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By Ehab Gomaa, K. A. El‑Nagdy

Backfilling material such as tailing (mine wastes) mixing with cement or gypsum has grown throughout the world’s underground mines. However, despite their growing popularity, the typical hydraulic and mechanical fill types utilized in many mines still exist. Deep underground mining has increased due to the lack of commercial minerals nearby. Mine wastes were considered the main part of backfilling to prevent environmental pollution, ground subsidence after mine abandonment, and mine collapse during deeper extraction phases. The cemented backfill technique is the principal technique used in underground mines, which include cement with fly ash and/or filter dust, cement with tailing material and fly ash, gypsum with fly ash, and synthetic anhydrite with fly ash and have been reviewed. It has concluded that a backfilling material must be selected based on further goals, available material near the mine site, and economic factors. This paper analyzes different backfill material mixtures to create a technique that will increase safety in underground mining conditions and foresees an appropriate formula that gives high uniaxial compressive strength. The multiple linear regression (MLR) on the collected data from the experimental works to construct the relationship between the uniaxial compressive strength (UCS) of the mixture and the components of the backfilling and the prediction formula for expected compressive strength was obtained. The results revealed that the predicted regression equation was robust and reliable to predict the (UCS) for the new components of the filling (cement (CE), filter dust (FD), water content (WC), and time (T)).

Oct 3, 2023

By B. Samanta, S. Bandopadhyay

Ore grade modeling of a placer gold deposit is considered a challenging problem even today. Typically, gold-bearing deposits show extremely erratic and unpredictable grade variation. This is principally because of isolated occurrences of gold metal within a few high-grade pockets. Because of sporadic fluctuation of high-gold-grade distribution, the overall spatial continuity of the deposit becomes poor. This type of variation is often described by a lognormal or highly skewed frequency distribution, and the spatial continuity of the deposit is characteristically represented by a variogram model that exhibits a high nugget component. Further complicating the matter is the sparse availability of drillholes to capture such erratic variation. Due to the prohibitive cost of drilling, gold grade and reserve estimations are typically carried out using insufficient exploratory data. Use of a traditional estimation technique on limited data often produces an unreliable ore grade model. Kriging and its nonlinear variation, lognormal kriging, are frequently used in gold deposit modeling (Rendu, 1979; Journel, 1980); however, kriging is flawed by the smoothing effect. Moreover, though kriging is known to be globally unbiased, it is not free from conditional bias. As a result, if a deposit is modeled using the traditional kriging technique, a high-grade zone can be predicted to be a low-grade zone and vice versa. Predicting high grade as low grade results in a missed opportunity to exploit profitable ore. On the other hand, prediction of low-grade ore as high-grade ore might result in financial loss for a mining operation. Thus, the use of kriging for ore grade mapping of a gold deposit may not be recommended. Estimation techniques (including kriging) can produce an estimate with minimum error in a minimum-error variance sense. However, no matter which estimation technique is used for grade modeling of a highly erratic deposit like gold, the error margin is bound to be high. Therefore, any grade prediction made for an unknown location is associated with high uncertainty. Since various mine-planning operations, including derivation of grade tonnage curve, pit design and economic analysis of the deposit, are based on estimated grade rather than true grade, a mining project evaluation that depends upon estimation will be subject to considerable risk. Thus, the successful design of a mining project not only depends upon accurate grade estimates, but also upon accurate assessment of uncertainty, for predictive grade is equally important. Although the kriging standard deviation is used as an indicator of uncertainty for prediction, it is flawed by characteristics related to a data-independence property. This study, therefore, investigates the stochastic simulation technique of conditional simulation for ore grade modeling of a placer gold deposit at Nome, Alaska. It is expected that the research will provide a better solution to ore-grade deposit modeling than the kriging technique because (1) conditional simulation preserves spatial continuity as well as reproduces the data histogram, thus avoiding the smoothing problem; and (2) generation of multiple realizations helps to accurately quantify the uncertainty about grade values in the deposit.

Jan 1, 2009

By Richard L. Bullock

At the beginning of the previous chapter, variation in the types of room-and-pillar stopes were briefly identified so that the reader could begin to understand the extensive application of this mining system. These same variations will be reviewed and elaborated upon in this chapter while discussing the methods by which the material is extracted. FULL-FACE SLICING BY DRILLING AND BLASTING When the entire mineralized thickness is taken in one pass of mining, it is known as full-face slicing. There is no mineral of economic value intentionally left either in the floor or the back. Theoretically, there is no limit as to how high the face could be in a single pass. But there are practical limitations to equipment size, pillar stability, and control of local loose slabs on the face, roof, and pillars that dictate current practices. The range of face height of the 15 typical room- and-pillar mines covered in Table 8 (Sec. 1, Subsec. 7) from the Dravo report, 1974 was 1.7 m (5.5 ft) to 9.8 m (32.1 ft), the average being 5.2 m (17.1 ft). In the Gaspe copper mine, Murduchville, Que., the drill jumbos were constructed with an extendable tower for full-face mining up to 15.2 m (49.9 ft) if necessary. However, a more common practice at that mine is to take the ore in multiple slices (Hall, 1959). In many metal mines the normal practice when starting stoping is to drive a single development drift into the ore zone a distance that will allow at least four or five rooms to be opened on each side of the initial drift. Since this opening will probably serve as the main haulage drift, it should be kept as straight and as level as possible. As to the other criteria for the stope developments, it might be well at this point for the mine planner to review the remarks presented earlier (Sec. 1, Subsec. 7) on drifts, entries, and crosscuts for production. After the initial drift is driven, it should be slabbed to the full room width if it is not that wide already. Next, the future pillars should be marked on the ribs and the rooms driven between them. If the ore extends beyond the length of the initial drift and is to be mined in the normal sequence of mining, care should be taken in a random room-and-pillar system to see that the extension of the initial roadway remains straight. If pillar "spotting" is left to the shift foreman or to the miners, they probably will not realize the importance of maintaining a uniform pillar line next to the roadway, and of keeping the line straight. Invariably, a pillar will end up in the middle of the future drift extension resulting in a "dogleg" in a possible main haul- age road. While this error is not catastrophic, too much weaving in and out of pillars will certainly slow down future production haulage and needlessly increase costs. Face or Breast Drilling and Blasting Practices To drill and blast the initial advances into the rock, usually some form of cut pattern must be used. In room-and-pillar mining, when a pattern is drilled in a face and that face is the only exposed surface to which all of the rock must be broken, the pattern is known as a "face round" or "swing." The face to which the rock must be broken is referred to as a "free face." Thus the most common way to advance a room into virgin rock with only one free face is by drilling swings. If, after breaking around a pillar, taking down back, or taking up bottom, there is a second (or third) face exposed, a group of holes drilled nearly parallel to a free face and breaking to it is known as a "slab round" or "slabbing." In some mining districts it is also known as "slashing." Obviously, because two faces exposed offer less resistance to fragmentation than a single face, slabbing requires less drilling and fewer explosives than required to break the same tonnage with the same degree of fragmentation with swings. Therefore, it behooves the driller (or supervisor) to plan the combination of

Jan 1, 1982

By R. Gradnik

The continuous need for infrastructure expansion forces the mining industry to develop underground structures into more sensitive and violent environments. When underground excavation enters into deep and high stress rock conditions, geotechnical designers find themselves faced with the difficult task of selecting a rock support system that will meet today's stringent requirements. Whatever support system is chosen, it is mandatory that potential rock movement be controlled. All this has intensified the development of safety systems and new, unique solutions which can enable construction companies to go both, further and deeper, without reducing the level of safety. To address these needs, Atlas Copco developed the Roofex® bolt, specifically designed to withstand large static and dynamic deformations under prespecified load conditions. The Roofex Rock Bolt acts as stiff support up to a designable load level before the sliding effect of the mechanical frictional system is triggered and allows highly controlled rock mass movement without loosing the support function.

Jan 1, 2011

By R. O. Ifill

Alternating current electrocoagulation (AC/EC) may be an alternative to chemical coagulation in the thickening of tailings in the mineral, coal and petroleum industries. In the current study, ground silica (d50 as received, 20 µm) served as the model tailings solid with characteristically poor settling behavior. Its natural initial settling rate was 1.6 x 10-3 cm/min. This paper discusses the indirect application of AC/EC via aluminum electrodes to the enhancement of the settling behavior of the silica. ?Indirect AC/EC? means that the silica was coagulated upon dispersion in water that had been treated by AC/EC. The effect of AC/EC retention time on settling behavior is examined and the water chemistry is analyzed. Test results show that the initial settling rate of silica was enhanced by a factor of 4,000 at its maximum of 6.4 cm/min, under the conditions employed. Settling enhancement was achieved upon contact between the AC/EC-treated water and the silica, and the enhanced initial settling rate was dictated by the floc size. The aluminum dosage increased with AC/EC retention time, reaching a maximum concentration that was equivalent to 13.1% of that predicted by Faraday?s law. Complex aluminum ions are considered to have adsorbed onto the silica surface and an adsorption mass ratio of 0.3 mg Al/g silica was the most effective in enhancing settling behavior. Owing to pH change or slow kinetics, aluminum hydroxide precipitation did not lower the ion product ([Al3+][OH-]3) below the KSP of 6.3 x 10-32.

Jan 1, 2011

By A. Rajoria

A dynamic process model for predicting the fluidized bed roasting of copper concentrates with high arsenic content is presented in this work. Mass and energy balances were performed based on roasting mechanism derived from thermodynamic data of species. The reaction rate data, obtained from pilot plant studies on major reactions, were used in tuning the kinetic parameters of the model. DEAS, a dynamic simulation package, is used in setting up the fluidized bed roaster plant model. The effect of process parameters such as concentrate feed rate, concentrate composition, airflow and cooling water rate on the temperature and conversion of roasting reactions are analyzed. The plant model captures the highly transient conditions occurring during the plant start-up, shutdown and other process perturbations. The application of this model in constructing a virtual plant environment for validating the operating procedures and operator training is presented.

Feb 23, 2014

By J. W. Wilson

In the ultrafine coal cleaning process, small coal particles (-400 mesh) are produced during the pulverizing operation. After the coal cleaning process, these ultrafine particles are very difficult to dewater, and create problems in coal transportation, storage and handling at utility plants. Research work on ultrafine coal dewatering and briquetting has resulted in the development of a single stage operation that is potentially more effective and economic than conventional thermal drying and briquetting processes. The final coal product have a moisture content of less than 15% and possess high strength and water resistance that satisfies coal transportation, storage and handling requirements.

Jan 1, 1995

By J. R. Boldt, C. M. Kleemann, V. N. Mackiw, B. Meddings, J. A. MacLellan, E. J. Maney, D. E. Halter, D. J. I. Evans, J. Honkasalo, M. F. Dufour

The technology employed in the processing of nickel- and cobalt¬bearing ores is a continually developing process for utilizing the natu¬ral resources available at a given time. Most of the nickel-bearing ores of commercial importance fall under the category of sulfide or lateritic ores. While the occurrences of both types of ore have been known for over a century, until 1960 sulfide ores largely from Cana¬dian mines have constituted the chief source of production for com¬mercial nickel in the Free World. Sulfide ores are readily amenable to some physical means of concentration, but the lateritic ores are not readily concentrated, and to compete economically with the sulfide ores only the high-grade garnierite ores (a type of lateritic ore) found in New Caledonia were processed to any great extent up to the time of World War II. Around 1960, a gradual shift in emphasis towards latetitic ores as a commercial source of nickel became evident. See Fig. 1. New discoveries of sulfide ores, usually located underground, had not kept pace with the steadily increasing demand for nickel. The lateritic ores, found near the surface of the earth, could be readily surveyed and mined. Lower cost processes were being developed for the extrac¬tion of nickel from lateritic ores and for a time it appeared that, based on pre-OPEC economics, the lower grade lateritic ores-down to 1.2% nickel-could satisfy an increasing share of the world demand for nickel. The sharp increase in the price of crude oil in the 1970s, however, slapped an economic damper on the lateritic ores. Processes for treat¬ing lateritic ores consume large amounts of energy; as the nickel industry entered its deep recessionary period in 1981-83, nickel pro¬duction from lateritic ores became less and less profitable than from sulfide ores. At the same time, new sulfide ores were discovered and brought into production, notably in Australia and South Africa. The chapters in Section 17 were prepared in 1973. Recent improvements in mineral dressing techniques since then include Inco's pyrrhotite separation process which can remove additional pyrrhotite from nickel concentrates, resulting in reduced sulfur dioxide emission from the Sudbury smelter, by some 25%. The chapters in this section dealing with nickel are grouped in line with the two broad categories of nickel ores. Starting with the description of nickel and cobalt ores in Chapter 2, Chapter 3 deals with the processing of sulfide ores, and Chapter 4 deals largely with lateritic ores. The intermediate steps to produce a suitable feed for extraction plants and refineries are similar in principle within the same categories, although dissimilar between sulfide and lateritic ores. Thus Inco in Canada employs flotation and so does Western Mining Corp. in Australia, although Inco's plants were started before World War II, and that of Western Mining only recently in the late 1960s. Treatment subsequent to the ore dressing step, however, shows some interesting differences. The earlier parts deal mostly with established and conventional practices, leaning heavily on pyrometallurgy in both

Jan 1, 1985

By Charles T. Ford

Limestone neutralization of acid mine drainage has been proposed and examined as a method of treatment. However, limestones differ greatly in composition and structure. The reactivity of 14 selected limestones was studied, both by chemical and physical properties and by relative effectiveness in neutralizing synthetic and actual mine waters, and several guidelines for the choice of limestones are given. The particle size should be at least as small as 74 microns and preferably smaller. The most effective limestones are most nearly pure calcium carbonate. Stones which have a relatively low calcium con- tent but which contain calcite and have a large surface area are also effective neutralizers. Magnesite and dolomitic limestones are much less reactive. Disadvantages in using even the most reactive lime- stones for neutralization are indicated.

Jan 1, 1974

By G. ALBORNOZ, S. Utili, A. AGOSTI, C. VALDERRAMA, R. PELL, N. Morales

The steepness of the slopes of an open pit mine has a substantial influence on the financial return of the mine. The paper proposes a novel design methodology where overall steeper pitwalls are employed without compromising the safety of the mine. In current design practice, pitwall profiles are often planar in cross-section within each rock layer; i.e., the profile inclination across each layer tends to be constant. Here instead, a new geotechnical software, OptimalSlope, is employed to determine optimal pitwall profiles of depth varying inclination. OptimalSlope seeks the solution of a mathematical optimization problem where the overall steepness of the pitwall, from crest to toe, is maximized for an assigned lithology, geotechnical properties, and factor of safety (FoS). Bench geometries (bench height, face inclination, minimum berm width) are imposed in the optimization as constraints which bind the maximum local inclination of the sought optimal profile together with any other constraints such as geological discontinuities that may influence slope failure. The obtained optimal profiles are always steeper than their planar counterparts (i.e., the planar profiles exhibiting the same FoS) up to 8° depending on rock type and severity of constraints on local inclinations. The design of a copper mine is first carried out employing planar pitwalls, secondly adopting the optimal pitwall profiles determined by OptimalSlope. The adoption of optimal slope profiles leads to a 34% higher net present value and reductions of carbon footprint and energy consumption of 0.17 Mt CO2 eq and 82.5 million MJ respectively due to a 15% reduction of rockwaste volume.

Feb 9, 2022

By Brian Y. Lattimer, Kray D. Luxbacher, Jonathan L. Hodges, Manuel J. Barros-Daza

The abstract discusses the importance of predicting fire response time (the time before conditions become untenable for firefighters) in underground coal mines. It introduces a data-driven approach using artificial neural networks (ANNs) to predict response time and fire size based on measurable parameters like CO concentration, mine geometry, air velocity, and time since fire detection. The approach was tested using simulations from fire models (FDS and FSSIM), showing promising results for real-time predictions during mine fires.

Mar 6, 2022

By E. A. Schmidt

The Thompson Creek molybdenum deposit is located in the Bayhorse mining district about 20 air miles southwest from Challis, Idaho. Cyprus Mines Corporation began exploration on the property in 1968, and at present is in the pre-production stripping phase. Molybdenum mineralization is contained in a late Cretaceous (88 m.y.), composite, biotite granodiorite-quartz monzonite intrusive complex. The complex intruded carbonaceous argillites, causing local silicification and hornfelsing of the sedimentary rocks. The main portion of the orebody is unconformably overlain by up to 1000 ft of Challis volcanic rocks (49 m.y.). Surface mineralization and related rock alteration can be recognized over a 2500- x 8000-ft area extending northwesterly from the discovery outcrop. Molybdenite occurs in a network of subparallel quartz-biotite-muscovite-orthoclase veins and veinlets within the intrusive rocks. The zone of +0.020/0 Mo mineralization outlines a steeply dipping tabular configuration in the main orebody and gradually changes into a hood-shaped configuration along strike to the northwest. The overall grade [ ] of the deposit is controlled by the abundance of the quartz- molybdenite veins and veinlets in the intrusive rock. Very little disseminated mineralization is present, and no economic mineralization occurs in the argillaceous sedimentary rocks. Muscovite, biotite, and orthoclase/microcline define three alteration zones that conform with the overall zone of mineralization. The upper alteration zone contains dominantly muscovite and sericite associated with quartz veins. The central zone contains coarse biotite and locally orthoclase and microcline. Molybdenite is most abundant within this zone. The lower zone contains thin envelopes of orthoclase/microcline around quartz veins. Based on about 313,000 ft of drilling and 8592 ft of underground development work, computer-derived ore reserve calculations indicate 200 million tons averaging 0.1 1% Mo at a 0.050/0 Mo cutoff grade with a waste to ore stripping ratio of about 3:l.

Jan 1, 2013

By M. Kuchta, S. Iverson

INTRODUCTION Underground blasting with perimeter damage control can reduce the requirement for loose rock scaling and reduce the risk of ground-fall injury. In previous work Johnson and associates (2010) from the National Institute for Occupational Safety and Health (NIOSH) developed a rock blasting perimeter damage and control design method included in the DRIFT software tool (NIOSH, 2018). This blast design software tool utilizes current methods of perimeter blasting with the added design element of a buffer row coordinated with the perimeter row to reduce perimeter damage. The coordination is in applying one of several blast damage models to calculate the damage extent for each production, buffer, and perimeter blast hole in the design. The calculated damage is then applied to the spacing between the blast holes. The basis of the buffer row design is the practical damage limit and is supported in this paper with computer models of signature blasts and full blast rounds using PFC2D (Itasca Consulting Group, Inc., 2018). Practical damage limit The practical damage radius (Rd) is not to be confused with burden. The practical damage extends toward the burden side and inward into the rock mass. For all practical purposes, the practical damage radius equals one half the burden (Hustrulid, 1999). The maximum extent of radial crack damage around a blast hole (Rcrack) and the maximum extent of radial crushing damage (Rcrush) are illustrated in Figure 1. A “practical” radius of damage (Rd) lies between the crushing limit and the cracking limit. The “practical” damage means that if the rock mass lying outside of the practical damage circle were removed, the rock remaining within the circle would easily break apart. Blast pressure and strain Work by Johnson (2010) and associates included signature blast tests in large concrete blocks. A signature blast test in concrete as exampled by strain data in Figure 2 was designed to monitor the dynamic strain at incremental distances from a single fully coupled blast hole, measure crushing and radial cracking extent by wire-sawing to expose the damage, and relate the damage to overbreak that is common in underground blasting. The initial slope or rising strain as the stress wave passes by the strain gages shows very high strain and strain rates close to the single blast hole. Since the derivative of strain versus time is strain rate, a line tangent to each of the strain curves at the maximum strain rate is drawn and the strain rates identified. The high strain rates were then plotted on a semi-log in Figure 3 to show a linear trend. The strain rate is greater than 100 s-1 within 20 cm from the blast hole for the signature blast experiment. The strain rate at the blast hole wall is projected to be 300 s-1. Figure 4 illustrates the damage in the NIOSH block exposed by a wire saw cut.

Jan 1, 2019

By K. Voltaire

As can be guessed from the subtitle, this paper will not be a geologist's one since neither of the authors is a professional geologist. This will be an attempt to present the various aspects of the legal, economic and geological setting that condition the exploration for mineral resources in Haiti. Before going any further, we are going to present a summary of our rather short mining history. One must first mention that Caribbean countries are latecomers to the mining business. If we go back to the colonial period (1500-1600's) there are net many traces of mining activity. The major exception to this is the recovery of alluvial gold in Hispaniola, Puerto-Rico and Cuba and also some sporadic copper mining in Hispaniola (notably in Haiti near the city of Gonaives and in some of the old mining districts in Santo-Domingo). These easily minable deposits were quickly exhausted and mining quite naturally came to an end as agriculture rose as the major economic activity. In the Haitian case, aside from very erratic mention during the 19th Century of the presence of some minerals (i.e. haematite, chalcopyrite) no serious geological work was undertaken. In the meantime, the peasants in the northern part were still looking for alluvial gold in some streams but this amounted to next to nothing. One has to come to the 1920's to find the first comprehensive geological survey. At the time, the country was under the military control of the United States. The American Government sent a team from the U.S.G.S. headed by the late W. Woodring to initiate a survey of the country. Given the lack of transportation, this was a tough task and Woodring and his associates did it magnificently. As a result of their efforts, they published a very thick piece of work of more than 600 pages. "The Geology of the Republic of Haiti" although it was published in 1924 is a classic and is a must for anyone interested in the geology of Haiti.

Jan 1, 1978

By Subhash Bhagwat

The definition of high sulfur coal varies from plant to plant and is subject to changes over time as the environmental regulations change. In plants constructed after September 1978 hardly any domestic coal can currently be burned without the use of Flue Gas Desulfurization (FGD) because a 70 percent reduction in potential S02 emission is mandatory even when the potential emission is less than 0.6 lbs. S02/106 Btu heat input. On the other hand older power plants may in many cases be legally emitting up to 6 lbs. S02/106 Btu or more under some State Implementation Plans (SIP). The Clean Air Act requires, however, that in the long run all stationary sources of S02 pollution must abide by the ceiling of 1.2 lbs S02/10 Btu heat input.. Coals with a potential to exceed this limit must, therefore, be considered as high sulfur coals. Such a definition would be unrealistic because many SIP's aim at an emission level of 1.8 to 2.0 lbs. SO2/106 Btu for existing plants and a stricter limit does not appear to be enforceable in the next 5 to 10 years. As a result, coals with an emission potential of greater than 2.0 lbs of S02 per 10 Btu heat input are considered high sulfur coals for the purpose of this investigation. Based on this definition and the 1984 actual consumption of coals by U.S. electric utilities (table 1) only 7 out of the 24 coal producing states could be categorized as low sulfur states. The actual S02 pollution, however, depends on total tonnage of a type of coal as well as the sulfur content of the coal. Table 1 also indicates that only 6 states--Kentucky, Illinois, West Virginia, Ohio, Pennsylvania and Indiana-account for nearly 77 percent of S02 pollution potential in the United States. These 6 states have therefore been chosen for this investigation.

Jan 1, 1986

By Wi11iam I. Rau

In the recent past, there have been numerous projections of the capital requirements for additional mine and processing capacities required to meet projected non-Communist World demand for minerals during the next 15 years. A consensus judgment would place this currently at $14 billion per year, with the largest expenditures required for increases in aluminum and copper production. Until the early 1960's, internally-generated funds were used to finance nearly all of the large mining companies' capital requirements. Since that time, however, the enormous rise in capital costs per ton of capacity, accompanied by a relative stagnation in the financial condition of many of the major mining companies, have caused the industry to begin to rely on external sources for financing expansion, most notably, project financing from commercial banks and institutional lenders. Seven years ago, it was estimated that, in the five-year period 1972-1976,75% of the total capital requirements of $10 billion per year would come from internally-generated funds, with the balance to be provided by external sources. l There is little doubt that, in the coming decade, the percentage provided by external sources will increase. There should be little difficulty in attracting the necessary investment within the U.S., particularly in view of the expected continuation of companies in diverse industries seeking mineral investment opportunities. The more serious question lies in the availability of internal and external sources of funds for the expansion of mine and smelting capacity in the developing countries. Recently, R. F. Mikesell projected an average annual investment of $4.7 billion (in 1977 prices) over the 1977-1990 period would be required from foreign sources to expand metal producing capacity in developing countries. Mikesell concluded that external financing of mining projects in the developing world has averaged less than $1 billion per year over the past five years, thereby requiring more than a quadru1ling of this financing over the next decade.

Jan 1, 1978

By Willard A. Murray

Investigations of underground coal gasification (UCG) have been actively pursued by the U.S. Department of Energy since the early 1970s; the Lawrence Livermore National Laboratory (LLNL) conducted its first field experiment in 1976. In conjunction with field and laboratory studies of the process of recovering energy in the form of combustible gases from the burning of coal deposits in situ, an extensive environmental program has been conducted at LLNL under the auspices of both the U.S Department of Energy and the U.S. Environmental Protection Agency. This program has focused on groundwater contamination and ground surface subsidence caused by UCG. The material presented in this paper is a summary of results from the environmental studies of UCG activities conducted by LLNL as described by Mead, et al. (1981) and supplemented by the author's experience with this program in 1980 and 1981. The following sections describe specific groundwater contamination and surf ace subsidence problems encountered during field experiments in northeastern Wyoming, and will discuss implications of preliminary findings from geohydrological investigations of a site in southwestern Washington. Finally, possible environmental control technologies that are most likely to be effective in mitigating environmental impacts are discussed.

Jan 1, 1982

By Thomas R. Mills

Under contract to the U. S. -Department of.-Interior, Office of Coal. Research, Hittman Associates recently completed a 16 month study of the environmental-factors associated with the siting, construction, and operation of large scale coal refineries. (1 2) The objective of this study was to identify the environmental effects and to define the magnitude of these effects and the approaches which can be used to mitigate adverse environmental consequences. In the study, the five major coal bearing regions of the country having reserves adequate to support a number of large scale coal refineries were considered. The regions considered were: Appalachian, Eastern Interior, Fort Union, Powder River, and Four Corners.

Jan 1, 1975

By J. Moreno, P. Sinton, R. Vogwill

Groundwater models are often used to estimate and optimize dewatering volumes, locations, and rates, as well as water supply production for mines. Projections of future dewatering rates are highly dependent on the conceptual model used as a basis for the numerical model. This paper presents an audit of modeled dewatering rates versus actual data and an examination of the reasons for discrepancies. The setting is a gold mine, in a collapsed volcanic crater breached by the ocean. Mining will occur to about 200m below sea level within a remnant geothermal system. A detailed hydrogeologic investigation was conducted to support dewatering and geothermal depressurization design for future mining operations. These studies included development of a three-dimensional model to simulate heat and density-coupled groundwater flow, using software developed by URS/Dames & Moore. In this paper, the groundwater modeling results prepared 5 years previously by others was audited using operational dewatering data. This model was then extended using the new mine schedule and actual locations for dewatering wells to predict future dewatering requirements. In addition, a comparison between “hot” and “cold” model results was used to assess the need to use a heat-coupled versus a standard groundwater model for dewatering predictions.

Jan 1, 2003

By S. W. Swerson

25.1-THE PURCHASING FUNCTION The word "purchase" means "to obtain by paying money or its equivalent; to buy for a price." This definition is all too simple to properly describe the word "purchase." One should consider the basic objectives of purchasing-to procure the right material, of the right quality, in the right quantity, at the right time, from the right source at the right price. Here it should be noted that in purchasing it is recognized that the best quality or grade of materials may be the only grade to purchase for a specific job. However. there are other jobs where the medium, standard or poor grade of materials is the right quality, as the lesser grade will perform adequately. In some companies the purchasing department makes the decision as to the right quality, while in others this function is shared with other departments. Years ago in business the owners controlled the purchase of required materials and services by performing the task themselves. The owners obtained the type materials and services they needed and paid the price demanded. This function was performed along with all the other duties requisite to the owner of a business, namely, clerical, hiring and firing of personnel, financial management, etc. As companies grew in size, clerical duties were assigned to other personnel and the owner usually retained the purchasing authority. This same division exists today in many small companies. However, in the larger companies, the purchasing function has been assigned to a separate department. Importance of the Purchasing Function-The larger the company the greater the need for a separate department to perform the purchasing function. Today the purchasing agent for most larger progressive companies has policy-making authority. Therefore, it is reasonable for him to report to the president or vice president of his company. Studies of several hundred companies show the purchasing department head reporting to top management 87.5% of the time. This fact in itself is indicative of the importance of recognizing that effective purchasing requires a department staffed with specialists in the field of procurement with knowledge and experience sufficient to perform in today's complicated business atmosphere. In the manufacturing field it is a common axiom that usually half of the company's dollar expenditures are for materials. If a purchasing department functions properly and the right product or service is purchased and a savings accomplished, the contribution to the net profit of the company is much greater, as opposed to sale dollars, where only a small portion is carried forward to net profit. Thus, the importance of the purchasing department and the contribution it can make to the success of the company are obvious. 25.1.1-THE PURCHASING FUNCTION 1. Responsibilities of the Purchasing Function-The purchasing agent in performing his function is acting as an agent for the company and others in the company requiring materials and services. Accordingly, he is committing the company for materials and services subject to certain limitations which the board of directors may stipulate. In view of this, the purchasing agent,'^ position must of necessity he

Jan 1, 1973