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By Jonathan Ludlow, Robert A. Jankowksi

Introduction A longwall operator can make few changes to increase output, significantly reduce respirable dust, and decrease power consumption. Reducing drum speed, and thereby cutting with increased pick penetration, is one. This article defines the benefits of deep cutting in terms of reduced dust production and power consumption. It also identifies the practical aspects of high pick penetration in terms of shearer performance and coal loading. Before examining some practical aspects of reducing drum speed and looking at the theoretical background, it is worthwhile to summarize what is meant by high penetration and deep cutting, and what potential benefits and pitfalls may be expected. Deep cutting (in the sense of high penetration rather than wide web) can be defined in one or more of the following ways: • Cutting with an average pick penetration distance higher than that used in the past. • Cutting with a pick penetration higher than the longwall operator would have used if the advantages of deep and slow cutting were not considered. • Cutting with a well-designed shearer drum below 40 rpm. All these definitions are slightly arbitrary. They are given to provide a basis for discussion and to make the point that any move towards deeper, more efficient cutting can result in operational benefits. The benefits of deep cutting appear in many different areas. The most noticeable benefit, provided suitable instruments are available, is the reduction of airborne respirable dust. During an experiment on a longwall in the Pittsburgh seam, a nearly four to one reduction in dust levels was seen when drum speed was halved. Not all studies have shown such a big reduction, but it seems that some benefit is almost always obtained when drum speed is reduced. Production rate and specific power consumption are also affected (in a positive sense) by reducing drum speed or increasing pick penetration. Although these changes may not be as spectacular as those in dust level, they contribute to the economic return of the longwall operation. Similarly, improved washability through fines reduction may have a beneficial economic effect. Cutting with shearer drums operating at lower speeds does have some possible deleterious impacts that an operator should be aware of. For example, cutting reactions - loads imposed on the picks by the coal being cut - will be increased as a deeper cut is used. Steps must, therefore, be taken to ensure the stability of the shearer and provide an adequate haulage effort. These increased cutting reactions also result in higher loads on the power transmission system (gearboxes, ranging arms, pick boxes, etc.) from the shearer motor(s) to the pick tip. These higher loads must be anticipated and provided for with the necessary hardware. In particular, extra haulage power must be provided with low drum speeds, since haulage effort required increases roughly in proportion with pick penetration. Because the drum will be rotating more slowly or will have fewer picks, the load on shearer components will also be more variable. If suitable, robust equipment is not used, this increased vibration will decrease reliability. Benefits of Deep Cutting Lower dust levels, decreased specific power consumption, and improved product washability are the most noticeable benefits of reduced drum speeds. Although the benefits will vary greatly with mining conditions and the type of coal, some examples of what can be expected are described below. Reduced Dust Levels Figure 1 shows principal results of a study on the effects of reduced drum speed conducted on a longwall in the Pittsburgh seam (Ludlow, 1981). This figure shows that average dust production was reduced by about 70% when drum speed was halved. By making some assumptions about such quantities as coal density, it is possible to apply this proportional reduction to the quantity of respirable dust liberated per ton of coal mined. When this is done, two kinds of results are obtained: • At 70 rpm, about 1 g (15 gr) of airborne respirable dust is created for every ton mined (roughly one part per million). At 35 rpm, only 0.28-0.37 g/t (3.9-5.1 gr per st) of coal mined become airborne respirable dust. • At 35 rpm, nearly four times the amount of coal may be mined before the compliance level is exceeded, compared with 70 rpm.

Jan 3, 1984

By G. T. Newell, J. G. Stone, V. M. Mejia

Introduction Advances in drilling have reached a point where large diameter cores can be recovered from "tight," or weakly indurated placer gravels. In such ground, core drilling can provide more reliable data regarding tenor than can be obtained using churn drilling or similar classical techniques. It can also provide metallurgical and geological information that is not available from samples obtained through alternate methods. In 1985, Coastal Mining Co, a subsidiary of M. A. Hanna, and Western Gold Reserves began to review a Tertiary placer deposit owned by San Juan Gold at North Columbia, CA, about 14 km (9 miles) northeast of Grass Valley. The deposit is one of the largest remaining unmined portions of the formerly extensive early Tertiary ancestral Yuba river system. It has been known since the 1850s, has been the subject of much technical literature, and has been the object of at least four previous drilling programs. The eastern one-third of the 6 km (3.7 mile) stretch of the channel between North Columbia and Badger Hill was partially stripped by large scale hydraulic mining in the late 1870s and early 1880s. Mining ceased in 1884 when the Sawyer Decision prohibited further discharge of hydraulic tailings into the Sacramento and San Joaquin Rivers. By that time, about 30 to 45 m (100 to 150 ft) of relatively low grade upper gravels had been removed over some 81 hm2 (200 acres). About 90 to 105 m (300 to 350 ft) of higher grade middle and lower gravels were left at least partially stripped. In 1914, a few churn holes were drilled along a widely-spaced line. In 1938-1939, Selection Trust conducted an extensive drilling campaign to evaluate the deposit. Particular attention was directed toward the partially stripped eastern portion. In 1968, the US Geological Survey drilled three churn holes in the eastern part of the deposit. The US Bureau of Mines conducted experimental mining and drilling in the Badger Hill area. In the late 1970s, Placer Service Corp. acquired a lease on the deposit. Between 1979 and 1984, Placer Service drilled 28 large diameter BADE (a German-manufactured machine) drill holes on the eastern portion of the deposit. The surviving records from the widely-spaced 1914 drilling program are fragmentary and the reported grade not well substantiated. The 1968 holes were drilled for scientific purposes. Again, drilling details are not available. However, detailed records for both the churn drilling program and the BADE program were available and formed the basis for the initial evaluation of the property. Geology The geology of the auriferous Tertiary gravels of California have been described by Whitney (1880), Lingren (1911), and, more recently, Yeend (1974). In general, the Tertiary gravels in the North Columbia area occupy a broad channel cut into pre-Tertiary igneous and metamorphic rocks. The upper, or white gravel is overlain conformably by volcanic tuffs and volcaniclastic rocks. A middle gravel is characterized by the presence of silicified and carbonized wood. A lower blue gravel unit has relatively coarser cobbles and contains a higher proportion of igneous and metamorphic cobbles than the other units. The upper gravel consists of interbedded pebbly sand and silty, or clayey sands with prominent cross bedding. Most of the pebbles are well rounded and consist mostly of white vein quartz and quartzite. The upper unit is moderately well compacted. Exposures in the walls of the old hydraulic mine pits stand at 45° and 50° angles. The gold content of the unit is well below an economic cutoff. The middle gravel - included with the upper unit by Yeend (1974) - is coarser grained, with carbonized wood, and 75 to 100 mm (3 to 4 in.) cobbles of metased-imentary and metavolcanic rocks in a sandy matrix containing abundant lithic fragments. The upper contact appears to be conformable, but the lower portion of the unit appears in places to consist of reworked lower gravels. The unit contains less clay than the upper unit and is somewhat more friable than the underlying lower gravels. The gold content, while somewhat higher than the upper level, is too low to be of ore grade. The lower gravel averages between 30 to 45 m (100 to 150 ft)

Jan 9, 1988

By Richard R. Klimpel

INTRODUCTION The process of froth flotation as a means of upgrading the quality of coal by removing water and/or ash and/or pyrite has been receiving increasing attention since the 1960's by the world-wide coal industry. Historically, coal preparation practice has involved the use of primarily gravity based separation techniques, screening, and/or water washing in removing the larger fragments of inert material from coarse raw feed coal. The preparation, handling, and cleaning of raw coal finer than 500 micrometers was deliberately avoided. Most often the fine coal that was produced would be judiciously incorporated into the bulk of the larger sized cleaned coal product or simply discarded. Since the 1960's and accelerating in the 19701s, the desirability of performing some type of coal cleaning or upgrading on the finer raw coal feed materials has been increasing. There are several apparent industrial incentives for implementing fine coal ((500 urn) processing alternatives including the gradually increasing value of the normally lost fine coal product; the additional amounts of fine coal being produced due to changes in mining techniques; the need to appropriately dewater and upgrade fine coal streams in order to meet increasingly stringent environmental regulations on water quality. storage of fines, etc.; the increasing need to lower undesired sulfur and ash contents in some coal end-use applications; and the increasing development of speciality coal use processes that require inherently finer and cleaner coal particles such as in gasification, liquefaction, fuel injection, etc. Looking at the over 70 years of mineral oriented fine particle separation successes at the industrial level, the use of froth flotation in fine coal would also seem to be a natural process to utilize at the industrial level. In the world of strictly mineral processing technology and, more importantly in the world of mineral commodity economics, there are clearly a number of advantages to using froth flotation that are well documented and universally accepted. These include: the relatively low capital and space requirements required when compared to other process options available to upgrade minerals; the extreme flexibility inherent to the froth flotation process both with regard to throughput and differences in raw feed characteristics; the general availability of a wide range of equipment sizes and reliable suppliers of equipment; and the general availability of appropriate and relatively cheap chemicals that are the major driving forces behind mineral flotation. Just how has the implementation of froth flotation fared in coal upgrading over the last 20 years? It is towards answering this question that this paper will be oriented. As such, the discussion will be somewhat philosophical. The comments offered are based on the personal experiences of this author who has worked in the basic side of inventing new flotation chemistry and flotation analysis techniques with the major overriding goal of implementing this new technology successfully at the industrial level in both coal and mineral flotation throughout much of the world. This paper is not intended to be a complete reference to all published scientific work on the subject of industrial froth flotation of coal. As a brief background to what will be expanded upon in this paper, the author would like to offer the following summary comments: 1) the industrial practice of fine coal froth flotation is qualitatively similar but quantitatively different than industrial mineral froth flotation when viewed from either research considerations, operating viewpoints or economic evaluation criteria; 2) it is clearly possible to invent and/or develop new technology (such as new chemistry and/or equipment) that can do statistically better than current plant practice in well-controlled tests in either coal or mineral flotation circuits; 3) there is so much inherent variability to the raw feed coal streams and other operating parameters that even if there is a documented technical incentive to implement some new piece of technology, the people, instrumentation and control methodology required are often simply not available or not economic unless the potential payoff is quite large.

Jan 1, 1988

By Wayne P. Myers

Electricity, as normally thought of by a layman's definition, is a man- made force that has no color, no odor, is not visible, cannot be heard, yet man can control it and make it perform his work for him. A person dare not cheat on it, for it then will master him. It has laws, rules, and regulations that must be strictly adhered to if it is to perform in the safest, most reliable, and economical manner. In the early days of coal mining the power source was man himself. Later man employed steam and compressed air to reduce his muscle power requirements needed in the extraction of coal. By the turn of the 20th century, electricity had become the most common source of power for the coal industry. Steam and compressed air are still found in use in isolated cases today for specific applications; however, they are rapidly being replaced by electricity. Electrical laws and regulations are quite extensive and somewhat complex for the proper generation and use of electricity. Man-made laws, rules, and regulations for the care and handling of electricity, designed for man's safety, are rigidly enforced by Mine Safety and Health Ad- ministration (MSHA), as well as the various state departments of mines.

Jan 1, 1981

By Mark E. Smith, Gerald J. Gierszewski

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

Jan 1, 1987

By G. T. Lineberry, W. J. Wiehagen

A paradigm shift As market economists, mine managers are interested in "staying in business" and even "prospering." Bottom-line results are the "order of the day." Concern for bottom-line results can invoke questions such as how much was produced, at what cost, and was anyone injured? While these questions will always carry everyday emphasis at the mine, they are fundamentally lacking in that they often appear, especially to lower levels of management and to the production worker, as pre-eminent concerns. Driven by market economics, many organizations have decentralized, outsourced, downsized, restructured and formed new coalitions. Investments in new technology and a concentration on productivity have been an essential part of these changes, contributing to improvements in both safety and efficiency. A reason for (and, perhaps a result of) these improvements has been the development of a highly experienced and flexible work force. This situation presents an interesting opportunity to modify a few paradigms about work life in general, and perhaps a few paradigms about how we invest in the miner, in particular. Studying these issues now night offer some interesting opportunities for the Future, as the foundations for the next generation of mine workers are laid. With the average "experienced" U.S. miner five to fifteen years from retirement, perhaps a key determinant of future success in the world market might be how well we make the transition from today's work force to the work force of tomorrow. We can ill afford to wait a few more years to tap into the special knowledge residing within a veteran mining work Force. The Work Crew Performance Model (WCPM) is one approach to defining, capturing and transferring this expertise. The WCPM suggests a paradigm shift, moving from a focus on products (quantity) to a focus on the process (quality). It also involves a slight change in the process -- a different way to think and work. Of great importance, it entails a belief and a commitment to people within the organization; a commitment to collect simple (but interesting) data; exploration for insightful ways to plot and integrate that data; and finally the search for creative methods to reinvest that knowledge back into the organization, that is, back into the work force. As a way of thinking, the WCPM subordinates information about how much is produced daily or how many were injured over a year's time to concern and detail for how the individual's and the work crew's performance contribute toward meeting organizational goals. Drawing upon the expertise within members of that work crew to address and solve operational problems is one approach for meeting organizational goals. This everyday focus on how we work is hypothesized by the WCPM to make the difference between a good section supervisor and the average one; an exemplar continuous-miner operator and the run-of-the-mill one; a mining crew that produces consistently high quantities, but, inadvertently, degrades the production of other units. The WCPM permits objective and reliable performance data to be collected. This data then can be used to define variability within tasks of work crew members; to relate observed variability to a cost consequence; and to more meaningfully analyze performance through the integration of traditional analytical tools, such as the production simulator, CONSIM, with a recommended behavioral approach, the WCPM. Basis for engineering test Most equipment selections and purchases involve a field trial test of the system. Technology is routinely bench tested at the manufacturer's facility, government agencies and the mine site. The WCPM suggests a similar, but more practical, method for testing the performance of a work system -a test for how technology is used (by people) at the work site. It implies learning from veteran mining personnel that have gone well "beyond the book." This approach to learning and to integrating training with everyday work life can help answer questions such as: What constitutes desirable perfor¬mance as, for examples, a shuttle-car operator, a miner operator or a bulldozer operator; what characterizes exem¬plary performance within the context of the work system, the technology and the work crew; how can investments in the worker he linked to measurable results within the organiza¬tion? This paper explores these questions and makes recom¬mendations for enhancing the cost-linkages between invest¬ments in people and the mine's monthly cost sheets.

Jan 1, 1996

By John T. MacWaters, George H. Milly

INTRODUCTION This paper describes operational procedures and instrumentation which were developed to implement a patented method of detecting and locating uranium ore bodies. All equipment was designed for accurate and reliable use under difficult field conditions. Data resulting from these developments, some of which are incorporated in a companion paper for this conference*, have contributed to our understanding, of the nature and significance of naturally-occurring radon exposures. Further applications of these techniques may prove useful in studying anthropogenic radon in evaluating the significance of variability in radon flux, or in taking proper account of the contribution of natural radon to ambient radon concentrations, particularly in mineralized regions. OPERATIONAL PROCEDURES The uranium search technique employed by GEOMET begins with a [reconnaissance mode], which utilizes a series of ambient air filter samples taken from a moving or standing vehicle to detect anamalous concentrations of radon. These samples are normally taken at night under stable (temperature inversion) or neutral atmospheric conditions, to avoid the rapid dispersion of the radon clouds typical of daytime turbulent conditions. Radon daughters formed by the decay of radon quickly attach to ambient aerosol particles and are captured on the surface of a membrane filter during a five-minute sampling period. The filter is placed in a portable alpha scintillation counter, and a five-minute count is started at a preset time after the completion of the sampling period. This count is followed immediately by a second five-minute count to enable determination of the cloud age, based on consideration of the relative half-lives of the initial radon daughters. With the use of tables computed from the mathematics of the uranium decay process, and with the measured radon background accounted for, the first and second counts are used to estimate the time which has elapsed since the pure radon initially emanated from the earth's surface. This cloud age, combined with wind speed and direction measurements made at the beginning and end of each sample, as well as with an understanding of the effects of local terrain features on wind trajectory, provides an estimate of the Probable location of the source. In practice the reconnaissance mode involves one vehicle with a single operator to take the five-minute samples, moving or standing, along any available roads or trails in the area to be covered. Under the usual nighttime stable conditions the cool air flows downslope and downvalley, and the operator will pay special attention to locations along the axes and confluences of the natural downflow patterns. These facilitate sampling of large air drainage areas. Successful detection of sources of radon based on high concentrations of radon daughters, and substantiated by evidence of "fresh" clouds, are frequently made on the basis of a relatively few samples taken judiciously by an experienced operator during a reconnaissance. At this point he has a choice: he can continue to cover the designated reconnaissance area, leaving the localization of any anomalous radon detections to await the ranking of detections by priority, or he can continue with localization of the detected source, either alone or by radioing for another vehicle to assist in triangulation operations. Localization of a source involves skillful and experienced use of the continuously collected data on concentration, age, and wind speed and direction, together with knowledge of the terrain from observation and contour maps, and exploitation of the available roads. Since age ratios are dependent solely on proximity of the source and wind speed, decreasing [cloud ages] based on statistically significant increasing age ratios represent the most useful. information in localization. The [magnitude of the concentration] of radon daughter products is also useful; however, it may vary widely for a variety of reasons during sampling for a localization, thus making this a secondary parameter for localization. Large values of concentration are of course useful; typical results for successful detection of a strong anomaly would yield alpha counts ranging from 10 to 25 times background values. These counts are convertible to ambient radon concentrations by appropriate consideration of the decay process. Successful localizations of radon sources using this procedure with ambient atmospheric measurements normally will narrow a source location to within an area ranging in size from less than a quarter of a square mile to at most a square mile, and frequently give information on the probable shape of the source (in plan view) on the earth's surface. Judicious use of this procedure could undoubtedly provide finer localization, but the coarser results just described are adequate for this Phase of GEOMET's uranium exploration procedures.

Jan 1, 1981

By Fred Leighton

Miners have long known that rock noise, or the popping and cracking of the rock commonly heard during mining, can indicate instability of the mine structure. For many years, miners have "listened" to the rock talk and many times have interpreted changes in rock noise activity to be a warning of failure and have retreated from the potential failure area. Microseismics, or the study of rock noises, was begun in the early 1940s partly because of this historical fact. Microseismics is based on using geophysical equipment to detect and analyze rock noises on audible and subaudible levels. Thus, these systems are much more sensitive than the human ear and "hear" even more of the rock "talk" than do miners. Research has shown that micro-seismics can be used to precisely locate portions of a working area that are generating rock noise, and that rock noise release rate information from each area can be used to analyze its stability. On the basis of research results, microseismic systems are now commercially available and have been purchased and installed in many mines around the world. Introduction When a rock mass is subjected to changing stress conditions, such as those caused by mining, small-scale adjustments occur within the rock that release seismic energy. When this energy is in the audible range it is called rock noise. Those areas where stress changes occur are also the areas of the structure most likely to fail. Individual rock noises can be detected and analyzed to determine their precise location relative to the mine structure. Over a period of time plots of this data provide a pictorial representation of where stress changes are occurring. This is because rock noise activity tends to concentrate in those structure areas most actively adjusting to the changing stresses. Since these areas are the most likely to fail, they can be pinpointed and mapped relative to the structure. Rock noise rates, or the number of rock noises occurring per unit of time, also tend to vary dramatically before failure. Thus, the ability to locate the source of individual rock noises provides a means of determining where failure may occur, and rate counting within each area offers a means of assessing when failure may occur. This information, properly treated, can be used to avoid, control, or warn of impending failure. Successful applications of this technique have provided recent impetus to the effort of developing microseismics into a practical, reliable, and economically feasible tool. The phenomenon of naturally occurring rock noise was discovered in 1938 by Obert and Duvall (1939) who were measuring seismic velocities in mine pillars. Seismic energy other than what they were generating continually registered on their recording equipment. Further study by these researchers showed that the extraneous seismic energy being recorded was from rock noises that were being generated within the rock in highly stressed areas. Pursuing this interesting phenomena both in the laboratory and in the field, Obert and Duvall documented the dramatic change in rock noise rate before failure (Obert, 1941). They also established the fact that, in many instances, rock burst failures could be predicted by monitoring and listening to the rock noise activity in rock-burst-prone areas (Obert and Duvall, 1945). These early efforts clearly showed that microseismics had great potential as a tool for measuring or estimating mine structure stability. Extended testing, however, showed that sometimes rock bursts occurred with no apparent microseismic warning, and that at other times sharp increases in microseismic data were not accompanied by failure. Also, because precise location of individual rock noises was not possible at that time, one never knew where the failure was going to occur, only that failure near the observation point was likely. Thus, while the technique clearly offered promise, it was not considered practical then. In the mid-1960s, the Bureau of Mines undertook a new research effort to improve the microseismic technique. Major improvements were judged possible mainly because new and vastly improved electronic and system components had become available as an offshoot of space program instrument development. Thus, in 1967, development of a multichannel, broad band, microseismic system began (Blake and Leighton, 1970). Application of this system in rock-burst-prone mines showed it worked well and provided the incentive to develop methods whereby the source location of individual rock noises could be easily and directly calculated (Leighton and Blake, 1970; Leighton and Duvall, 1972; Redfern and Munson, 1982). The improved system and the ability to locate rock noises quickly showed that the microseismic technique offered new and increased poten-

Jan 8, 1983

By Scott G. Britton

Today's coal mines are turning more and more to belt haulage as the key system to transport coal out of the mine. The advantages of using belt for both intermediate and main haulage are numerous. Belts are reliable high capacity haulage mediums that can be utilized at low operating costs. They are adaptable to rolls in the coal seam and can handle relatively steep grades. They are not noisy, operate with a minimum amount of labor, and are easily extended and retracted. Belts require the least power of any haulage medium, are easily controlled by push- buttons and interlocks, and present low fire and safety hazards. Belt haulage can also be the largest source of operational problems for mine management when not installed correctly. This chapter and Chap. 6 concentrate on proper installation of both the static and dynamic components of efficient belt haulage. This chapter looks at the static system component, composed of the belt drive, its installation and operation, as well as belt takeup and transfer point construction. These components do not change readily over the life of the installation, hence the term static system. Chap. 6 discusses the dynamic system of belt line construction, which includes belt splicing, breaking down and reanchoring the tailpiece, and extending the belt line, plus operational and maintenance items for belt operation. This chapter is divided into three major topics. The first is a general discussion covering the points necessary for designing an efficient haulage system; the second looks at installing belt drives, head rollers, and takeups, and the third discussion centers on installing efficient and effective transfer points along the haulage system. Each will address the design and construction issues facing the engineer or construction supervisor in charge of installing the system. Before looking at the static system, it may be useful to review some of the basic design factors and considerations of all belt haulage systems. The material usually carried on an underground belt is a mixture of raw coal and rock, commonly known as run-of-mine (ROM) coal. This mixture has certain material characteristics that guide the mining engineer to the final and proper belt selection. The characteristics looked at most frequently are: 1) Angle of Repose is the angle at which the material will freely make a pile. Most ROM coal averages a 35"-39" angle of repose. 2) Angle of Surcharge is the angle a pile will assume on a moving conveyor. This ranges from 5" to 15" less than the angle of repose in most materials. For ROM coals, 25" is the average angle of surcharge. 3) Flowability, which determines the maxi- mum cross-sectional area needed to carry any given material on a belt. It is measured by the angles of repose and surcharge of a material and also serves as an index of the safe incline angle for the belt line. These characteristics, plus practical engineering research and development work are defined by an organization called the Conveyor Equipment Manufacturers Association (CEMA) for most bulk materials. CEMA also provides a code for each classified material depending on the characteristics above and the subjective considerations of the average size to be encountered, the abrasiveness of the material, and any miscellaneous characteristics (oil resistive, corrosive, etc.) which may be important. For coal, CEMA (composed of member companies in-

Jan 1, 1983

By Wilbur I. Duvall

INTRODUCTION This chapter outlines some general principles and procedures that a rock mechanics engineer can use to arrive at a logical design for the size, shape, and spacings of underground openings in various types of com¬petent rock formations. Rock formations can be clas¬sified into six major categories which are useful for design purposes, provided the necessary input data are available regarding the geology of the site, the physical properties of the rock, the in-situ stress field in the rock, and the general geometry of the underground openings. The US Bureau of Mines (USBM) developed a rock classification system based upon these concepts and found it to be useful for design purposes (Obert, et al., 1960; Obert and Duvall, 1967). This classification sys¬tem is given in Table 1; however, an extra column has been added to the original classification to indicate the appropriate design techniques for each rock class. In this classification system, two main classes of rock are considered: competent rock and incompetent rock. Competent Rock Competent rock is any rock which, because of its mechanical and geological characteristics, is capable of sustaining underground openings without the aid of any structural support except that provided by unmined rock in the form of pillars and side walls, etc. Roof bolts and small random props which support only local loose rock are not considered structural supports. Competent rock is further subdivided into three ad¬ditional categories-massive, laminated, and jointed¬ which are defined as follows: Massive Rock: Massive rock is a rock formation where the spacing between mechanical defects such as joints, partings, or faults is equal to or greater than the critical dimensions of openings, or where the strength of the bond across these mechanical defects is com¬parable to the rock strength. Laminated Rock: Laminated rock is a rock forma¬tion divided by approximately parallel planes of weak¬ness into lamina whose thicknesses are small compared to critical dimensions of openings and where the bond between the lamina is weak compared to the rock strength. Elastic Rock-This is rock whose physical proper¬ties will permit the use of elastic theory to predict strains and displacements from applied stresses. Inelastic Rock-This is rock whose physical proper¬ties are such that strains and displacements are a strong function of time as well as applied stress. Incompetent Rock Incompetent rock is any rock which, because of its mechanical and geologic characteristics, is not capable of sustaining underground openings without the aid of additional structural support in the form of linings, steel sets, or systems of props. This classification system for rock depends not only on the mechanical and geologic characteristics of the rock but on the size of openings and their depth be¬low the surface. For example, many competent rocks at shallow depths for small stress fields could become inelastic rock at greater depths and larger stress fields, or competent massive rock for small openings could be¬come laminated or jointed rock for larger openings. Thus, to classify rock for design purposes requires a knowledge of the geologic properties of the rock mass, the physical properties of the rock, the general size of the openings to be mined, and the average range of the in-situ stress field to be encountered. To design underground openings and structures in rock, it is necessary to use criteria of failure for rock under various stress conditions, and to use safety factors to allow for differences between laboratory and in-situ strength properties of rock and to account for errors in¬troduced by the various assumptions made in the design process. Criteria of failure based on maximum stresses have been found to give reliable results in practice. Thus, it is assumed that rock will fail in tension when the tensile stress exceeds the tensile strength of the rock as determined by a standard modulus of rupture test on samples of rock core. If the tensile stresses in rock are small, then it is assumed that the rock will fail at a value of compressive stress equal to the compressive strength of the rock as determined by standard uniaxial compres¬sive strength tests on samples of rock core. These fail¬ure criteria, when used with appropriate safety factors, can be expressed as:

Jan 1, 1982

By Jozef S. Stachulak

Inco Limited operates 11 underground mines in the Sudbury District. The mines are located on the rim of the Sudbury Basin, an oval with the axis in the range of 27 and 60 km. The ore dips to at least 3000 m below surface. The ores are mined primarily for nickel and copper. Total ore production from underground is in excess of 55,000 tons per day. Over 40 surface fans have been installed since the late 1960's. All of the fans are adjustable pitch, axial flow units. A major factor influencing ventilation design in the last 30 years has been the introduction of diesel equipment underground. Volumes per fan have ranged from 60 to 330 (cubic metres per second), with motors from 100 to 2500 hp. Fans of the axial flow type have been in common use for main fan installations at Canadian mines for many years. The standard arrangement has been to mount these fans horizontally, i.e. with the fan shaft and the long axis of the housing horizontal. This is a natural arrangement for an underground fan, but for a surface installation, a vertically mounted fan has definite advantages. The surface area taken up by a typical vertical fan installation is generally about one quarter of that with a horizontal fan of the same capacity. (1) This is not a problem with isolated fans and flat surface outcrop sites, but where the installation is to be near existing buildings, or where there are poor surface soil conditions, space and cost considerations greatly favour vertical fans. MAIN FANS INSTALLED There are 43 main surface fans in operation at 11 mines. Twenty-seven of these fans are supply units, and 75% of them are vertical installation. The remaining 16 units are main surface exhaust fans, with predominantly horizontal installation. Within the last five years, some 20 main booster fans have been installed underground at several mines. Axial flow fans, with adjustable pitch blades, are used for both surface and underground installations. Exhaust fans are equipped with stainless steel or cast aluminum blades. Main underground fans are arranged horizontally and the majority of them have a floating shaft between the fan shaft and the motor. (2) The size of the fans in service varies from 1.8 m to 5 m in diameter, with the majority ranging from 1.8 to 2.5 m. The pressure produced by these fans varies from 0.25 kPa to 2.0 kPa. At Inco Limited, two main fans in parallel are preferred, rather than a single fan, so that if one fan fails, the remaining fan can still supply up to 70% of normal air quantities, while the damaged fan is repaired. This requires closure doors on each of the fans so the fan can be isolated in case of failure. It is more expensive than a single fan, but results in less production interruptions. The fan installations are well away from sharp inlet and outlet bends. FAN DESIGN INTEGRITY Both the mine operator and the fan manufacturer must understand that the main fan is critical to the mine operation, and that everything technically possible must be done in design and manufacture to ensure the highest degree of reliability. Some of the design parameters and criteria, based on Inco experience, are outlined and discussed below. RESONANT FREQUENCIES AND HOUSING MODEL ANALYSIS Any fan assembly will have many different resonant frequencies. It is a challenge to the designer to arrive at a design in which forcing frequencies do not coincide with any of these resonant frequencies to produce unacceptable vibration levels in operation. Finite element analysis is a useful tool that can be used to identify the most critical of these, so that housing and blade stiffness can be adjusted to change any resonances that might be close to forcing frequencies. Shaft critical speeds should be at least 25% above the fan operating speed, and there should be sufficient separation between other resonant and forcing frequencies to avoid excitation that might result in high vibration levels. QUALITY ASSURANCE Radiographic Blade Examination Mine fan blades are normally cast in small lots. To ensure that the castings are sound, a full radiographic examination is recommended in the highly stressed lower 1/3 of the blade.

Jan 1, 1995

By Y. Xi, T. M. Yegulalp

An availability approach for determining the capacity of a major Bucket Wheel Excavator (BWE) system to be used in the Yuabaoshan Surface Coal Mine in China is presented in this paper. Based on previous studies on other BWE systems for delay times due to internal and external causes and a four-level maintenance and repair schedule, a model was developed, taking into account all random failures, to present a fast and simple solution to this highly complex problem. A procedure employing the Markov renewal theory for determination of the distribution of random failure times and failure rate of the whole system, as well as some other factors useful for calculation of the overall system capacity, are discussed.

Jan 1, 1994

By Michael W. Lewis

The latest downturn in the world’s metal prices has forced many mining companies to focus on cost reduction and productivity improvements to remain competitive. Maintenance represents a significant part of the overall operating costs in the mining industry. A structured, proactive maintenance approach, such as reliability-centered maintenance, can achieve cost reductions and productivity improvements. Remote condition-monitoring systems are becoming commonplace because they have the potential to identify problems before failure. Detection of equipment degradation before failure will enable repairs to be scheduled, thereby reducing costs and interruptions to production. The mining industry is in the midst of uncertain times. Decreasing commodity prices, demand, head grade and capital availability along with increasing environmental requirements have put pressure on mining companies to significantly reduce costs to maintain short- and long-term viability. Corporations are looking at expansion, merger/consolidation and financial/hedging programs to achieve these cost reductions. Mines are reducing costs by eliminating all but essential expenditures, increasing production levels, reducing overtime and investing in new technologies.

Jan 1, 2000

Crushing and grinding The Crushing and Grinding Technical Committee There was substantial activity in the area of crushing and grinding with the improved outlook in the industry. The Crushing and Grinding Committee's theme of 1987's Annual Meeting in Denver was Fundamentals of Mechanical Design for Crushing and Grinding Mills. At the meeting, the concept of gearless sag mill variable speed drives was taken out of the closet, dusted off, and a wraparound motor was put on hard copy with the ordering of a 9.7- x 4.5-m, 8.2-MW (32- x 15-ft, 11,000-) sag mill. Additionally, comparisons were provided to the mill design of the gearless drive sag mill and the twin 4.5 MW (6000 hp) do variable speed drives placed on the drawing board for the Kennecott UCD Expansions. This involved a 10- x 4.5m, 8.9 MW (34- x 15 ft, 12,000 hp) sag mill. Differing opinions were also expressed in technical presentations pertaining to the need for quality mill specifications in design and manufacturing. J. Berney-Ficklin of Bechtel presented a paper outlining the need for quality design and manufacturing specifications in large scale grinding mills, and the overall cost reduction in capital and operating costs. Others presented differing specific views, but all emphasized a central theme. Quality and reliable designs in large mills should not be compromised. In-pit crushing applications continue to increase, overseas and in the US, in both mining and quarrying. Different in-pit crushing units have been classified as mobile, with integral transport systems; semi-mobile, typically modular systems that use separate transporters for repositioning; semi-stationary plants that must be dismantled before transporting and typically require earthwork and concrete foundation work; and stationary in-pit crushing plants, which remain in place for the life of the pit. The majority of equipment sales of conventional crushing and grinding circuits in 1987 were to gold mining operations. There have been two areas of end use: • conventional jaw/gyratory or impact crushing circuits preparing low grade ores for heap leaching; and • small scale sag/ball mill circuits preparing higher grade ores for classical cyanide leach/ recovery circuits. There was also continued interest in ABC and SABC circuits for both base and precious metals plants. High pressure roll crushers continue to attract considerable interest in the cement industry. There, they are used to prepare raw materials or clinker for subsequent ball/tube milling. However, they have apparently been less successful in the mineral industry where the energy savings/ capacity gains have been much less dramatic. Several papers covering the theory and application of these machines were presented in an industrial minerals session on cement at the SME Annual Meeting. The Third International Conference on Hydrocylones, held in England in October produced several papers of interest on classification and screening to the mining industry. In particular, a new cyclone apex, developed by Mt. Newman Mining, shows promise of reducing water flow to the cyclone underflow stream and, therefore, increased underflow pulp densities and reduced fines recycle in grinding circuits. A twin vortex hydrocyclone was also described and shown to give a sharper split than conventional units. Similar double cyclones were introduced in the 1950's, but failed to achieve widespread acceptance. This was due mainly to a proneness to blockage in the transitional zone. It will be interesting to see whether this latest variant can overcome this problem in the typical operation. A number of North American iron ore plants installed the capability to produce fluxed iron ore pellets. Most chose to grind the stone on site with circuits designed to grind taconite ore. Methods employed to grind this material included single-stage ball mills closed with fine screens, rod and ball mill circuits - both open and closed with cyclones - and semiautogenous mills, followed by ball mills closed with screens or cyclones. Dewatering and tailings disposal B.M. Moudgil and D.L. Sober, University of Florida Due to tightened environmental regulations and the need for more efficient land use, a greater interest in dewatering and waste disposal has developed. As a result, research efforts in flocculation, surface chemistry, and polymers science have focused on the problems related to the dewatering of solid mineral wastes. A few studies have been conducted to examine the flocculation process. Hogg et al., (1987) discussed the formation and growth of flocs. They determined the size and density of flocs was controlled by the physical conditions of the system (e.g. agitation,

Jan 5, 1988

By C. Widham

Introduction It is generally accepted as dogma that sampling contributes most of the error in gold fire assays. Differences in assay results on pulps from the same sample interval are frequently regarded as evidence of the presence of the so called "nugget effect" of relatively coarse gold particles. It is true that coarse gold particles can contribute to substantial sampling fluctuations. But, while the process of sampling is probably the major source of error, the analytical process cannot be completely ignored as a possible contributor to erratic assay results. To maintain a stable assay process, the analytic part of the system must also be kept in control. One method of monitoring the performance of the analytic system is to systematically assay standard materials, whose sampling characteristics are carefully controlled. Gold assay standards are not prepared, nor can they be prepared, to account for both sampling and analytical errors. It is not possible to send coarse material to a lab for both preparation (i.e., comminution and splitting) and fire assaying and then come to conclusions only about the fire-assay process. Because most gold ores are very heterogeneous, sampling errors would, in most cases, completely mask the contribution of the analytical errors. Assay-standard material is prepared only to assess the accuracy and variability in the fire assay process. Because the objective of the assay standard is to provide information about the fire assaying, it is necessary to control the sampling error of the standard material, so that it is only a minor constituent of the discrepancies observed in any assay results. To do this requires that the particle size of the standard material be reduced to a point where the relative standard deviation of the sampling error (i.e... the standard deviation of the errors divided by the average gold content of the material) is 2% or less. For all but very homogeneous mineralization, this means that the material must be reduced to 100% -150 mesh before the sampling errors are adequately controlled. However, even reducing the particle size can contribute to sampling problems. The liberation of gold may cause segregation that can cause large sampling fluctuations that are not easily controlled while maintaining the desired grade. Because, in most cases, the standard material would already be in the "pulp" state when it is submitted to a lab for assay, it is not possible to entirely conceal the nature of the sample from the lab. This is a problem inherent in using assay standard material. Because of the contribution of sampling to error generation in the assay process, the use of "coarse" material does not solve the problem of submitting a totally "blind" standard to the lab. In the sections that follow, the selection, preparation, testing and use of gold fire assay standard material is discussed. While some may dismiss the production of standard material as folly, it is possible to produce and utilize standard material to stabilize and improve the fire-assay process to produce more reliable assay results. Material selection It is desirable to use material that has as nearly the same metallurgical characteristics as the samples with which the standards will be included. However, this is usually difficult. For many reasons, including the particle size at which a significant amount of the gold mineral is liberated, the sampling characteristics of even -150-mesh material may preclude the use of geologically and metallurgically similar ore as a standard. It is usually easier to get material having desirable grade characteristics with the necessary sampling properties than it is to find geologically and metallurgically similar material with the required sampling characteristics. High-grade standards are especially difficult to find and prepare. This is because, as grade increases, the size of the gold particles usually increases. Larger gold particles are liberated and tend to segregate during comminution, and the homogeneity of the material cannot be maintained. For grades much above 3 g/t (0.088 oz/ton), it is very difficult to find material that has the proper sampling properties. Old mill tailings are likely candidates for assay standards. Some of these have sufficiently homogeneous mineral contents, so that the sampling errors can be effectively controlled. Where mill tailings are either not available or are not acceptable, mineralization that has exhibited homogeneous results in reassays of the pulp material is also a good candidate for the standard. Finally, the mineralized rock being sampled may (and should) be used if adequate homogeneity in the -150mesh material exists. "Adequate" ("acceptable") homogeneity is defined below.) It is important to use standards having a wide range of grades. This alone may preclude the material being

Jan 1, 1997

By J. C. Yingling, R. J. Sweigard, S. T. Mickleborough

The geostatistical technique of conditional simulation is applied to measure the uncertainty inherent in the factor-of-safety (FS) predictions for soil slopes in which there is an incomplete knowledge of the spatial distribution of soil properties. A case study that considers a natural clay slope in Canada is presented. The sequential-indicator simulation technique was used as a means of generating a series of spatial maps of the soil properties at the site. These maps approximate random draws from the conditional probability density function that characterizes existing uncertainty in the spatial distribution. FS determinations were made for each simulated map using the simplified-Janbu approach. In using this approach, computations are performed using a new computer program that can operate on irregular spatial profiles of soil properties specified on a discrete grid. A unique capability of this approach is the ability to generate the FS distribution conditional to the type, position and value of existing site-characterization measurements. This capability should reduce uncertainty in FS predictions.

Jan 1, 1996

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

By K. S. Venkataraman

The physical properties of clay-water systems depend on the complicated system of forces between the clay particles themselves, and between the clay particles and the ions in the liquid phase. The kind and distribution of ions in, on, and between the clay particles and the size and the shape of the particles are the basic factors determining the macroscopic behavior of clay-water systems. Understanding the system requires a knowledge of the nature of the clay particles, their size, structure, composition, and surface properties, and of the manner in which they interact with ions [and molecules] in the surrounding liquid [or other medium]. The validity of Professor Brindley's words (Brindley, 1958), written three decades ago in the context of making pottery, whitewares, and electrical porcelains, transcends time, and the basic message is perhaps all the more important in the considerably expanded use of ceramics for structural, thermal, tribological, electronic, and other applications. Silicon carbide, silicon nitride, and sialons have been studied in the last two decades for high- temperature structural and tribological applications, particularly for using in internal combustion engines. Titanates, zirconates and niobates of barium, strontium and lead, have high dielectric constants, and are extensively used in the formulations for making capacitors. Hexagonal ferrites (molecular formula MO.6Fe2O3) are in use for making permanent magnets for fabricating miniature motors, and for assembling loud speakers, particle accelerators etc. Cubic ferrites such as magnesium-zinc ferrite and nickel-zinc ferrite are used as transformer cores, and for other high-frequency applications. In this context, Richerson's recent book (Richerson, 1984) on the general scope of traditional and technical ceramics is a good starting point for an overview of contemporary ceramics technology. Glasses are a whole class of amorphous materials used widely as sintering aids, and for making glass-bonded ceramics and glass-ceramic composites. Composites are yet another burgeoning field where two or more particulate components are used for improving the performance of ceramics. For all these applications, the inorganic starting materials are almost always submicron and near-micron powders. Understanding the powders' physicochemical properties, and their surface chemical interactions with the surrounding liquid/gaseous medium is-necessary for making reliable ceramic parts at competitive prices. Even though ceramics science and engineering has attained its separate identity in universities and the industry, ceramists themselves would concede that ceramics science is a cross-disciplinary field, having incorporated and assimilated within itself many principles from several apparently disjointed disciplines. Principles of material science, graduate-level physics and chemistry, polymer science, surface and colloid chemistry, transport phenomena, particle technology, unit operations commonly used in chemical engineering and mineral processing, and statistics and applied mathematics are integral part of any ceramics curriculum in universities. Added to this is the fact that all bench-scale successes in making ceramic parts are to be scaled-up for larger throughput operations. Understanding and applying process engineering principles of comminution, classification, drying, calcination, etc. then becomes essential. CERAMIC FORMING: Despite the diversity of the materials and processes, conceptually, the steps involved in making ceramic parts have remained the same over several decades: The different components for making the pan (usually one or more powders plus other forming and sintering additives) are proportioned and mixed thoroughly, and the well-mixed formulations are consolidated into desirable shapes known as "green bodies." Usually binders such as wax, clay, organic polymers and surfactants, whether dispersed or dissolved in a suitable liquid are used during mixing the batch for giving strength for the green bodies. In the dried green state, the inorganic powders typically occupy only 55 to 60% of the bulk volume of the body, depending on the particle size distributions of the powders and the forming history, with mostly inter- particle voids accounting for the rest of the void volume. SINTERING: The formed bodies are then fired in high- temperatures kilns/furnaces during which the parts are exposed to a predetermined temperature profile, and "soaked" for a certain duration at the final high temperatures, typically between 1200 K and 1900 K, and then cooled to room temperature. The gaseous atmosphere in the furnace is controlled (oxidizing, reducing, or inert) when necessary. During the initial stages of firing, volatile liquids evaporate, and during the intermediate temperatures between 400 and 600 K, the the organic polymeric additives pyrolize and oxidize into water vapor, CO, C02, and other gases. At still high temperature, the glasses, when present, soften, and simultaneously, the ceramic particles rearrange into a network of grains with definite grain boundaries so as to reduce the total interfacial free

Jan 1, 1990

By Alim Elibol, Rosbel Jimenez, Mustafa Can Suner, Deniz Tuncay, Zach Agioutantis

LaModel is a displacement-discontinuity variation of the boundary element method, which is utilized to analyze the displacement and stress in flat-lying, tabular orebodies. In coal mining, the accuracy and reliability of LaModel are proven with the utilization of the Mark- Bieniawski coal pillar strength and the stress gradient formula with proposed methodologies. However, the application of LaModel is limited to coal mines due to the default coal pillar strength properties. Recently, the stress gradient functions have been established and used to derive concentric rings of zones to simulate stone mine pillar yielding in LaModel. Also, to gain a robust understanding of the brittle failure mechanism and its effect on pillar strength, an S-Shaped failure envelope, where rock mass' cohesive and frictional strength components mobilize as a function of confinement (e.g., minimum principal stress), has been implemented into FLAC3D software. Therefore, by employing brittle failure criteria, the pillar strength stages at various width-to-height ratios and different pillar dimensions are estimated for four benching stages. The estimated pillar strengths and stress gradients for benching conditions are then used to calibrate LaModel in-seam material properties and compare stresses computed by LaModel and FLAC3D. Consequently, this study aims to increase the safety measures in the stone mining industry by implementing the practical use of LaModel for underground stone mines' initial development to various stages of bench mining.

Jun 25, 2024