the contract specifications for the East Branch Interceptor Sewer tunnel for the Port Richmond Water Pollution Control Project in the hope that measured data would indicate possible future cost effective design or construction improvements. The monitoring program involved measuring ground deformations, circumferential liner stresses, longitudinal liner stresses, and jacking pressures. This paper presents some of the findings from these measurements, with emphasis on liner stresses. PROJECT DESCRIPTION The 3.05m (10 ft) diameter steel lined tunnel was constructed under two separate contracts. The easterly portion, Contract 3A, was approximately 3109m (10,200 ft) in length and constructed by Peter Kiewit Sons Co.1 The westerly portion, Contract 3B, was approximately 2160m (7100 ft) in length, and constructed by Richmond Constructors, a joint venture of Grow Tunneling Corp., MacLean Grove & Company, Inc., Grove, Shepherd. Wilson & Kruge, Inc., Morrison-Knudsen Company, Inc., and Andrew Catapano Co., Inc. Each contract included three test stations, hence a total of six along the tunnel route. Excavated material along the tunnel route consisted primarily of dense fine to coarse sand with varying percentages of gravel and silt with occasional boulders. Pockets of organic silt were encountered on Contract 3B and a large portion of Contract 3A ran through slightly weathered serpentine and schist. Overlying material and the groundwater level varied considerably along the route as did the height of overburden. Figures 1 and 2 show the generalized soil profiles for Contracts 3A (PK stations) and 3B (RC stations) respectively.

Jan 1, 1979

By Gus H. Goudarzi

The apatite deposits in east-central Bahia, Brazil, are about 50 km east of Serra do Jacobina, between lat. 11° and 12° S. and long. 39°40' and 40° W. Apatite was found and examined at 15 localities in a mineralized belt that strikes about N.25° - 30 W. over a distance of about 100 km, but the size of the deposits has not yet been determined. Rocks of the region consist of a sequence of Precambrian migmatites and quartz feldspar paragneisses containing units of pyroxene-rich rocks, lenses of sillimanite-bearing quartzite, amphibolite zones, and calc-silicate marble lenses. Pegmatites and granites of probable anatectic origin (anatexite) are abundant. Faults are numerous. Clastic sediments of Tertiary and Quaternary ages cover most of the area; thus, outcrops of the Precambrian rocks are sparse. Trenches and pits have exposed mineralized zones as much as 10 m wide at some localities within the region. A layer of apatite at least 1 ½ m thick is exposed in a trench at Lagoa do Mel. At Rio do Peixe, vertical shafts have penetrated a mineralized zone, to a depth of as much as 13 m, of which about 25 percent is apatite. Apatite is massive or forms minute crystals, stringers, and disseminated grains in alkalic host rocks. At Apolinario, euhedral crystals of apatite are as much as 50 cm long and 30 cm in diameter. Associated minerals include vermiculite, biotite, scapolite, opal, chalcedony, epidote, montmorillonite, chlorite, kaolinite, sericite, magnetite, and limonite. Apatite is everywhere intimately associated with pyroxene-rich rocks, alkalic syenite pegmatites, and alkalic granite pegmatites. At Panelas-Retiro, apatite is associated with pegmatites, crystalline calcite (vein calcite), vermiculite, and pyroxene-bearing rocks. Analyses of samples show that the apatite is radioactive and rare-earth-bearing. Traces of vanadium and titanium were detected by spectroscopic analysis. The apatite was probably derived from phosphatic zones in the original Precambrian sedimentary rocks and has been crystallized and some mobilized by high-temperature regional metamorphism (higher amphibolite facies) into veins, pegmatites, and along certain preferred beds and fractures. Radiometric and geophysical surveys, in addition to conventional geological investigations, are planned to further investigate the nature and economic potential of the deposits.

Jan 1, 1972

By Omar A. Muhtadi, David A. Milligan

7.1 INTRODUCTION The circulating fluid (generally an aqueous cyanide solution) is the major active component of heap leaching. This solution extracts precious metal from the heap and is then processed to recover the precious metal. The solution continuously recycles through the heap to recover more precious metal. With each cycle through the heap, water and chemicals are lost. Make-up water and chemicals must be added to balance losses from evaporation, seepage, entrainment, and chemical reaction in the heap, as well as any other processing losses. Adding more water than needed can result in a build-up of water as inventory in the various ponds and can eventually require that water be discharged from the operation. This is an extremely undesirable condition, and can often be very expensive in both investment in water treating facilities and operating costs. Most heap leach facilities operate under a zero discharge permit; thus, a situation requiring water discharge can result in permit violation. Chemical solution control must be managed to maintain the optimum pH, dissolved oxygen content, and cyanide content, while preventing "scaling". Calcium-based precipitation is a common scaling problem in mill reclaim waters. A buildup of these deposits can result in severe reduction of water flow, loss of pump efficiency and eventual interruption in metal production. In addition, problems of plugged sprinkler heads, blinded leach pads and fouled carbon (in carbon recovery systems) can be and often are experienced when scale control is not maintained. The remainder of this chapter discusses chemical solution control and scaling prediction/prevention in the context of a "closed" system. Sizing of ponds to accommodate precipitation and runoff, collection systems, and other similar design considerations are not discussed here. These topics are deferred to later chapters, e.g. Chapter 10 considers design of collection systems while Chapter 12 discusses pond sizing. 7.2 CHEMICAL CONTROL Chemical composition of the circulating solution must be controlled by the operator if the heap leaching process is to be successful. The primary or essential chemical control parameters in any heap leaching operation are:

Jan 1, 1988

By R. T. Webber, D. Moffett, D. W. Averill, G. H. Kassakhian

INTRODUCTION In Canada, radium-226 is removed from uranium mining and milling effluents by the addition of barium chloride to precipitate barium-radium sulphate. Large settling ponds are provided to effect solid-liquid separation. However, the barium-radium sulphate precipitate [(Ba,Ra)S04] is a fine, slow-settling crystalline solid which is not easily removed by simple sedimentation systems. Consequently, settling pond effluents may contain appreciable quantities of particulate or suspended radium-226, depending upon the specific pond system and/or the time of year considered. Also, current technology is expected to be incompatible with ultimate disposal goals since safe recovery of the barium-radium sulphate sludge from the settling ponds may be difficult to achieve (IAEA, 1976). Under the auspices of Environment Canada, a Government-Industry Mining Task Force established a Radioactivity Sub-Group in 1974 to assist in the development of effluent regulations and guidelines for the uranium mining industry. Among the recommendations of the Sub-Group was the need to seek more effective methods of radium removal, including development of physical--chemical treatment systems as an alternative to settling ponds. In response to the Sub-Group recommendations, a bench scale feasibility study began in March 1976 at the Wastewater Technology Centre (WTC) of Environment Canada. The bench scale work was expanded and accelerated in 1977 with financial assistance provided by the Atomic Energy Control Board. The results of this study indicated that coagulation and flocculation of the (Ba,Ra)S04 precipitate greatly increased settling rates (Wilkinson and Cohen, 1977). Subsequently, a research program with a larger scope was initiated in January 1978; this program is jointly managed and funded by government and industry as indicated in Table 1. The principal goal of the Joint Government-Industry Program is the development of a physical-chemical process to reduce the radium-226 content of uranium mining and milling effluents. The following effluent radium-226 activities were established as targets for the program: 0.37 Bq/L (10 pCi/L) as total radium-226; 0.11 Bq/L (3 pCi/L) as dissolved radium-226. Process development work was undertaken at bench scale at the WTC and at

Jan 1, 1980