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By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

By Ryne Goodnough

INTRODUCTION The Wyodak Mine is located in the northern Powder River Basin of Wyoming and is thought to be the oldest continuously operating coal mine in the United States, having recorded yearly production since 1923. The mine has expanded over its 80-plus-year history from an underground room and pillar operation to a modem surface snip mine. Currently, the mine feeds four on-site power plants using conveyors, has a filly automated truck load-out, and ships coal by rail to nearby Wyoming power plants. Annual production of 4.8 MTFT is accomplished on a single-shift basis with 53 fill-time production, maintenance and administrative personnel. In 1990 a decision was made to modernize the coal feed system to the power plant complex with modular conveying equipment. The project involved constructing coal storage silos, overland conveyors, feeder breakers and modular conveyors to follow the mining benches. Total project cost in excess of $25 million was used to construe or commission: A 2,150-TPH 48-inch overland conveyor with horizontal curve A 2,150-TPH 48-inch plant feed conveyor 13 shuttle conveyors, each 150 feet in length 2 front-end loaders equipped with large coal buckets 2 crawler-mounted feeder breakers with compatible discharge conveyor 1 high-lift shuttle conveyor with adjustable discharge height 1 high-angle conveyor 2 4,500-ton coal silos with automated truck load- out capabilities. Conceptually, the plan was to develop 33 MTONS of reserve adjacent to the plant complex and use front-end loaders to tram ROM coal to in-pit feeder breakers. The primary crushing would reduce coal to 6-inch minus and feed through a series of shuttle conveyors to an overland supply conveyor. The overland conveyor would feed the secondary crusher, and the crushed product would be moved to any of four storage silos using conveyors. The in-pit crushing and conveying system is a unique fit for the power plants and provided many interesting challenges for the coal mine to overcome. System integration, maintenance windows, equipment availability and the like had a major role in engineering the system to a power plant complex having no emergency coal stock- pile and limited plant storage. In crediting the designers and operators, the in-pit conveying system at the Wyodak Mine has delivered coal to the power generators since 1994 with little or no downtime. ENGINEERING CONSIDERATIONS Yearly production for the mine is based on 3.5 MTPY which requires about 1,550 TPH on an 8 hour, one-shift- per-day, 5-days-per-week basis. Considering efficiencies for starting and stopping, crew transport, loader trams, and operational delays, a nominal conveyor feed rate of 2,150 TPH or 72% system utilization was used in engineering design application. This ton-per-hour basis is used to size and commission the remaining components of the in-pit system. Coal quality requirements also played a role in equipment selection. The mine reserve grades vertically and coal from Upper and Lower benches needed to be mixed in any given proportion. This required equipment to be filly variable and capable of producing all or part of the 2,150-ton-per-hour requirement. Bisecting the 80-foot- high coal face for quality required construction of a high lift module capable of reaching a single overland conveyor 40 to 50 feet above the pit floor. Designers built a high-lift, adjustable and portable module specific for this application Having two bench systems capable of each producing the required feed for the power plants required computerized control processes to control on-line production. RSLinx, RSLogix, and RSView were chosen as the communication and control software, and a Data Highway Plus serves as the communication backbone. Each of the two in-pit benches are controlled with a data center equipped with Allen Bradley 5-40 processors. This

Jan 1, 2004

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

7.1 INTRODUCTION Subsidence measurements include surveys of surface and sub- surface movements and measurements and monitoring of surface structural damages. In this chapter the most commonly employed methods are discussed. 7.2 MEASUREMENT OF SURFACE MOVEMENT Surface movements involve two basic components, subsidence and displacement. The other five components (i.e., slope, curvature, strain, twisting, and shearing) are derived from the two basic components. Since all measurements are geometrical, they can be measured by common surveying techniques. 7.2.1 MONUMENTS The first step is to select the types of monuments or surveying stations. Various types of monuments have been employed. The simplest one is a paint mark on the pavement or concrete slab, although this should be discouraged because it is easily affected by artificial or environmental factors. Others use steel or wooden rods, 2 to 3 in. in diameter or square by 2 to 5 ft long, driven all the way down, flush with the ground. Both types of-monuments have no provision for frozen ground heave and both are tampered with easily. A more elaborate type is shown in Fig. 7.1 (O'Rourke et al., 1982). A rebar or iron pipe, 2 in. in diameter by 6 ft long is driven down to nearly flush with ground surface. The top 4-ft section of the hole around the pipe is enlarged to 1 ft in diameter and grouted with concrete. The most sophisticated type of monument is shown in Fig. 7.2 (Conroy and Gyannaty, 1983a). It consists of two hollow steel pipes. The inner I-in. pipe, 10 ft. long is enclosed by a 2-in. pipe at the upper 5-ft. interval. The gap between pipes is filled with foam to isolate the inner pipe from external environmental effects. The bottom of the inner pipe can either remain open or be sealed with a cap, depending on whether the ground is dry or wet. 7.2.2 MONUMENT ARRANGEMENT AND SURVEY LINES The monuments are installed strategically to form survey lines. Three basic types of monument arrangements are shown in Fig. 7.3. Figs. 7.3a and 7.3b are for single and multiple panels respectively; Fig. 7.3~ is mainly for research purposes with multiple survey lines both in longitudinal and transverse directions. This is extremely useful for defining the distributions of ground deformation, as shown in Figs. 2.10 and 2.11. Basically there are two types of survey lines: the whole basin survey line which extends across the panel width or length (e.g., lines R, R, in Fig. 7.3 a&b), and the half basin survey line which covers approximately half of the panel width or length (e.g., lines R, F in Fig. 7.3 &b). Survey lines are generally straight and linear. They should be located on the major crosssections of the subsidence basin. Survey lines must extend beyond the edge of the opening for a sufficient distance so that the edges of the subsidence basin can be determined accurately. The minimum distance from the edge of the opening must be [ ] where h = mining depth, [ ] = angle of draw, and [ ] = 5 to 7 times monument spacing for whole basin surveys, or 8 to 10 times the monument spacing for half basin surveys. There must be at least two controlling monuments at each end of each survey line. Controlling monuments must be located beyond the influence of underground mining. The spacing of the monuments ranges from 15 to 200 ft. The UK National Coal Board (NCB, 1975) recommended a spacing equivalent to 0.20 of the mining depth. Peng and Geng (1982) recommended that the maximum monument spacing be dependent on mining depths as follows:

Jan 1, 1992

By Leonard Harris

Good morning ladies and gentlemen. I would like to thank Hans von Michaelis for inviting me here today to talk to you on the subject of "How to Reduce Capital and Operating Costs at Heap Leaching Projects". The comments I have to offer are based on my experiences in gold heap leaching by Newmont Gold Company in Nevada, by Newcrest in Western Australia (formerly Newmont Australia), by Minera Yanacocha in Peru, by Penmont in Mexico, by Z­Mont in Uzbekistan, by the Hochschild Group in Peru and expe­riences in copper heap leaching by Magma Copper Company in Arizona and as a consultant on several other smaller projects worldwide. These comments are a mixture of technology, ideology, philos­ophy, strategy, logic and opinions and might be looked at by many of you as "one man's view". I like to look at them as "a view from the bridge". Let me add that it doesn't bother me that you may not agree with my comments and that I would be interested in your comments or criticisms at the end of this presentation. I don't intend to discuss the details of the mineralogical and metallurgical test work. I will assume, for this talk, that the test work has been conducted very carefully and that the ore in ques­tion has been determined to be heap or dump leachable. Let's begin with a couple of don'ts! ! First, "don't build the mill until you have a mine". This might sound to many of you as fundamental but believe me it continues to happen. The minable reserves should be well known and established as being economical to mine and leach. NPVs, ROis and IRRs are fine if you can be sure of the gold price you use in your calculations. The best method is to carefully estimate the capital cost and, most importantly, the operating cost per ounce of gold recovered which should include the repayment of the borrowed capital and the interest thereon. For example, if the operating cost under these conditions is $200 per ounce, you have a project. If it is $400 per ounce, walk away f om it. Incidentally, the same prin­ciple can be applied to projects involving other metals such as silver, copper, lead, zinc, etc. The second don't is - "don't be sheep". Don't climb on the bandwagon with respect to design and operation because every­one else does it a certain way. Keep an open mind, be versatile, look for change, look for every possible innovation for reducing capital and operating costs. Beware of those who will tell you "we tried that it won't work" or "it's too expensive" or "that's a crazy idea" , <' ''what do you know about mining - you are a metallurgist aren't you" or "what do you know about metallurgy - you are a miner aren't you" or "only geologists can handle this issue". Be careful about pigeon-holing people according to their basic degree and experience. In these days of low metal prices, we need to consider everybody's ideas, as crazy as some of them might seem.

Jan 1, 1998

By V. Rajaram

V. Rajaram&apos;s paper does not accurately reflect the status of Eastern oil shale development as it stands today. Today, there are numerous firms and research organizations pursuing development of the Eastern oil shale resource. Details of the current projects can be found in the symposium proceedings for the 1985 Eastern Oil Shale Symposium held in Lexington, KY, in November 1985. Additionally, the comparison of the Paraho and HYTORT processes is invalid because the HYTORT process design used it not an accurate representation of the HYTORT process. I will briefly address this point. It is disconcerting to see the author reference hydroretorting work conducted at the Institute of Gas Technology (IGT) (circa 1978) in the introduction, which indicates hydroretorting (the HYTORT process concept) yields of up to 250% of the Fischer Assay oil yield, and then turn around and base the remainder of the paper on the 1981 "Synthetic Fuels From Eastern Oil Shale Report," which depicts the HYTORT process as yielding only 115% of Fischer Assay. Indeed, work conducted at IGT to 1981 indicates commercial oil yields for the HYTORT process would be in the range of 200% to 210% of Fischer Assay. A more thorough literature search by the author would have yielded a 1981 publication by Schora et al., entitled "Update on the HYTORT Process" presented at the Symposium on Synthetic Fuels From Oil Shale in Nashville, TN, October 1981. This paper thoroughly discusses the shortcomings of the referenced study with respect to the HYTORT process. For the interested reader, I can also recommend a June 1984 report entitled "An Economic Comparison of Five Process Concepts for Using Eastern Oil Shale," published by Los Alomos National Laboratory (LA-10106-MS, Parkinson et al., NTIS No. DE84-015981). This study applies a realistic HYTORT design (circa 1982) and indicates that HYTORT is the most econom¬ical method of processing Eastern oil shales (about $20/bbl less than Paraho). Additionally, completion of the HYTORT feasibility study in June 1983 by Phillips Petroleum Co., Bechtel Group Inc., HYCRUDE Corp., and IGT has resulted in development of designs that will make the use of HYTORT on Eastern oil shales even more economic. HYCRUDE Corp., a second-tier subsidiary of IGT, is involved in commercialization and licensing of the HYTORT process and is active in efforts to develop the Eastern oil shale resource. As such, I am disappointed to see outdated material that has the potential for adversely affecting development of Eastern oil shale resources published in MINING ENGINEERING. I can only hope that this response will serve to mitigate any negative opinions that may have been formed. Today, activity in commercializing Eastern oil shales dominates the oil shale scene in the US and indicates a realization of the potential for these shales. Reply by V. Rajaram I appreciate the updates on the HYTORT process provided by R.C. Rex, Jr. I hope that Hycrude&apos;s work will lead to the funding of a pilot plant to demonstrate that yields of greater than 200% of Fischer Assay can be achieved in a commercial plant using eastern US oil shales.

Jan 1, 1987

By William A. Vogely, Robert T. Deacon, Walter J. Mead

Oil and gas together are the most important energy sources consumed in the United States. In 1983, 67% of total energy consumption was represented by these two resources (43% was oil and 25% natural gas), while coal accounted for only 23% of U.S. energy use. All other sources (hydroelectric, nuclear, geothermal, wood and other waste products) accounted for only 10% (U.S. Department of Energy, 1983, p. 1). In this chapter we address the issue of regulation in the oil and natural gas sectors of the U.S. energy industry. In addition, we analyze the cartel characteristics of the Organization of Petroleum Exporting Countries (OPEC). If OPEC is an effective cartel, then U.S. oil as well as substitute energy prices, production, and consumption will be affected by cartel action. Further, as the nation learned in the 1970s, major changes in oil prices produced major disturbances in other sectors of the U.S. and world economies. Consequently, it is important to determine whether the roughly ten-fold increase in nominal prices that occurred in the 1970s was the result of cartel action, the result of normal economic forces adjusting to the fact that the world was running out of cheap oil, or some combination of the two. Furthermore, the presence of OPEC- associated oil price increases is related to regulations that were imposed on the energy sector during the decade of the 70s. THE RATIONALE FOR GOVERNMENT INTERVENTION Economists have traditionally judged the state of an economy to be efficient if it is impossible to improve the welfare of any one member of society without harming someone else. Any economy that satisfies this condition is said to be Pareto efficient, after the Italian economist Vilfredo Pareto who first gave rigorous attention to such questions (see Chapter 4.10). It is easy to see why an economy that fails to satisfy this condition could not be efficient; by definition, it would be possible to alter the allocation of resources in a way that improves everyone&apos;s welfare.&apos; Much of the appeal of this weak efficiency criterion is that it avoids comparisons of levels of well being among different individuals. The cost of such generality, however, is that in any given economic system a wide array of different Pareto efficient states will typically be attainable. In the simple Robinson Cursoe economy of the undergraduate textbook, for example, efficiency could just as easily be attained in the case where most of the island&apos;s resources are owned by Friday as in the case where they are under Crusoe&apos;s control. However, the distribution of welfare among the two inhabitants would obviously be quite different in the two situations. Such equity issues cannot be settled on objective or scientific grounds. To decide which distribution of welfare is socially "best" an explicit value judgment is needed. It is useful to keep in mind the preceding distinction between efficiency and equity issues when analyzing the role of government in a market economy. Decentralized market processes will not necessarily result in a distribution of welfare that society considers just. For this reason a potential role for government lies in the redistribution of income among individuals. In some circumstances, however, market pro-

Jan 1, 1985

By Sadatoshi Ohmori, Karen Shinoda, Nobuharu Isago

Method of tunnel excavation is generally classified such as full-face excavation, bench-cut excavation, diaphragm with center pillar excavation, center drift advancing excavation and so on. These excavation methods should be selected properly in terms of ground condition, the area of tunnel section, the arrangement of construction. Center drift advancing excavation method in mountain tunnel construction, is often applied under quite weak ground condition, prior to the main tunnel widening excavation. Recently, the advancing drift is for the excavation of extreme larger section in main tunnel to confirm the stability of face. In addition, we make advancing deformation occur in drift when the construction was done in poor ground like squeezing and swelling ground, and the force occurred in main tunnel was expected to reduce by the effect of ‘dodging’ for support structure. Mechanical relation between drift and main tunnel has been mainly grasped on past experiences in Japan. However, the characteristic of tunnel enlargement was partially discussed by using the three dimensional analysis and the result was grasped by past experiences on the basis of each tunnel construction. It is considered that the natural ground stress is normally released as the advance displacement, but it was released after the main tunnel widening, causing a displacement and deformation increase. Up to now, the construction practice has shown the reduced displacement during the main tunnel widening excavation by the drift, and the numerical analysis agreed with above observation. However, there is no report found that the displacement and deformation at the time of widening excavation of the main tunnel become larger in case of a drift application. In Kyusyu area, western part of Japan, the northern part of Yoshinomoto tunnel was applied an advancing bench-cutting method for a center drift for the purpose of obtaining the geological information for support selection of the main tunnel in widening excavation. The shape of center drift was a perfect circle with a diameter of 4.0 m or 6.0 m and the same temporary support regardless of the geological situation. As a result, the displacement and deformation of the main tunnel after widening excavation were larger than the section where the center drift was not applied. Furthermore, the displacement and deformation of the main tunnel after its widening excavation was considered larger in the section where the cross-section of the center drift with 4.0 m diameter was applied than one with 6.0 m. Therefore, the center drift with a high rigidity support structure had the effect of suppressing the advancing deformation prior to the main tunnel widening excavation.

Jun 13, 2021

By R. C. Gilmore, A. Juganda, Jr. Bogin, S. Saki, S. A. Lolon, J. F. Brune, Jr. Zipf, J. W. Grubb

"Methane-air mixtures form explosive gas zones (EGZs) in underground longwall coal mine gob ventilation systems. Such EGZs have caused a number of fatal mine explosions, including the 2010 disaster at the Upper Big Branch mine, where 29 miners lost their lives. Researchers at the Colorado School of Mines use Computational Fluid Dynamics (CFD) modeling to predict EGZ locations. In this paper, a model of a bleeder-ventilated gob system includes two adjacent panels, an active longwall and a mined-out gob. Boundary conditions for the simulations match statutory ventilation conditions for methane concentrations and air flow rates at common measurement points throughout the mine. The modeling results predict the persistence of EGZs in the gob that may compromise the effectiveness of the bleeder ventilation system. INTRODUCTION Mine operators must maintain statutory air flow and air quality requirements at various locations throughout the mine ventilation network. U.S. regulations generally require mine operators to establish bleeder ventilation systems surrounding longwall gob areas (30 CFR §75.334.) Gilmore et al. (2015) documented that explosive gas zones (EGZs) exist within bleeder-ventilated gobs and may extend into the bleeder entries that must be examined weekly, presenting an immediate fire and explosion hazard to mine examiners. In a project funded by the National Institute for Occupational Safety and Health (NIOSH) under contract 200-2009-31409, researchers at Colorado School of Mines have developed a tool for modeling gob air quality using Computational Fluid Dynamics (CFD). This CFD model is capable of examining a wide range of longwall mine ventilation systems. Marts et al. (2014a, 2015) studied progressively sealed gobs with U-type ventilation and using a back return. Saki et al. (2015) examined the impact of gob ventilation boreholes and face air quantity effects on tailgate methane concentrations. Gilmore et al. (2014) modeled bleeder-ventilated gobs and Gilmore et al. (2015) studied the effects of bleeder regulator settings on EGZ volume and locations in the gob. CFD studies for progressively sealed gobs were generally carried out using 5000 ft (1500 m) panels because the gob atmosphere remains largely unchanged from about 500 m (1,500 ft) inby the face. A modular meshing methodology (Gilmore, 2015) permitted researchers to extend bleeder models to 10,000 ft (3050 m) to include the full extent of bleeder entries. This paper further expands the bleeder CFD model to study the gob atmospheres of two adjacent longwall panels, each 10,200 ft (3100 m) in length, with the active face located 6,700 ft (2040 m) from the startup room."

Jan 1, 2016

By Robert B. Fulton

Staurolite, an iron aluminum silicate mineral, is used industrially as a high value-in-use sandblasting agent, as a premium-grade foundry sand, and as the source of aluminum in portland cement manufacture in areas where the aluminum constituent is not economically available from shale or argillaceous limestone. Staurolite is produced by E. I. duPont de Nemours and Company as a coproduct of titanium mineral production in Florida, where it was first recovered on a commercial scale in 1952 from the Pleistocene sand deposit known as Trail Ridge. At this deposit it comprises about one-fifth of the heavy mineral suite, the principal other heavy minerals being ilmenite, leucoxene, rutile, zircon, kyanite, sillimanite, and tourmaline. Staurolite and other nonconductive silicate heavy minerals are separated from the conductive titanium minerals on high-tension electrostatic separators. The silicates are then separated from one another magnetically, the staurolite being relatively magnetic. In the total process, 200 t of raw sand yield 7 t of heavy mineral concentrate from which about 1 t of staurolite is recovered. The staurolite concentrates have the characteristics given in Table 1. Staurolite concentrates are sold in three grades having the size distribution of subrounded grains shown in Table 2. All three grades are used in sandblasting for cleaning metal, for paint removal, and for cleaning masonry buildings. Staurolite compares favorably with lower priced abrasives, such as boiler slags and silica sand, in applications where the abrasive is used only once and is not recovered and recycled. Because it is hard, fine, clean, and dense it can reduce blasting time, and less material is required to clean a specific area. Compared to higher priced abrasives such as aluminum oxide and garnet, which are recycled, staurolite costs significantly less per kilogram and cleans at equal or faster rates. Its breakdown factor, which is a measure of recycling performance, is intermediate between those for aluminum oxide and garnet. [ ] [Tables 3 and 4] show the blast cleaning performance of the three staurolite grades compared to other abrasives. [Table 5] compares the longevity of premium priced abrasives during recycling. [Table 6] shows a cost comparison between silica sand and staurolite. Fringe benefits claimed for staurolite&apos;s use as an abrasive are: 1. Good flow properties and freedom from the lumps or slivers found in silica sand or boiler slag abrasives. 2. Low dusting affords good visibility, permits faster working, and reduces cleanup problems. 3. Heated storage is not required during freezing conditions, because the mineral is nonhygroscopic.

Jan 1, 1994

By Geno Saccomanno

INTRODUCTION Cancer of the lung in uranium miners has been reported by many authors. In fact, it has become a most popular subject for comment by environmentalists, health physicists, and newspaper writers. I believe it appropriate for me to discuss cancer of the lung in uranium miners in a general way commenting briefly on etiology, incidence, tumor cell types, and particularly, tumor development and treatment. Cytological development of cancer in its incipient stages, particularly the in situ stage, relates to potential cure with photoradiation, early diagnosis, and treatment. Our observations of the cytological response to this modality of diagnosis and treatment will be discussed briefly since this area will also be covered by Doctor Balchum. The cause of lung cancer in uranium miners is primarily due to two carcinogens; cigarette smoking and radon daughter radiation. Cigarette smoking has continued at the same intensity over time among uranium miners; about 60% of the miners are cigarette smokers and 40% are non-smokers. Radon daughter exposure to the uranium miner has varied from mine to mine, but over the last 20 years has been reduced to a minor fraction (.3 WL) of the level that prevailed in many of the mines during the 1950-1960 era when many of the miners were exposed to more than 10 WLM. Sputum samples have been examined on as many uranium miners as responded on the Colorado Plateau since 1956. Autopsy studies were done whenever this was permissible. In some instances, only tissue from a biopsy of a node or lung was available for study. Smoking histories were also taken. Approximately 200,000 sputum samples were examined and histological examinations of lung lesions were made from 312 cases. METHODS The method of sputum cytological collection and preparation was reported previously. I The lungs were fixed with formalin and prepared for sectioning by Dr. Oscar Auerbach with his block-section method. RESULTS Three-hundred-twelve lung cancers from uranium miners were examined. The tumor types and the age at death for each type, smokers and non-smokers, are given in Table #1. The incidence of the tumors in the smokers and non-smokers is given in Table #2. The incidence of small cell carcinoma of the lung has decreased from an incidence of 76% of all the lung cancers found in uranium miners to 22% of these lesions (Figure #1). The decrease in incidence of this tumor is probably due to decreased exposure to radiation in the mines.

Jan 1, 1981

By Dr. -Ing. Rainer Alles

Introduction Belt conveyors usually run in a straight line. The correct alignment is a major criterion for reliable functioning and optimum running properties. Any misalignment of the conveyor, especially mistracking of the belt, can have many negative effects. As a consequence of this, damage can occur on the belt and conveyor, some material can he thrown off at the edges and serious running problems can arise. In the worst case the belt can come off its supporting structure so that operations are brought to a halt. The absolutely straight alignment of a belt conveyor is actually impossible so that on closer observation every belt conveyor contains more or less sharp curves. Whereas on non-movable conveyors the deviations from the straight can he kept within the range of practical measuring accuracy by careful installation, movable conveyors in surface mining or belt systems in difficult terrain often contain curves that are easy to see with the naked eye. Below ground, too, alignment cannot always achieved to a desirable extent or is changed by geological faults over the years. Despite these unwanted curves, most belt conveyors perform their function reliably. In view of the belt tension, the belt tends to take the easiest path. This in general is the shortest distance between the pulleys on which the belt is tensioned and bent. This means for example that for a curve radius of 10,000 m over a flight of 100 m the greatest deviation from the ideal line is as much as 12 cm. And a figure of this magnitude cannot normally he accepted. In such cases, use is made of guide devices to properly train the belt if it doesn&apos;t center itself anyway. If it&apos;s possible to control the belt position over such unwanted curves. it can also he expected that the belt tends to follow the desired curved path within certain limits. I&apos;d now like to cite a few examples to illustrate the ways and means of guiding a belt conveyor in curves. Cost-effectiveness The advantages of belt conveyors, in particular for longer conveying distances and mass flows have been known and undisputed for some time. However, there are often obstacles standing in the way of the concept of straight routing. If a belt conveyor is to bypass elevations or existing buildings or to follow the seam of underground deposits or a predetermined path, then substantially tighter radii and longer curves are desirable or necessary.

Jan 1, 1993

By F. N. Earll

4.1 – DEFINITIONS 4.1.1-MINERALS, ROCKS, ORES AND NONMETALLIC MINERALS F. N. EARLL A mineral is a naturally occurring inorganic element or compound. Most minerals are distinctly crystalline in habit, and many commonly occur in well-formed crystals or crystalline aggregates, but a few species are characteristically amorphous or noncrystalline. Nearly 2,000 mineral species are now recognized but, perhaps fortun¬ately, a rather large proportion are rarely encountered. Even so, the application of modern mineral-detection and identification techniques reveals the presence of surprising numbers of mineral species, even in deposits of outwardly simple mineralogy. A rock is an aggregate of grains or crystals of one or more minerals. Some rocks are essentially monomineralic. Thus, limestone consists largely of the mineral calcite, and many sandstones are dominantly quartz. Most rock varieties, however, contain a number of mineral species, and their classification or identification depends upon the minerals present, their relative abundance, and such factors as rock tex¬ture and origin. Ore is defined as an accumulation or deposit of one or more minerals that can be exploited at a profit. The profit criterion commonly is taken somewhat liberally, however, for, although always present, the specific mineral content that constitutes a viable product is transitory. The term ore, usually, but no longer consistently, is restricted to the description of mineral deposits that are of value for their metal content. Nonmetallic deposits, although similar to metallic ores in many respects, generally are valued for the minerals or compounds present, as with rock salt (halite), rather than for individual contained elements, though such minerals as elemental sulfur are of much value. 4.1.2-ROCK TYPES F. N. EARLL Rocks arc grouped tinder three broad headings: (1) igneous rocks formed by the crystallization of magma; (2) sedimentary rocks formed by surface processes of weathering and deposition; and (3) metamorphic rocks, which may have been derived from either of the foregoing, but have undergone important physical and/or chemical changes through the application of heat or pressure, or both. The three groups can he further subdivided for convenient reference. Thus, igneous rocks are described as being abyssal or plutonic if formed at great depth, hypabyssal if formed at intermediate depth, and extrusive or effusive if formed at or on the earth&apos;s surface. The sedimentary rock group is similarly subdivided into clastic sediments and chemical precipitates. The clastic sediments form by deposition of preexisting minerals that were released from their place of origin by weathering and were transported to their point of deposition as discrete grains

Jan 1, 1973

By C. I. Walker

"Understanding the influence of slurry particle size on wear is essential to maximising the life of centrifugal pumps operating in mineral processing plants and for hydrotransport duties. Particle size influences not only the wear rate but also the wear pattern in pumps. Complicating the situation is the different relative wear rate for different pump parts depending on particle size. A simple Slurry Wear Type (SWT) number based on particle size and size distribution is proposed as a means for categorising slurry pump wear patterns and optimising pump application and material selection. INTRODUCTION The wear of centrifugal slurry pumps in mineral processing and hydrotransport can be quite severe and is one of the major cost concerns of plant operators. Some typical worn slurry pump parts made from high chrome white iron are shown in Figure 1. Erosion wear rates of up to 2 mm/day are not uncommon [1].Centrifugal slurry pumps are required to handle solid-liquid mixtures with a very wide range of particles extending from sub-micron size clay through to 200mm and larger cobbles. Particle size distribution can also vary widely from uniform and narrow (such as graded sand) through to very broad (such as grinding circuit slurries). Slurry particles may also undergo size change due to attrition during the pumping process. The major wearing parts in a slurry pump are the impeller, side liners and casing liner. These are shown with reference to a Warman® MC centrifugal pump in Figure 2. While many particle properties (including size, shape, hardness, density and concentration) influence pump wear, this work will consider only the effect of particle size as it is the major determinant of wear characteristics. Particle size influences the wear pattern, material performance and relative component wear rate in centrifugal slurry pumps. SLURRY WEAR TYPE (SWT) BASED ON PARTICLE SIZE Soil and grain size classification systems abound. Many have different limits to the ranges for a particular classification, albeit not too"

Jan 1, 2018

By C. S. Caldwell, L. C. Thompson

Mine closure and reclamation issues have evolved over the past several years becoming further complicated by stricter standards and by the cmrent economic climate and price of gold. Many mine permits are calling for more stringent remediation of spent ore heaps and tailings. Leachate solution water quality must meet strict discharge criteria for multiple parameters besides cyanide at mine closure. Closure standards are more difficult to meet and most mines do not have an increased budget at the end of mine life to deal with the falling standards. These issues make additional gold recove1y at the end of mine life a critical problem in financing some closure operations. Another issue is the length of time that must be budgeted for heap rinsing operations. Innovative biological processes developed and applied by Pintail Systems, Inc. at numerous mine closures have indicated that bio-processes can significantly enhance gold recovery dur­ing heap cyanide bio-detox over the recoveries expected from rinse operations. If heap rinsing can be eliminated or reduced and if gold recovery can be improved at the end of mine life, many mine closures may be possible with processes that pay for themselves or generate profit. Additional advantages of the bio­detox technology is that it can be engineered to treat multiple contaminants besides Weak Acid Dissociable (WAD) cyanide. Other contaminants treated in heap process solutions include nitrates, ammonia, heavy metals, thiocyanate and total cyanide. At the Cripple Creek and Victor Mining Company Ironclad Mine near Victor, CO, 5 million tons ofrinsed, heap leached spent ore were biologically detoxified to reduce cyanide content, enhance gold recovery during bio-detox and to reduce dissolved metals in heap leachate solutions. During the bio-detox treatment period a substantial improvement in gold recovery allowed the detox oper­ation to operate as a profit center. Final gold recovery was at least twice what was expected from a normal rinse operation.

Jan 1, 1998

By Ken Deter

My recent experience has been with wet tropical environments, which address a different set of concerns than Nevada, for example. Spraying rather than drips are required in most cases just to assist the water balance requirement, and my comments address those situations. Most tropical operations are faced with both extremely wet and very dry seasons, with some having multiple cycles during the year. A heap leach operation must address the rate of solution building up within the ponds and how long the volume must last until the next rainy season. It may be easier to construct sufficient pond volume to permit yearlong operation than by discharging excess solution now, and searching for water later. Pond capacity requirements can help with the following: Consider using the solution that could be locked up in the pads during leaching to assist water balance problems. Spray excess solution continuously onto spent ore that would not be normally be sprayed again to "soak up" the excess solution, and stop spraying when the water balance is low, allowing the solution in the spent ore to return back into the system-with or without cyanide added. In a multiple lift system, if intermediate HDPE liners are planned for solution grade control, have drips placed before the HPDE liner is in place. The drips can distribute excess precipitation throughout the lower lifts to control the water balance, while the most current lifts are being leached and sent directly for the preg pond. With only 2 million tonnes (2.2 million tons) of spent ore involved, up to 200,000 cubic meters (53 million gallons) of solution could be moved into or out of the water balance system, if proper methods are designed and installed. That is equivalent to one meter (40 inches) precipitation falling on a 200,000 square meter (2.2 million square feet) pad area. You may even get a little bit more gold recovered as the solution drains from the earlier pads, as well.

Jan 1, 1999

By Richard M. Crowell, John W. Shield, Todd J. Harvey

Facing lower gold prices, gold mine operators throughout the world are seeking ways to reduce costs for treating refractory sulfide ores. Many gold mines, currently producing oxide ores, have underlying refractory sulfide deposits and the operators could extend the mine life by processing these sulfides. Fortunately, oxidation technologies for refractory ores need not be complex nor costly. GeoBiotics, Inc. has developed heap biooxidation technologies (GEOCOAT™) for treatment of refractory gold ores and concentrates. GEOCOATTM uses simple low cost unit operations common to heap leaching of gold and copper ores. The cost effectiveness and flexibility of GEOCOAT™ heap biooxidation allows mining companies to expand operations to include refract01y ore. In addition, GEOCOAT™ can be utilized to supplement or replace existing oxidation processes such as roasting, autoclaving, or agitated tank biooxidation, and to treat a lower resource grade than is currently economic. The GEOCOAT™ technology involves coating sulfide concentrates onto sized support rocks and biooxidizing the sulfides in a heap configuration. After biooxidation the concentrate is stripped from the support rock, neutralized and leached in conventional cyanidation facilities. Ashanti Goldfields Company Ltd. (AGC) produces over 885,000 ounces of gold per year from oxide and refractory ores at its Obuasi operations in Ghana. Refractory flotation concentrates are treated in roasters and in the world's largest stirred tank biooxidation facility. In August 1998, AGC and GeoBiotics jointly constructed and operated a field demonstration heap of the GEOCOATTM process at the Obuasi operations. The heap contained I 082 tonnes of support rock and 125 tonnes of concentrate. The heap was approximately 19 meters square at the base and 3.8 meters high, and ran for 88 days after which the concentrate was reclaimed. Average gold recovery after biooxidation was 91.5% with a NaCN consumption of 8.6 kg/t. Baseline gold recovery by direct cyanidation of the concentrate was 53.8%.

Jan 1, 1999

By J. E. Stephens

US Borax began exploration in southeast Alaska in the summer of 1971 with a staff of two, Barry Watson and Jackie Stephens, based in Spokane, WA. After several years of prospect examinations in the islands west of Ketchikan, AK, and with the able assistance of prospector-pilot Kenny Eichner, Jackie Stephens proposed and organized a boat-based, helicopter-supported, stream sediment sampling program to explore the western contact of the Coast Range intrusive complex on the mainland. The rationale for this included: the extension of known mineral trends from Canada into southeastern Alaska along favorable geologic zones; the western contact area of the Coast Range intrusives with metasediments and metavolcanics was a favorable target for metals exploration as shown by numerous small prospects. Knowledge gained from mineral exploration of the island areas indicated that they were well mineralized, but already well explored, and that the known mineral potential was probably present in subeconomic quantities (Brooks, et al.). Publications of the US Bureau of Mines and other scientific agencies were gathered and every prospect within the southem southeastern Alaska area was plotted. This in itself lent a tremendous amount of knowledge to the total exploration program. US Borax&apos;s McLean Arm copper property on Prince of Wales Island was drilled and dropped while this compilation of data was proceeding. We found that many companies had followed one another around the island areas, but little or no exploration had been done on the west side of the Coast Range, and that many explorationists flatly stated that the west side of the Coast Range was barren of mineralization. In summary, although some of the shoreline and near shoreline areas of the mainland had been explored, no organized boat-helicopter exploration program had been done along the favorable western zone of the Coast Range Intrusive complex. We decided to use a boat as a floating hotel for the crews that were to be helicoptered into the back country from tidewater for our initial stream-sediment sampling program.

Jan 1, 1991

By G. Johnston, P. K. Chatterjee, S. Holguin

M. Gershon The article is an excellent one, providing a basis for the conduct of the major tasks in underground mine planning and scheduling. I object to only one word. Since that word is featured in the title, it is a strong objection. The word used erroneously is optimization. There is no optimization performed in the article. By definition, an optimization is an approach that yields the best of all possible solutions. We can limit this definition to its narrowest sense, the optimization techniques of operations research such as linear programming, or broaden it to include less sophisticated search procedures that find the best mine plan or schedule. No such claim is made in the article. No optimization of any kind is performed. The major tool used is a simulation, which, in truth, is probably more appropriate for this problem than any optimization. But simulation does not optimize. It only allows for quick evaluation of many options, leading to the selection of the best plan "of those tried," not "of all possible plans." This article, of course, is not the first mining paper to make this error. It seems that the word optimization is used more to impress people than to refer to any act of optimizing. It is used in a way that implies "try to optimize"or "come close to optimize," but not to really do it. It is usually appropriate to quietly ignore this error. But when Dr. Chatterjee, who has done so much good work in true optimization projects and whose work I have come to admire, makes such an error, it is time to publicly dispute it. Having discussed this very point with him in the past, I am sure that he will agree. If he can support this position in his response, perhaps people will start to listen and we can reduce the number of false claims of optimization. This should be easy for him in this case since the approach he used, simulation, is the more appropriate one anyway.

Jan 8, 1988

By J. Katzur

Recently in Germany, novel artificial humus in combination with water-storing hydrogels has successfully been tested as soil improving material in pot experiments. The novel artificial humus (NOVIHUM®) is derived from lignite through oxidative ammonolysis. The product contains c. 5.5% nitrogen. Using NOVIHUM® as fertiliser, variable amounts of the contained nitrogen are gradually available to the plant, depending on the hydrolysis rates of the different N-compounds. NOVIHUM® is a unique combination of long-term fertiliser and stable humus. The tested hydrogel is a polyacrylate/poly-amide copolymer, which is able to improve the water holding capacity of the soil and to retard significantly desiccation. In May 2000, in Germany (Aue region), an experimental site was set up under extremely difficult growing conditions for trees (Quercus petraea) on a uranium spoil pile of the mining company WISMUT Ltd. The oak plots treated with NOVIHUM® and hydrogel showed increased total increment-values, which mainly are the result of the second growth phase. Obviously, these plants profit from the favourable physical parameters of the substrate and the use of the novel artificial humus as nitrogen source. The oak trees can better use the plant-available NH4-nitrogen released from the artificial humus during the planting year. In Greece, an experiment was set up in December 2000 aiming to the afforestation (Pinus nigra, Abies cephalonica) of a bauxite spoil pile of SILVER & BARYTE ORES MINING Co. S.A., on the Parnassus Mountain. The scope was to demonstrate, that the above soil improving materials are also appropriate for the recultivation of sites under extreme climatic and orographic conditions. The first results revealed, that where NOVIHUM® and hydrogel were combined, plant growth was greater than the average. The one-year results of the experiments fulfil our expectations concerning the application of NOVIHUM® in combination with hydrogel for the afforestation of problematic sites, such as the south-exposed steep slopes of the spoil piles of uranium mining. They may represent the most detrimental sites for afforestation in Europe.

Jan 1, 2002

By N. E. Norman

A great deal of effort has been put forth in recent years to arrive at accurate projections of the demands for tunnels in this decade. In reviewing the statistics presented by various study groups, it appears that during the decade from 1960-69, there were some 1,335 miles of rock and soft ground construction tunnels excavated throughout the world. These statistics indicate 645 miles of rock tunnel and 689 miles of soft ground tunnel were excavated during the last decade. This is expected to increase to 3,000 miles in the 1970&apos;s according to projections made by various authorities in the field of construction tunneling. I point out the above statistics to emphasize two main points: (1) The demands for tunneling in this decade alone are staggering. (2) There is a fairly even distribution in the amount of rock tunneling and soft ground tunneling being excavated. Unfortunately, available statistics do not answer many important questions about tunneling which are of great importance to the contractor and to boring machine designers. For example, how many of the projects completed in the past decade involved both soft ground and rock tunneling? How many projects involved tunneling in conditions where both rock and soft ground were contacted at the same face. While the answers to these questions remain unanswered, I am sure most of you can think of numerous examples of projects where mixed face conditions are prevalent. Many of you are probably familiar with the Austin Crosstown Interceptor which was bid last month in Austin, Texas. Here is a project involving 57,000 feet of tunnel where limestones and clay shales are expected to be traversed. Another project, the Tonner Tunnels No. 1 and No. 2, near Los Angeles, involve some 22,000 feet of tunnel where mixed soil and rock conditions are expected. A large majority of the projected tunnels for this decade are to

Jan 1, 1997