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By I. Jashnani, J. R. Lawter, D. B. Lindsay

INTRODUCTION It is a generally recognized obligation of owners and operators of underground mines in the United States to protect the health of miners against the potential carcinogenic effect of airborne alphaemitting radionuclides, in particular the shortlived radioactive daughter-products of 222Rn. In meeting this obligation, a number of strategies are available, and it is largely up to the operator to decide what combination of methods he will use. Because the officially accepted theory of the mechanism by which radon and its short-lived progeny cause cancer of the respiratory system, the objective in controlling this health hazard is to limit as far as possible the inhalation of the three daughter products, 218Po, 214Pb, and 214Bi. (214Po, the fourth nuclide in the chain, is so short-lived that its alpha-decay occurs almost instantaneously after the beta-decay of its parent, 214Bi.) The reasoning in this theory leads to the conclusion that the alphadecay of radon itself causes essentially negligible physiological damage, and that the amount of daughter product activity attributable to decay of radon within the respiratory system is also very small. As every ventilation engineer in this business knows, however, focusing remedial actions exclusively on the daughter products themselves is choosing the hard way, and will probably never lead to effective control of this health hazard in underground mining activities. The reason for this troublesome fact can be seen in Figure 1. Let us say that we have air containing 222Rn and its short-lived daughters in near-equilibrium, a not-uncommon condition in the air of a poorly ventilated underground mine, and that we wish to purify it by removing only the toxic daughter products. If we have been very thorough in our purification efforts, we might expect to achieve a 99% reduction of the Working Level (WL) measurement. Figure 2 allows us to examine closely the early regrowth of WL's in this freshly cleaned air. Within the first five minutes, for instance, the WL's are back to 10% of their full equilibrium value, and ten minutes later the level has more than doubled and is on its way to an almost complete defeat of the clean-up effort by the end of the next hour. It is quite clear from this simple illustration that no matter how effective an air-cleaning system may be if it is directed only at the selective removal of the toxic trio, 218Po, 214Pb and 214Bi, its beneficial effect will be quite temporary. A truly effective attack on the problem must always include methods aimed at the control of the unruly parent, radon. Then, even if its progeny have an inalienable "right to life", their brief lives will not be spent in the miners' lungs. The object of the present paper is to show how state-of-the-art environmental air

Jan 1, 1981

By Jeanne A. Zimmer, Andrew B. Cecala, Edward D. Thimons

The Bureau of Mines performed a study to determine optimal locations for longwall face methane monitoring equipment. Longwall face ignitions continue to be a concern to the U. S. coal mining industry. Gains in ignition control technology have been somewhat negated by advances in longwall productivity. More reliable, heavy-duty longwall face systems generate higher production rates and increase methane liberation rates. These rates will likely continue to rise in the future as further coal production increases occur. Efforts will continue in the areas of methane drainage, methane containment, and methane dilution. In addition, more effective face methane monitoring is needed. As technology continues toward automated longwall panels, the importance of monitoring for methane at the shearer becomes even more significant. Since the primary face ignition source is the shearer mining machine, the Bureau ran tests to determine the best location to take methane readings on the shearer. This information, in conjunction with data obtained from field studies, provides a knowledge base for selecting optimum face methane monitoring locations. INTRODUCTION Face ignitions are a concern for many U. S. longwalls and will continue to be until better control technology can be developed or until reliable and accurate knowledge can be provided as to when an explosive concentration is present. Figure 1 shows the number of reported U. S. underground longwall ignitions from 1980 to 1990. The number of longwall panels have remained relatively constant over this time period with 89 reported in 1980, compared to 96 in 1990. In spite of technological improvements in mining over the past decade, no significant reduction in the number of longwall face ignitions has occurred. One factor contributing to this is that the U. S. coal industry is mining gassier seams causing methane liberation rates on the average to be higher (Trevits, 1991; Grau, 1987). Another factor is the development and implementation of more reliable, heavy-duty longwall face systems that are generating higher production rates and higher methane liberation rates. Additionally, the average panel width in 1991 was 714 ft; a 44 pct increase from 1980 when faces averaged a width of 495 ft (Merritt, 1992). Increases in panel width also impact total face methane emissions. [ ] There have been substantial advances in control technology in the areas of methane drainage, methane containment, and methane dilution. In some mines, efforts to drain methane in advance of mining have proven effective, but geological conditions often preclude this approach (McCall, 1984; Ely, 1989). In other cases, ventilation air quantities of over 100,000 ft3/min and face velocities of over 1,000 ft/min have been implemented on longwall panels. Other operations have implemented anti-ignition spray systems which use water sprays behind each cutting bit to quench the hot streaks created from cutting noncoal materials (Courtney, 1990). Other control techniques, such as the modified shearer clearer system and ventilated drum, have been used to increase the amount of turbulence and ventilation around the cutting drums, thus reducing methane levels (Cecala, 1986; Divers, 1987).

Jan 1, 1993

By Milton H. Ward

The percentage of exploration dollars flowing out of the industrialized countries continues at an all-time high, and this shift in risk money is justified (Samuels, 1990; Bosson, 1977). Clearly, the odds of discovery are greatly enhanced if exploration is pursued in an untrampled and geologically more prospective environment. Unfortunately, such major unexplored areas of the developed world are shrinking rapidly, not just because the lands have already been tested, but primarily because they have been withdrawn from exploration. Experience, however, has shown that foreign investment is not without risks. These include problems that result from working in isolated areas, a lack of infrastructure, untrained workers and an absence of a solid geological data base. While the likelihood of once-feared expropriation is now low, mining enterprises still must take precautions to gain the support of their constituencies, including a variety of governments (local, provincial and central), workers, environmentalists, cultural organizations and local landowners. Is foreign exploration worth the risk? In most cases, Yes! But maintaining tenure and long-term success in a less developed country is only likely with good planning and the use of appropriate safeguards. These include measures such as buying country-risk insurance, building multinational relationships, forming joint investments, developing a broad customer base and seeking local investors in the company (Cobbe, 1979). The primary theme of this lecture is: How a company handles its constituency is critical to long-term tenure in a foreign country. Ideally, a multinational mining company would be so effective in dealing with the local people and managing its relationship with the host government that it would become a preferred employer, preferred over its competitors and over local companies. Historical overview Our earliest ancestors lived in Africa three to four million years ago. The Paleolithic or Old Stone Age, about 500,000 years ago, marked the beginning of mining in its earliest form (Gregory, 1980; Wolf, 1984). Various types of rocks and stones were used for tools and weapons, with flint being the preferred resource for implements. Initially, these materials were collected as stones (eoliths) lying on the surface. Then they came from excavations of flint in limestone beds that were exposed in river banks or naturally occurring excavations. Later, more formal mining developed. No doubt tribes of the Old Stone Age experienced conflict over rights to eoliths, food and land. The question of who owns the land and has rights to its contents and what is fair value for extraction is still the subject of intense debate and, occasionally, open conflict. As time passed, the search for useful minerals required migration from one area to another. Later, voyages from Europe to the New World and travels to other remote regions were funded and supported by absentee investors, kings and such (Rogers, 1976). The primary objective of such funding was to obtain gold, silver, food stuffs and other commodities that could be collected from the new countries. This is still the objective of foreign investment.

Jan 1, 1994

By Joseph M. Hans

INTRODUCTION The decontamination of a uranium mill site is principally an earth moving task. The task can be completed effectively and efficiently by selecting the proper earth moving equipment and using that equipment within its particular constraints for decontamination work. A considerable backlog of experience concerning earth moving equipment selection and use was acquired during the decontamination work conducted at the former Shiprock, New Mexico, uranium mill site. This experience was acquired by trial and error while decontaminating over a wide range of depths and different types of terrain. Many different types of excavation equipment were available to the Shiprock decontamination project because the Navajo Engineering and Construction Authority (NECA) heavy equipment operator's school was doing the decontamination work and a variety of equipment was needed for training. The equipment is described and its general use for decontamination is discussed. Its on-the-job use for decontaminating the Shiprock millsite is presented. [EARTH MOVING/ANCILLARY EQUIPMENT DESCRIPTION AND DECONTAMINATION CONSTRAINTS] Tractor-Scrapers The most commonly used scraper is the bottom-dump type because it is highly mobile and can dig, carry, and spread loads. In this type a tractor is connected with a yoke to a scraper bowl (Figure 1). The bowl is essentially a rigid box that pivots around the rear axle, opening to the front. A hardened steel knife, or bit, is attached to the leading edge of the bowl. When scraping, the bit penetrates and forces the earth overlying the bit edge into the scraper bowl as the scraper moves forward. When the bowl is full, it is closed by lowering a closure or apron against the cutting bit. The load is dumped while moving, by raising the apron and forcing the load from the bowl with a crowd, or ejector. The height of the dumped load is controlled by adjusting the bit height when dumping. Scrapers of this type generally do not have enough power to scrape up loads unassisted. Loading assistance is generally provided by a crawler tractor that pushes from the rear against the scraper's push block. Some scrapers are made self-loading by adding an additional engine for the extra power needed for the load pick up. Another commonly used self-loading scraper uses a loading elevator (Figure 2). The elevator consists of two roller chains carrying crossbars. The elevator is rotated when scraping and the crossbars cut the earth above or in front of the bowl bit, forcing the load into the scraper bowl. No apron is used on elevating scrapers. The load is dumped by partially retracting the bowl floor and pushing with an ejector. Scrapers have some undesirable characteristics when used for decontamination work. One characteristic is that a clearance space exists around the apron when it is in its closed position. Part of the load (especially sandy soils) dribbles through the

Jan 1, 1981

By Dale W. Norris, John A. L. Campbell

INTRODUCTION Optimization, as measured by productivity gains, is being achieved in our nation's preparation plants principally by automation, economy of scale, and reducing maintenance. Increased tonnage per unit of investment, a1 so a measure of optimization, is being achieved by breaking with tradition and increasing plant utilization, while simultaneously improving on equipment re1iabi1ity. Improved qua1ity, another measure of optimization, is being attained through technical innovation. This review will concentrate on state-of-the- art unit operations and processes, and will not discuss emerging new technologies which are covered in detail in other presentations of this symposium. It is recognized that much of what will be discussed in this paper is not unique to the fine coal circuit, but is equally applicable to the entire preparation plant. However, the areas reviewed are the principal means being used today to optimize an existing preparation plant's fine coal circuit. Prior to designing and optimizing a fine coal cleaning circuit, the question that must first be answered is, "Do we even want a fine coal circuit? This is a much more complex question than is readily apparent, as most of you recognize. It can be argued either way. In the final analysis the "bottom line" will hold (Blackmore, 1982). If your sales contract a1 lows it, and the selling price can justify it, you most assuredly want to build a fine coal circuit into your preparation plant. In general today, most new preparation plants do have fine coal cleaning circuits. The purpose of fine coal cleaning has changed with time. Some of the early applications of fine coal cleaning were to prevent stream pollution rather than to prepare a product for market (Merrill a Pennington, 1962). During the 50's, 60's and into the 70's the principal thrust of fine coal cleaning was in metallurgical coal preparation. More recently, the emphasis of the 80's is on steam coal processing. There is no consensus on what is fine coal. Traditionally, fine coal would be anything below 1/4" in size. Because of the advances being made in intermediate coal cleaning processes, i.e. heavy media cyclones and jig cleaning, this discussion will define fine coal as -28 mesh (Aplan, 1987). Figure 1 is a plot of the size ranges of a large number of U.S. bituminous coals. In general the -28 mesh fraction of a bituminous coal will be between 7% and 18% of the ROM feed with reported ranges of 5% to 40% (Aplan, 1987). Consequently, traditional fine coal cleaning is general 1 y involved with a re1 actively mi nor portion of the coal feed. Every effort is taken both in terms of mining and preparation to keep it that way because of the high cost of cleaning fines. The move away from this posture, namely, the techniques to fine grind coal to liberate pyrite is probably the biggest challenge facing fine coal processing today.

Jan 1, 1988

By Thomas J. Peters

At the request of the U.S. Bureau of Land Management, the U.S. Bureau of Mines and U.S. Geological Survey conducted a mineral investigation of about 85,470 acres of the Diablo Mountain Wilderness Study Area, to determine mineral resources and mineral resource potential. The study area is in Lake County Oregon, about 45 miles northwest of Lakeview. Fieldwork was conducted in 1986, 1988, and 1989. Geologically, the study area lies on the northwest edge of the Basin and Range physiographic province, a region of fault block-formed mountains and basins characterized by interior drainage. Consolidated rocks are mostly Tertiary basalt and tuffaceous sedimentary rocks. Low-lying areas are covered by Quaternary alluvial-fan, sand-dune, playa, lacustrine, and landslide deposits. The principal structural features are normal faults that have large vertical offsets; these faults are typically concentrated at the margins of large horst and graben. The 34,310-acre "additional," or western part of the area was mainly studied in 1988 and borders the east shore of saline Summer Lake. In this portion of the study area there is potential for undiscovered resources of soda ash, boron compounds, and sodium sulfate as well as by-products, which include magnesium compounds, salts, potash, bromine, lithium, and tungsten associated with brines. No energy or mineral resources were identified in the study area, but brines within lake and playa sediments contain concentrations of chemical components suitable for the production of soda ash, boron compounds, and sodium sulfate. Possible by-products include potash, salts, bromine, lithium, magnesium compounds, and tungsten. Limestone from the study area could be used in the recovery process of brine components or agricultural uses. The study area includes thermal springs and lies within 2 miles of the Summer Lake Known Geothermal Resource Area. Local production of geothermal power by the binary process may be economically feasible. Soda ash, soda ash products, and boric acid are widely used in fluxing metals, and have important applications and markets in the aluminum industry in the northwest and the developing Chinese counterpart. Evaporite commodities are essential to many "backbone" industries, and to many new applications and advanced materials. Markets for brine mineral products appear to be undergoing steady and strong growth, especially in the Pacific Northwest and in Pacific Rim countries. Soda ash produced near the study area would be 55 percent closer by rail to marine export at Portland, Oregon, than trona deposits of the Green River District (Wyoming). Soda ash and a daughter product, caustic soda, and sodium borohydride will receive increased application in the bleaching of paper pulp for environmental reasons. Caustic soda, used in many industrial processes, is more environmentally friendly when derived from soda ash than when produced from sodium chloride salt, because of the absence of a chlorine by-product. Producing soda ash from brines, a type of in situ" mining, can be done with minimal environmental degradation, and mining natural soda ash is environmentally as well as economically preferable to synthetic soda ash production.

Jan 1, 1995

By K. Vanbroekhoven

An industrial site near the centre of a historical town in Flanders is characterized by chromium(VI) contamination of the groundwater. Since the potential for natural attenuation was not sufficient to control the groundwater plume as observed from different analyses like pH, Eh, soil oxidation and reduction capacity, the feasibility of site remediation based on in-situ bioprecipitation of chromium was examined in lab microcosm studies. By creating reducing conditions by addition of a carbon source in the groundwater, chromium(VI) can be biologically reduced to chromium(III), which can form an insoluble precipitate of chromium(III) hydroxide. Microcosms contained contaminated groundwater and aquifer samples that were collected and manipulated under anaerobic conditions. Three different carbon sources were screened for their effectiveness: molasses from sugar beet production, molasses form starch production (protamylasses) based on potato juice and lactate. The different concentrations of sugar molasses were used to determine the optimal conditions. A range of parameters was monitored at regular time intervals for a period of six months such as chromium(VI) and chromium(III), DOC, methane, pH, Eh. A rapid conversion of chromium(VI) to chromium(III) was observed for all conditions tested. However, extensive removal of chromium from the groundwater by precipitation was only achieved for a limited number of conditions. In general, lactate was found to be the most effective carbon source, slightly better than the lowest concentration of sugar molasses that was tested, which itself was more effective than molasses based on potato. Higher concentrations of molasses, although causing a steeper decline in redox potential, resulted in less effective Cr removal. Typically, the fermentation of molasses results in acidification and this drop in pH may hinder the formation of insoluble chromium(III)hydroxides. It may also be possible that certain components of the molasses form soluble complexes with chromium(III). For lactate, no significant drop in pH was observed. Following the laboratory microcosm tests, the stability of the chromium precipitates was examined by performing a sequential extraction on the microcosm materials. These were performed under redox conditions which were similar to the ones observed in the field. Since the outcome of these tests was positive, a pilot scale test was implemented in the field in April 2005. Two different substrates, the cheap waste product sugar molasses and lactate are being injected creating two different impact zones and typical parameters are monitored. Frequent re-injections were important to maintain reduced conditions and are carried out three-weekly up to now. To study the stability of precipitates formed, mesocosm socks containing aquifer material were installed in the monitoring wells and harvested at regular intervals. Results of this pilot test will be presented and compared to the laboratory results. Based on the success of this field test, a full scale process will be designed and implemented as remediation strategy.

Jan 1, 2006

By Everett M. (Original by) White

Numerous articles have been written about the coal mine ; he has been compared with brave men who have gone out to conquer some unknown hazard throughout the ages. Now, however, modern mining is no longer something to be avoided, but a vocation that anyone can be proud of, with remuneration equal to or above that of almost any industry. A concerted effort is being made each day by both the operators of most mines as well as the state and federal inspectors to see that coal mines are as safe or safer than work places in any other industry. How- ever, it is true that mine fires and explosions have occurred in the mining industry since its inception. One of the earliest accurate references to such an occurrence is found in the transactions of the Royal Society of London. In the Colliery Managers' Handbook, 1898, Caleb Pamely writes of an explosion that occurred at Fatfield, England in August 1708, resulting in the deaths of 69 persons. He states: "Three of them, viz, two men and a woman, were blown quite up from the bottom of the shaft, fifty-seven fathoms deep, into the air, to a considerable distance from the mouth of the pit." It is interesting to note that women were employed as miners in England during that period. In the United States, the first record of coal mined was in 1702 by the Jamestown colonists. In 1786, about 20 years after Pittsburgh, PA was first settled, a charter was granted by William Penn to mine coal on the banks of the Monongahela River. Then in 1803, the first important coal company, the Lehigh Coal Co., was established. From this beginning, the mining industry continued to grow in importance and the first state mining law was formulated in 1870. The Pennsylvania state mining law was followed in 1873 by Ohio's, and in 1876 by Maryland's. A mine explosion at Avondale, PA in 1869 stimulated state legislatures to formulate these early laws. In 1907 near the town of Fairmont, WV, there were, within three days, two explosions resulting in a total of 597 deaths. Because of this, the federal government took action and by 1910, the US Bureau of Mines had been established. This was perhaps the most significant change in health and safety for miners during the first half of the 20th century as the USBM concentrated its entire efforts in the study, research, and testing of methods totally directed towards improvement of the health and safety of the miner. After a mine disaster in Illinois, a federal law was passed in 1940 calling for federal coal mine inspectors. Then in 1969, Congress enacted the Federal Coal Mine Health and Safety Act following a mine explosion near Farmington, WV. In 1977, the law was revised and included noncoal mining operations.

Jan 1, 1981

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

1.1 INTRODUCTION When underground mining involves total extraction, it induces overburden strata movements. If not properly planned it causes surface subsidence and affects surface environmental conditions. Total extraction usually refers to longwall mining and room and pillar mining with pillar extraction. Surface subsidence has long been a subject of intensive research for scientists all over the world and considerable achievements have been obtained. However due to its difficulties and complicated nature, research into overburden movement has been thus far incomplete as compared to that into surface subsidence. Since surface subsidence is a manifestation of the results of overburden movement, the processes and mechanisms of overburden movement must be fully understood in order to establish the mathematical prediction models of surface subsidence. In this chapter the processes of overburden movement and its zones of movement in the overburden will be discussed. 1.2 PROCESSES OF OVERBURDEN MOVEMENT When total extraction is used, it produces a large void in the coal seam and disturbs the equilibrium conditions of the surrounding rock strata. The roof strata bend downward while the floor heaves. When the excavated area (or gob) expands to a sufficient size, the roof strata will cave. As a result, the overlying strata continue to bend and break until the piles of the fallen rock fragments are sufficiently high to support the overhanging strata. At this tie the overhanging strata no longer cave, but bend and rest on the underlying strata. Strata bending and subsidence develop upward until reaching the surface and forming a subsidence basin. The whole overburden strata and the surface subsidence basin will further go through a period of compaction and gradually become stabilized. Current knowledge regarding the process of overburden movement has been derived from several sources. One is the direct observation of the mined sections and their surfaces, another from field monitoring of strata movements in the overburden, and others from computer modeling and scale modeling in the laboratory. For example, Figs. 1.1 and 1.2 show the results of a borehole monitoring experiment (Borehole B-2) using the full profile borehole inclinometer and full profile borehole extensometer for lateral displacements and vertical subsidence, respectively (Conroy and Gyarmaty, 1983). The panel, 400 ft wide by 5000 ft long and 630 ft deep, was extracted from the Pittsburgh No. 8 seam which had an average thickness of 54 in. The face advanced from east to west. Curve A in Fig. 1.1 shows the inclination of the borehole when it was 45 ft ahead of the face at Position A (i.e., on the solid coal side). The total deviation of the borehole was 1/7000. The first shear movement occurred at 130 ft above the coal seam and appeared to have occurred along the bedding plane with large contrast of rock strata on both sides, i.e., sandstone vs. shale. When the face had passed the borehole 79 ft at Position B, the borehole deformed to assume a bow shape with a maximum deviation of 2.8 in. from the center line. Numerous shearing planes occurred and extended further upward. When the face had passed the borehole by 75 ft at position C, strata subsidence was measured. Fig. 1.2 shows that a strata separation as large as 4 ft occurred at 40 to 100 ft above the coal seam. Above this level, the strata subsided more uniformly with bedding separations within 1 to 2 in. Another example is the longwall mining with complete caving in the Soviet Union's Karaganda Coalfield (Kolebaeva, 1968). The coal seam was 6 ft thick and 154 ft deep. The immediate roof was the fine-grained sandstone interbedded with coarse-grained sandstone. Above this, there was sandy shale and shale. In order to monitor the process of overburden movement, 15 stations were established in a borehole (Fig. 1.3). The elevations of those stations were measured when the face was at various distances from the borehole. Fig. 1.3 shows the movement history of each station. For convenience of comparison, the movement history of each station was plotted on a common reference point as shown in Fig. 1.4. Clearly, the movements of strata from Station # 1 to #6 were nearly simultaneous, i.e., they behaved as a single unit. The strata from Station #7 to # 15 had differential separations, the maximum of which was 40 in. As the face passed by and moved away, strata separations in general reduced gradually. But strata separation between Stations #12 and #13 remained the same until the end of the monitoring period. The total bed separation between Station #7 and # 15 reached 11 in., approximately 18% of seam thickness, when the face was 132 ft beyond. Based on the above observations, the author developed a conceptual model of overburden movement due to underground longwall mining (Fig. 1.5). The above-mentioned two case studies illustrated that subsidence in the overburden strata propagates upward and subsidence velocity decreases from the bottom to the top. When the subsidence velocity at the surface reaches the maximum value, the subsidence velocity at the bottom portion of the overburden strata has decreased considerably and the strata have begun to compact. Bed separation occurs within a certain distance above the coal seam and reduces from the bottom to the top. When the face has moved away, bed separation reduces gradually, so that eventually some beds completely close and others partially close. Bed separation reaches the highest level when the subsidence velocity at the surface is at its maximum value. For instance in Fig. 1.4 surface subsidence velocity was maximum when the face was 38 to 44 ft past the point. At this time the total bed separation was maximum. Overall the strata directly above the opening are subjected to tension in the vertical direction. But above a certain level, all strata move nearly simultaneously. There is also significant shear movement along some bedding planes. 1.3 ZONES OF MOVEMENT INTHEOVERBURDEN After the extraction of a longwall panel or room and pillar section of sufficient width the strata in the overburden are subjected to various degrees of movement from the bottom to the top. According to the movement characteristics, the damaged overburden can be divided into four zones (Fig. 1.6). Caved Zone. After the extraction of coal, the immediate roof caves irregularly and fills up the void. The strata in this zone not only lose their continuity completely, they also lose their stratified

Jan 1, 1992

By B. S. Crocker

Introduction Pebble mill grinding can be defined as wet or dry grinding in a tumbling mill with pebbles used as grinding media instead of metal balls. To clarify the terminology, which may differ in the various mineral industries, it is essential to point out the differences between grinding with partially sized and carefully sized grinding media. When grinding with partially sized media (limited top size only), sizing is usually the setting of the primary crusher and there is no bottom sizing-all fines, chips, sand, and small sizes are included. This is often called run-of-mine or primary autogenous grinding. In this case the entire crusher discharge is fed into the mill and part of it becomes "charge" and the rest is "sand feed." When grinding with carefully sized media (i.e., media that is sized both at the top and bottom of the size range), no fines or chips should be present in the charge to the mill. This is pebble milling and the grinding media consists of ore, rock, Danish pebbles, gravel pebbles, or ceramic pebbles-media that is appreciably lighter than steel balls. The grind¬ing charge is fed to the mill to maintain a load and it does the grinding usually on a classified sand product of suitable size. If the grinding media is the same as the raw material being ground, this type of grinding is also called secondary autogenous grinding. Pebble milling was in common use in the gold mines of South Africa in the early 1900s. It is by far the oldest form of autogenous grinding. It may be used to grind crushed ores from a maximum size of about -% in. to all -10 µm. Pebble milling differs from "run of mine" grinding in many ways, but two big differences are: I) The ore itself is crushed to approximately t/4 in., or it may be further reduced to 8 or 10 mesh in a rod mill or fine crusher. 2) Then carefully screened ore of the correct size is placed in one or more stages of pebble mills to grind the ore to its desired final size. The media size is carefully selected to be the most efficient for the particular size reduction required in each stage. This is a precise practice, and pebble mills can be designed as easily and accu¬rately as can rod mills or ball mills by those skilled in the art. A pebble mill is a rotating cylinder in which the grinding media has been carefully sized for the job in hand. If the mill is doing coarse grinding, the media will be correspondingly large and the mill might be called a primary (or intermediate) pebble mill or a lump mill. If it is to do fine grinding, the media selected is proportionately smaller and the mill would be called a secondary or tertiary pebble mill. The media itself is normally rounded pebbles made from screened ore from the primary-crusher discharge. The charge could be rounded pebbles of the correct size removed from a primary autogenous mill, or the pebbles could be carefully sized gravel, sized beach pebbles, or sized flint pebbles. Usually the specific gravity of a pebble is from 35 to 55% of that of steel balls. It commonly varies from 2.66 to 4.2, depending on the specific gravity of the ore. Primary pebble mills, or lump mills, would be charged with pieces of rock up to 10 in. maximum and 5 in. minimum and would grind crusher discharge with a maximum size of 1/4 in. on hard ores and 3/8 in. on softer ores. Secondary pebble mills could be charged with pieces of rock screened on 4 x 4-in. square mesh and retained on 1 1/2 x 1 1/2 in. These mills would grind material with a top size of 8 to 10 mesh. Tertiary pebble mills could be charged with screened rock that would pass 3 x 3 in. and be retained on 1 x 1 in. and grind 35 to 48 mesh sands. Principles of Operation Media Size To operate a pebble mill efficiently it is imperative that great care be taken in screening the rock feed that is fed into the mill to round up into a range of pebble sizes that are suitable for grinding of the sands that are also being fed into the mill, having regard to the top size and the average size of such sands. Since most mill operators know from experience the correct size of steel balls to be used in a mill to handle different sizes of sand feed, it is convenient to relate the pebble size to a ball size. If the correct ball size is not known, it is relatively simple to conduct batch laboratory tests in 12- to 18-in. laboratory mills to determine the optimum size of steel ball. Also, Bond and others have published formulas for calculating the optimum ball size. When the correct size of the individual steel ball is known, then a pebble of the same weight will perform the same grinding if used in the proper size of mill and under the proper conditions, which will be defined later. This media size, as defined by the average weight of the individual pieces, is most important. The following tests will serve to clarify this point. Laboratory Tests Using Steel Balls or Pebbles. Several labora¬tory grinds were made to compare a given size and weight of steel ball with a pebble of the same weight but larger size, and of the same size but lighter weight. The following loads were used in the 12-in. laboratory mill: All 1 1/4-in. steel balls weighing 122 ± 10 g. All 1 3/4-in. pebbles weighing 123 ± 5 g. All 1 1/4-in. pebbles weighing 45 ± 3 g. To equalize capacity, the test with the steel balls was run only 34% as long as the tests with pebbles. In each case, 2,500 g of primary sands were ground at the optimum gravity for the media being tested at a speed 78 1/2 % of critical speed, and for 170, 340, and 680 revolu¬tions for the steel balls, and 500, 1,000, and 2,000 revolutions for the pebbles. The first test is summarized in Table 74 by showing the 80% passing size in microns for the feed and for each of the grinds. Two other tests are summarized in Table 75 showing the screen analyses

Jan 1, 1985

By M. R. Shahan, W. R. Reed, P. Zhang, J. P. Rider

Roof bolter operators may be exposed to high respirable dust concentrations on continuous miner sections with blowing face ventilation when bolting is performed downwind of the continuous miner. One solution to reduce the high respirable dust concentrations is to use a canopy air curtain (CAC) to deliver clean air from a filtered blower fan directly to the bolter operators under the canopies. The influence of CAC installation in the airflow and dust dispersion around the location of the roof bolter operator can be evaluated by using computational fluid dynamics (CFD). This study, performed by the National Institute for Occupational Safety and Health (NIOSH), considers two scenarios: (1) a roof bolting machine in the center of the entry for installation of the fifth row of bolts from the face, and (2) a roof bolting machine positioned close to the face for the installation of the last row of bolts. In both scenarios, the bolting machine is placed in an environment which contains 6.0 mg/m3 of respirable dust and is ventilated by a blowing curtain with 3,000 cfm (1.42 m3/s) of air. This environment is used to simulate the roof bolter machine operating downstream of a continuous mining machine. Two operation positions are simulated at the same bolting location: dual drill heads in the inward position for two inside bolts; and dual drill heads in the outward position for two outside bolts. The influence of the CAC on airflows and dust dispersion is evaluated with the CAC operating at 250 cfm (0.12 m3/s). INTRODUCTION Roof bolting machines are used to drill bolt holes into the roof of an underground mine at specified patterns and to install roof bolts to secure the roof from falling. At times, the roof bolting machine is also used to install rib bolts; however, only the roof bolting operation was considered in this simulation. Though the roof bolting operation drastically improved the safety of miners from roof falls, some roof bolter operators are experiencing health problems associated with respirable dust. Based on recent statistics [1], following a low point in the late 1990s where the prevalence of coal workers’ pneumoconiosis (CWP) in miners with 25 years or more tenure was approximately 5%, CWP now exceeds 10% for miners with 25 years or more of tenure. In central Appalachia (Kentucky, Virginia, West Virginia), 20.6% of long-tenured miners have CWP and 4.5% have progressive massive fibrosis, the most severe form of CWP. CWP, also known as black lung, is an occupational respiratory disease that has no cure and can ultimately be fatal. The only way to prevent the incidence of black lung is to eliminate miners’ exposure to respirable coal mine dust. Current U.S. regulations require that the overall respirable dust concentration to be 1.5 mg/m3 or less for a full-shift sampling if respirable quartz is not present; otherwise, the standard is 10 divided by the percent quartz (not to exceed the 1.5 mg/m3 respirable dust standard) [2, 3].

Jan 1, 2019

By Warren W. Frech

11.1-FRAGMENTATION PRINCIPLES WARREN W. KRECH Energy must be supplied to rock by direct or indirect means to fragment that rock. The amount is dependent upon the properties of the rock and the type of loading system. Fragmentation energy is consumed by three main mechanisms: (1) creation of new surface area (fracture energy), (2) friction (plasticity) and (3) elastic wave energy dispersion. The loading method determines the relative proportions and the amount of energy consumed in fragmenting a given rock type. Unconfined tensile failure consumes the least energy, with an increasing amount of energy required as the rock is more highly confined within a compressive stress field during fragmentation. The way energy is applied by tools to cause rock or mineral fragmentation is important in determining fragmentation efficiency. Rock-tool interaction should always be controlled to the fullest extent possible to optimize the combination of rate of energy input and fragmentation energy requirements. While the direct cost of energy, (dollars per horsepower-hour) may at first glance appear low relative to other hourly costs, there is a tie between power consumption and other cost factors, such as maintenance and wear. A high rate of energy application utilizing a rock-tool interaction producing only unconfined tensile failure would provide the most efficient fragmentation. No tool that produces a combination of high energy input rate and a loading system that produces failure under only conditions of unconfined tensile load now exists. Present mining methods and tools produce failure under conditions of high compressive confinement, high dispersive energy loss and high frictional energy requirements. Thus, while generally applying energy at high rates, they are highly inefficient with respect to energy consumed in fragmenting rock. To best design fragmentation tools and optimize fragmentation systems it would be desirable to know how rock properties influence breakage. At present, rock properties directly- related to fragmentation are undefined. Although present testing technology- does not permit rigorous definition of in-situ rock fragmentation behavior, several properties (mechanical and physical) commonly are used to indicate the relative case with which rock breaks. The more common test is for uniaxial compressive or tensile strength of cylindrical cores taken from intact portions of the in-situ rock (testing details are provided ill Sec. 6.2, Rock Mechanics). The strength of rock is influenced by the environmental conditions imposed on the rock. Those of most importance in rock are (1) confining pressure. (2) pore fluid pressure, (3) temperature and (4) rate of load application. Increase in confining pressure, as with increasing depth beneath the earth's surface or under the action of a fragmentation tool, causes an increase in rock strength. Increases from atmospheric pressure to 10.000 psi commonly cause up to n fourfold increase in the compressive strength of crystalline rock. Apparent rock strength decreases as pore fluid pressure increases, since it decreases the effect of confining pressure. The effective stress concept of soil mechanics generally applies to rock. Although chemical effects of pore fluids influence rock strength, they geologically are small compared to the confining pressure effect, except for a small minority of rock types. Increase in rock temperature causes a decrease in rock strength. This effect

Jan 1, 1973

By P. Bissert, L. Yan, D. Yantek

"Federal regulations require refuge alternatives (RAs) in underground coal mines to sustain life for 96 hours while maintaining an apparent temperature (AT) below 35°C (95°F). NIOSH research has shown that heat and humidity buildup is a major concern with RAs because they have a limited ability to dissipate heat, and high internal air temperature and relative humidity (RH) may expose occupants to heat stress. The heat transfer process within and surrounding an RA is very complex and is not easily defined analytically nor experimentally. To investigate heat and humidity buildup in RAs, NIOSH conducted multiple in-mine 96-hour tests on a 6-person metal-type RA, a 23-person tent-type RA, and a 10-person tent-type RA. The test results show that when moisture was introduced into the tests to represent perspiration and respiration from miners (wet tests), the average measured internal air temperature at midheight increased by 7.7°C, 9.4°C, and 10.5°C, and the RH approached 95%, 94%, and 88%, respectively. For the dry tests when no moisture was introduced, the average internal temperature increased by 8.4°C, 10.3°C, and 12.6°C, respectively. INTRODUCTION If an accident occurs in an underground coal mine, miners who cannot escape from the mine can enter an RA for protection from adverse conditions, such as high carbon monoxide levels. One of the main concerns with the use of mobile RAs is the thermal environment inside the RA. The metabolic heat of the occupants and the heat released by the CO2 scrubbing system will cause the interior air temperature to increase. Moreover, the humidity within the RA will increase through occupants’ respiration and perspiration and from the chemical reaction within the CO2 scrubbing system. The internal thermal conditions can subject miners to heat stress, which can lead to heat exhaustion, heat stroke, or even death depending on duration and magnitude of exposure. MSHA regulations require that RAs must be designed to ensure that the internal apparent temperature does not exceed 35°C (95°F) when the RA is fully occupied [1]. Apparent temperature (AT) is a temperature-humidity metric for the perceived temperature caused by the combined effects of air temperature, relative humidity (RH), and wind speed. It is used to assess the perception of indoor temperatures when workplaces are not sufficiently heated, cooled, or insulated to provide comfortable or healthy conditions. Before a mobile RA can be deployed at an underground mine for emergency usage, it must be tested to show that its internal AT when fully occupied will not exceed the 35°C (95°F) limit. For practical reasons, RA manufacturers usually conduct their tests at above-ground test facilities. To investigate RA thermal response in an in-mine environment, NIOSH conducted multiple 96-hour tests on a 6-person metal-type RA, a 23-person tent-type RA, and a 10-person tent-type RA in its underground coal mine facility. The work described in this paper could be used by RA manufacturers to determine a final temperature rise based solely on dry and time-reduced testing."

Jan 1, 2017

By A. Konya

"INTRODUCTION Underground Mining and Construction projects face several challenges unique to the underground environment compared to the surface environment. The goals of an effective blasting program are the development of a blast pattern that minimizes cost, maximizes face advance, and provides the desired fragmentation and face profile. In addition to these , mines must consider the very real operational problems that occur such as drill deviation, ground vibration, and deadpressing or sympathetic detonation of explosives. These operation problems lead to increased costs and decreased production in the mine, hurting the bottom line. Combined with these problems, underground blasts can be difficult to monitor and evaluate due to tight spaces, dust during the blast obscuring video, and the need to evacuate the immediate area with no clear line of sight to observe performance. This creates an environment which most mines know that blasting can be improved in their underground operations, yet most are hesitant to change due to fears of the shot freezing, misfires occurring, and questions about how to achieve a proper design. The explosive companies and shot service providers also offer limited assistance or knowledge in the underground blasting realm, unlike with surface blasting, leaving many underground operations in the dark. It is for these reasons that most underground operations are operating decades behind the current technologies and incurring significant costs, both real and hidden, from there blasting operations. This paper discusses philosophical changes to designs that need to be made for effective design and modifications of underground blast patterns to prevent deadpressing and precompression of modern explosive products (Konya, Konya, & Worsey, Modern Blasting Method Selection, 2017). While numerous different methods of underground heading development exist, there are two types that are used in a majority of mines around the world, these are the burn cut and the V-cut. The burn cut is a parallel cut that has holes that go straight into the face and are primarily used in smaller opening (due to ability to pull deep rounds with smaller room widths) and metal mines (helps in ore dilution and greater fragmentation). The V-Cut is a cut that uses angled drilling towards the center of the face (in the traditional method) and is primarily used in underground aggregate operations due to its cheap cost but larger fragmentation sizes and greater throw of the muckpile. This paper will discuss these two forms of cuts specifically with the understanding that modifications to these can be applied to other parallel and angled underground heading techniques. BURN CUT The burn cut is one of the older forms of underground blasting, where a part of the blast (burn) of a shot is heavily loaded with additional empty (relief) holes with the goal of smashing (aka burn) and pushing/blowing all the material in the burn out to provide relief for subsequent blastholes. The relief holes in the burn cut provide a parallel relief for the holes that are firing to break too creating a second plane of relief, however these are extremely small and can cause many problems if a design is improper."

Jan 1, 2018

By S. K. Kawatra, S. Valluri

With increased CO2 emissions from steel and coal industry, there is great opportunity to capture CO2 and utilize the captured CO2 for economic advantage. The primary goal is to develop energy-efficient processes that reductively couple CO2, an abundant and renewable carbon source, for the production of value-added chemicals such as methanol, ethanol, and oxalic acid. These chemicals can be used elsewhere in the refining process or sold as valuable by-products. Geological sequestration, in comparison, has no economic return. Current CO2 utilization research at Michigan Technological University focuses on electrochemical reduction of CO2. Electrochemical reduction of CO2 to hydrocarbons and other chemicals is a complex multistep reaction with adsorbed intermediates. The exact reaction mechanisms leading to various products are not clear from the literature to date and will likely change over the range of conditions like pH, electrode potential, electrolyte medium, catalyst, etc. This paper will present different ideas and their viability to utilize CO2 as a feedstock to produce several value-added compounds. INTRODUCTION Carbon dioxide (CO2) emissions into the atmosphere will need to be drastically reduced to curb the various undesirable effects of climate change. The large contribution in total CO2 emission originates from coal or natural gas power plants, and a considerable amount from steel plants. Average CO2 emissions in the US from steel industry is 2.9t CO2/t steel (Global CCS Institute 2013). Capturing the available CO2 from the steel and coal industries for economic advantage is a win-win situation for the industry. This technology is not only important scientifically but is also vital for a sustainable future. The various ongoing investigations can be categorized as biochemical, thermochemical, photochemical, and electrochemical approaches. Among these, the electrochemical method shows the most promise as an efficient form of CO2 conversion technology, because of its many advantages like high reactivity under ambient conditions and good extensibility from small- to large-scale processes1 BACKGROUND Extensive research has been done on the chemistry of transforming carbon dioxide into more useful products. Methods are known to convert carbon dioxide into a wide variety of substances, including methanol, isobutanol, carbohydrates, methane, carbonates, urea, formic acid, oxalic acid, carbon monoxide, epoxides, formaldehyde, and so on. Several of these (carbohydrates, formaldehyde, isobutanol etc.) are primarily results from biological processes. The rest are results of strong reduction reactions or electrolytic reduction. In short, electrolytic reduction can be used to form methane, methanol, formic acid, oxalic acid, carbon monoxide from carbon dioxide2. CO2 reduces at the cathode in an electrolysis cell. These processes have the general form of generating the carbon dioxide anion radical (2⋅−) and allowing it to react with itself or the electrolyte3. Catalysts can be added to influence the formation of the anion radical or to suppress side reactions3. The electrolyte, catalyst, voltage, electrode material and CO2 content are all known to effect the reaction pathway. Table 1 shows the overall reactions and energy requirements for electrochemical conversion of CO2 to chemicals

Jan 1, 2019

By Patrick J. Maessen, Bart H. M. Hendrix

"In the Netherlands, the Dutch Tunnel Standard has been developed for the execution of road tunnel projects. The Dutch Tunnel Standard is based on the Dutch Tunnel Law (WARVW) which sets the minimum safety standard for road tunnels in the Netherlands. Uniformity of tunnels and the uniform perception of a tunnel by its users leads to a higher safety level. Therefore, several items in the Dutch Tunnel Standard are specifically developed to ensure safety in emergency situations through the plane of human factors. Taking human factors into account is important in regular traffic situations, but in an emergency situation, this can be vital. In this paper, the development and contents of the Dutch Tunnel Standard are explained in the first part and in the second part, the systems related to human factors are described. DUTCH TUNNEL LAW AND THE NEED FOR THE DUTCH TUNNEL STANDARD In the 90’s of last century and the early 2000’s, several serious accidents occurred in tunnels in Europe. These accidents involved fire and led to many casualties. After these accidents, the European Union decided to change the regulations for tunnel safety, leading to the EU directive 2004/54/EG. The primary goal of this directive is, that casualties are minimized in case of a calamity in a tunnel (e.g. fire). In the Netherlands, the EU directive led to the development of the tunnel law and a tunnel regulation (in 2006) to prescribe the safety features required for a safe tunnel. The new Dutch tunnel law – cited in the Building Act – was meant to make tunnel design easier – but resulted at first in the opposite effect. Tunnel design in the Netherlands and design standards of Rijkswaterstaat Rijkswaterstaat – the executing organisation of the Dutch Ministry of Infrastructure and the Environment – is the owner of the tunnels in the Dutch national motorway network, thus owner of the majority of the road tunnels in the Netherlands. In the early years of tunnel building in the Dutch national highway network (approximately 1950 – 1970), Rijkswaterstaat was not only owner, but also designer of its own assets. Most Dutch tunnels are submersed tunnels, because these type of tunnels provide a cost-effective way to make river crossings. Another type of tunnel that is used, is a TBM tunnel. Because of the geological conditions in the Netherlands (merely clay/peat and sand), there are no NATM or drill-and-blast tunnels in the Netherlands. Since the last decades, land tunnels (covered roads) are becoming more common to provide for multiple use of surface space and to meet more modern environmental regulations as traffic increases near densely populated areas. Tunnel lengths in the Netherlands vary from 500 m to 1.5 km for river crossings up to 6 km for the Western Scheldt (bored) tunnel. The lengths of land tunnels vary from 500 m up to 3 km. With these increasing lengths, some safety topics and design topics arise, topics that were non-issues in the past, when tunnels were only used as river crossings."

Jan 1, 2016

By Stanley W. Ivosevic, Mark E. Emerson

Introduction Metals exploration activity is moderately intense in the US despite the 1981-82 recession in the mining industry. Although exploration expenditures are significantly less than the 1981 peak, quality and goal orientation of current exploration are high. Major foreign and domestic mining companies as well as nonmetal companies had been attracted to invest in minerals in the US. Its capital markets, political stability, and open-door policy are attractive for hard mineral explo¬ration, particularly in the West. The approach was to either acquire fully or partially developed ore bodies through merger and acquisition, or to fund large scale exploration programs in search of bulk-tonnage deposits. The acquisition avenue has not proved the panacea envisioned for overcoming lack of substantive mineral development. Nor has the decentralized approach to exploration that was adopted proven cost-effective. The emphasis on gold, to the exclusion of all other commodities, raises questions regarding the direction of exploration in the US for the rest of the 1980s. Overview More than 300 companies are active in metals exploration in the US. These range from established, well-known domestic and foreign corporations to hurriedly incorporated, unknown companies financed by having promoted mineral prospects or occurrences of questionable quality. Junior companies tend to be oriented toward property entrepreneurship, rather than the development of minerals resources. These companies have helped maintain mineral lands from prospect to production by holding blocks of property intact during periods of exploration inactivity. This was done through promotional efforts, even though few properties that attracted major venturers proved to have real merit. Exploration objectives of senior and junior exploration companies currently embrace only a few commodities. Arranged in approximate descending order of priority, these are: precious metals; base metals in monometallic or polymetallic deposits; strategic metals, such as tungsten, tantalum, and tin; unconventional commodities, such as diamonds; ferrous metals; and commodities such as beryl, zeolites, and barite. Prospectors are important in searching for these targets. In the US, where they have been financed by major and junior companies, prospectors have proven a useful and low-cost means of locating gold prospects (e.g. the Alligator Ridge property in Nevada). The tenacity and persistence of independent prospectors have been factors in discovering properties rejected or overlooked by professional geologists (e.g. Lornex mine in British Columbia). This has been realized by the provincial government that grubstakes bona fide prospectors. Exploration interest in commodities has always been cyclic with price. Gold is one of the few metals that has experienced a price increase of such magnitude to warrant attention. The potential for realizing adequate operating margins for the Nevada open-pit "no-see-um" gold deposits is great. The present declining gold price, however, may indicate that this emphasis on gold, to the exclusion of other commodities, is shortsighted. Mining companies have been the last to react to price increases, opening new mines based on unrealistic market expectations. A scenario could develop where plummeting gold prices bring recession to the gold industry, as lately occurred in uranium, molybdenum, and copper. US Exploration Climate The US is attractive for exploration due to its strong mineral po-

Jan 4, 1984

By Paul L. Russell

Numbers, arithmetic, mathematics. Where did it all come from? Who needs it? We all do, but many school children would not agree. Engineers, physicists, accountants, tax collectors, and most of the general public are well aware that modern society in large part is founded on mathematics and could not exist without it. Mathematics developed over thousands of years as part of the everyday life of man. No one area, tribe, or race can claim discovery. In fact, from the records that do exist, it is evident that many facets of mathematics were discovered independently of others who made similar discoveries. While proof may be lacking, the earliest concepts of mathematics may be some 250,000 to 300,000 years old. The first use of numbers as written records is unknown. However, a bone of a young wolf deeply cut with 55 notches-arranged in two rows, one of 25 notches, the other of 30-was found in Czechoslovakia. Dated to be about 30,000 years old, it is perhaps an early record of some kind of counting system. It is assumed that mathematics was developed as an answer to man's practical needs, but anthropological studies suggest that the art of counting may have developed in connection with the needs of primitive religious ceremonial rites. Fingers probably were first used to designate sets of from one to five objects. Using both hands, up to 10 items could be indicated. As civilization developed, so did the need for more flexible systems to record items of trade, measure land, and figure taxes. Aristotle noted that the development of the decimal system comes from the fact that most people are born with 10 fingers and 10 toes. In spite of this obvious fact many early civilizations used binary, ternary, and quinary systems while vigesimal systems (with 20 as a base) and even sexagesimal systems (with 60 as a base) were used by others. A form of writing was in use in both the Mesopotamian and Nile valleys as early as 3000 to 4500 B.C. In Mesopotamia, wedge-shaped marks were impressed upon soft clay tablets that were hardened in ovens or by the sun. Many thousands of such tablets have survived. When the tablets were finally deciphered, some were found to be records of items shipped in trade, inventories, tax accounts, and mathematics exercises for teaching. Mesopotamia is also the source of an interesting board game dating back to about 2500 B.C. Its buttonlike men and triangular dice were apparently used with squares of shell, inlaid with lapis lazuli and red sandstone. Unfortunately, rules of play were not found. In Egypt, record preservation was somewhat better. Discovery of the Rosetta Stone in 1799 eventually led to deciphering the Egyptian hieroglyphics that included a numbering system at least 5,000 years old. These records show a competence in counting and numbering. A calendar of good accuracy may have been in existence as early as 4,200 years ago. Mathematics, as practiced today, developed when the Greeks added logic and proof to arithme-

Jan 12, 1982

By D. W. Gentry

Introduction Until recently, there were many jobs for mineral engineering graduates. Traditionally, minerals-related career paths available to university graduates stem from rather standard offerings in the disciplines of mining, metallurgy, petroleum, geology, geophysical engineering, and mineral economics. Within each discipline, there exists many career opportunities. In fact, until fairly recently, the most difficult decision for most graduates was choosing among all the available alternatives. Unfortunately, this relative state of euphoria changed quickly and dramatically within the minerals sector. Within two or three years, graduates from minerals-related engineering disciplines found it exceedingly difficult - and then almost impossible - to secure employment in the minerals sector upon graduation. At the same time, there were dramatic reductions in production requirements and, ultimately, mine closings. This resulted in massive layoffs and often termination of professional personnel possessing significant and valuable industrial experience. Obviously, this worsened an already bad employment picture for those hoping to begin a career in minerals. Certainly, there is no universal agreement on what the future offers for the US minerals industry or the career opportunities associated with it. However, there are a number of considerations that deserve study by students, faculty, administrators, and industrial personnel before minerals-related career paths are chosen. Where Will the Jobs Be and How Many? Everyone is aware that the minerals industry is very cyclic. It is not surprising then that job opportunities for new graduates also fluctuate. Typically, there is a two-to-four year lag between industry's engineering manpower needs and the academic community's ability to produce this talent. Due to the response lag, there will necessarily be years when supply and demand do not match. Employment Projections It is extremely difficult to anticipate what the key career paths might be within the next few years. Considerable speculation remains with respect to whether activity in the US minerals sector over the short-, intermediate-, and long-term will improve, remain status quo, or further deteriorate. However, some information pertaining to the prospects for career opportunities in mining engineering was offered by Tingley and Ghose (ME, Sept. 1983, p. 1274) from a study performed by Fenvessy and Schwab. Results from this industry-sponsored survey suggest that the US will experience 700 more mining engineers than needed in 1985, 1300 in 1990, and 1800 in the year 2000. These conclusions were based on the assumptions that future graduate levels remain flat and demand increases 3.6% annually. Educators quickly argued that depressed job markets, coupled with university student mobility, would negate the assumption of flat mining graduate levels in the immediate future. Student Mobility Student mobility from one engineering discipline to another is high. This is perhaps best demonstrated by the tremendous growth in programs such as computer engineering and computer science at the expense of enrollments in the more traditional engineering programs. The ability and willingness of students to quickly change engineering disciplines in response to existing and perceived job market demands have created significant student, faculty, and resource dislocations within universities nationwide. This has resulted in problems for university faculty and administrators responsible for effective use of

Jan 2, 1985

By Robert W. Wheeler, Richard G. Ames, Robert B. Reger

Introduction The use of diesel engines as an alternate power source in US underground coal mines has been a controversial step towards more efficent mining practices. Three areas of potential benefit are foreseen. First, diesel-powered mining equipment is believed to be more efficient, resulting in less occupational risk for any given amount of production. Second, diesels are believed to pose fewer intrinsic hazards, especially those related to fires and ignition of combustibles. Finally, several other health and safety advantages are foreseen for diesel units. Each of these issues will be discussed separately. Diesel Efficiency Proponents of the use of underground diesel-powered mining equipment point to increased production efficiency due to greater maneuverability, speed, power, and payload (Bradbury, 1973; Jorgensen, 1982; Ogden, 1978; Wenberg, 1979; Olsen, 1980). Units most suitable for diesel power include LHD vehicles, shuttle cars, personnel carriers, and maintenance vehicles. Estimates of increased efficiency vary from 60%-300% and more (Wenberg, 1979). Increased diesel mining efficiency is usually viewed as making it practical to mine lower grade seams. Diesel units can also operate over a wider range of floor pitches and distances than electrically-powered units, thus allowing mining of seams with special mining problems (Olsen, 1980). Diesel application is not a universal substitute for electrical power. For example, continuous miners are unlikely diesel candidates. Reduced Intrinsic Hazards Proponents of diesel use have pointed to the increased safety potential due to a reduction in electrocutions, arcs, burns, underground cable fires, and methane ignition (Ogden, 1978). Diesel units are frequently employed in maintenance activities for lifting and moving, and for coal cleanup. Thus the case is made that diesel use, by making powered operations more conveniently available, reduces the incidence of lower back injuries and related causes for days lost due to disability (Olsen, 1980). Several other issues relating to diesel safety have also been suggested. Diesel units can continue operation even during power outages allowing, for example, personnel evacuation. The maneuverability of diesel units make them particularly attractive in difficult physical conditions, such as long haulage from the face to the belt and in areas where water is deep. The case for diesel use focuses heavily on the potential for a "per worker" reduction in accidents, or a reduction in the severity of accidents. Since accidents are a major part of the health impact of mining, a reduction in accidents and resulting injuries would make a major contribution to the health of coal miners. Opposition to Diesel Use Opposition to the use of underground diesel mining equipment in the US seems to have started with the United Mine Workers of America. Some opposition to US diesel use has been based on a widely publicized 1976 letter from NIOSH to Mine Enforcement and Safety Administration administrator Robert Barrett (NIOSH, 1982). In essence, this letter warned that while insufficient evidence existed as to diesel effects on underground coal miners, NIOSH had concerns based on prior knowledge of separate diesel exhaust constituents and the potential for interaction between these constituents and coal mine dust. Current opposition to underground diesel use has focused on respiratory health effects related to diesel emissions, particularly nitrogen dioxide (NO,), carbon monoxide (CO), and particulate content (Barrett, 1982). Through

Jan 4, 1983