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By V. B. Achanti, A. W. Khair

Developing a laboratory procedure for the determination of coal-seam compressive strength has been a subject of discussion for many decades, and the reliability of laboratory determinations is often the subject of debate. A coal seam’s characteristics, the method of sample collection, the specimen preparation method and the testing procedure can often result in drastic variations in the test results. In this paper, the development of a laboratory testing procedure for the determination of coal-seam compressive strength is presented. Two samples that were collected from different coal seams were tested in this study. First, the inherent characteristics, i.e., coal structure. Hardgrove grindability index (HGI), Slake durability index, proximate analysis and mechanical properties (Young's modules and Poisson's ratio), of each coal were determined. The results were used to categorize the coal scams in terms of the characteristics liable to produce variable strength data in typical laboratory tests. Second, two commonly used laboratory specimen shapes, cylindrical and cubical, were tested to determine which shape produced the most reproducible strength values. Third, four loading rates within the static testing range, i.e., 0.007, 0.07, 0.34 and 0.7 MPa/s (1.0, 10, 50 and 100 psi/sec), were used to determine the ideal loading rate. Finally, the number of test samples required to produce strength data within the acceptable standard deviation, considering coal seam characteristics, were established. Background Laboratory testing of coal specimens for the determination of seam compressive strength has been a subject of discussion since the early 1900s. Many researchers have tested coal specimens to estimate coal-seam compressive strength for use in pillar design to insure the safe operation of [ ] mines (Unrug et al., 1986; and Hirt et al.. 1992). Yet, there is a lack of a universal procedure, and, no matter which approach is used, it becomes the subject of intensive debate. A great deal of work has been performed on the subject, and there is much information available in the literature. Some of this work concerns quantifying the influence of specimen shape and size on compressive strength (Bieniawski. 1968: Barron et al.. 1992: Khair and Xu 1994; and Khair 1994). As indicted in the literature, coal is a heterogeneous material, consisting of discontinuities such as cleats and bedding planes, which are likely to produce variable strength results (Holland, 1964; Unrug et al., 1986). Some of the prominent issues that influence the determination of coal-seam compressive strength in the laboratory are: •the collection of a representative sample from a coal seam; •the geometry, i.e., shape and size of the specimens to be prepared and tested: •the loading rate at which the specimens are to be tested:

Jan 1, 1997

By Timothy J. L. Haley

The population of the Rhondda underwent a spectacular rise as energy requirements increased in the mid-nineteenth century. The rapid expansion in the demand for coal transformed the population from a few hundred sheep farmers in the eighteenth century, into a thriving coal-based economy of some 168,000 inhabitants in 1924. Cochrane (1958) put forward the idea that the differential migration of more healthy men from the Rhondda Fach during the period since the peak of population in 1924, had resulted in a respiratorily-impaired residual population. This idea was used to explain the reduced survival of miners & ex-miners aged 55-64 years from this valley as compared with a similar group from Leigh in England. Using the proportional hazards regression model of Cox (1972), a comparison is presented of the exact twenty-year mortality of migrant and non-migrant miners & ex-miners and non-miners from the Rhondda Fach. The analysis makes allowance for the pattern of migration throughout the twenty-year period and adjusts for the confounding influence of variables such as age, occupation, and category of Simple Pneumoconiosis. The results indicate that contrary to expectation, any effect of migration upon mortality is manifest only at higher ages. Furthermore, for these men, the migrants have worse survival than the non-migrants. Also, contrary to expectation in the light of work by the British Coal Board (Soutar et al, 1982) there is no evidence of increased mortality amongst men with category I simple pneumoconiosis. INTRODUCTION [The Background] In September 1950, the Medical Research Council's Pneumoconiosis Research Unit at Penarth in South Wales initiated a study of lung disease in the smaller of the two Rhondda Valleys - the Rhondda Fach. At that time, the primary interest was in assessing the possible role of pulmonary tuberculosis in the aetiology of the serious (complicated) form of coal-workers' pneumoconiosis, i.e. Progressive Massive Fibrosis. The Rhondda Fach is situated in South Wales, some 25 km north-west of Cardiff. It is a narrow winding river valley about 12 km long, and approximately 1 km wide. The name means Little Rhondda; there is a corresponding Big Rhondda - the Rhondda Fawr. Together, these valleys make-up most of the Rhondda Municipal Borough. Following the initial survey in 1950-51, there have been several follow-up studies of this population. In addition, amongst numerous other investigations to unravel the aetiology of coal-workers' pneumoconiosis, Medical Research Council teams studied various other (sample) population groups, particularly in the industrial areas of the adjacent Aberdare Valley, and in Staveley and Leigh in England. The agricultural areas of the Vale of Glamorgan in South Wales, and Annandale in Scotland were also studied.

Jan 1, 1993

By Jasinder S. Jaspal

The United States has very large coal reserves in seams that are less than 1065 mm (42 in.) thick. These reserves are often not mined because of the difficulty of working in low heights and, in most cases, the seams are beyond the technical or economical limits. It is estimated that the remaining reserves in 355 to 710 mm (14 to 28 in.) thick seams that are not faulted or intruded, dip less than 15° and under a cover of 610 m (2000 ft) or less are nearly 910 Gt (1 trillion st). This is a very large percentage of coal reserves in the United States (Unrug, 1985; DuCarme et al., 1990). The US Bureau of Mines is attempting to develop new techniques for mining 150 to 1065 mm (6 to 42 in.) thick coal seams. Research has focused on existing thin seam mining techniques, testing of small- and full-scale physical models of the Bureau's slot mining concept and computer modeling of the full-scale process. In 1986, only 21 % of the total US coal production came from -1065 mm (-42 in.) thick seams. Of that, 355 to 710 mm (14 to 28 in.) thick seams share of production was minimal, probably 5% or less (Mayercheck et al., 1990). This shows that the thin seams, particularly the seams 355 to 710 mm (14 to 28 in.) thick, have not been mined to the extent of the available reserves. Thin seam coal, generally, is of premium quality. As thicker seams are depleted, the mining of thin seams could become more important. Thus, there is a need to develop technology to mine these seams (J. Farinelli et al., 1987). A successfully applied technology could change reserve categorization. In the Appalachian Plateau there are 417 Gt (460 billion st) in 355 to 1065 mm (14 to 42 in.) seams. This represents nearly 9.2% of the US remaining reserves (Unrug, 1985). At depths of 152 m (500 ft) or less, the Appalachian Plateau has about 214 Gt (236 billion st) of reserves in 355 to 1065 mm (14 to 42 in.) thick seams. Other mining areas also have substantial coal reserves in this seam thickness category. Those, however, are not as attractive for mining as the Appalachian Plateau. In the past, most of the reserves of the Appalachian Plateau region were partially worked by contour and highwall auger mining. It is estimated that these mining systems had left some 40,230 km (25,000 miles) (Readers Digest, 1987) of unreclaimed highwalls. The Bureau recognized this situation. In FY1987, the agency initiated a research project to develop anew technology for mining thin seams of the Appalachian Plateau.

Jan 1, 1992

By K. N. Han, H. C. Cho, R. Kim

"As an extension of the work on the effect of OH- on various cations in hydrometallurgical systems, the effect of Cl-, Br-, I- and CN- on cations such as Ag+, Cd2+, Hg2+, Cu+, Cu2+, Co2+, Zn2+, Pb2+ and Sn2+ has been examined in detail. The Newton-Raphson technique has been used to determine the amounts of various metal species as a function of the concentration of anions. Some practical implications of various metal complexes in metal separation and removal from solutions have been examined and discussed.IntroductionUnderstanding the role of anions in hydrometallurgical processes often determines the success of an operation. The successful choice of concentration and type of anions not only leads to an effective extraction and removal of metals (Esmaeili et al., 2005; Brbooti et al., 2011; Wei et al., 2005; Jeppesen et al., 2009; Al-Tarazi et al., 2004), but also to assisting effective separation of various metals, whether it is a solvent extraction, ion exchange operation or even a simple precipitation (Farha et al., 2011; Sevenich and Fritz, 1983; Rashid, 1974).This paper continues the discussion of the role of anions in hydrometallurgical systems beyond the role of OH-, which was the topic of a previous paper (Kim et al., 2014). Table 1 lists various anions that are frequently used in conjunction with metal ions to be selectively separated or precipitated. As seen in the table, many anions are used frequently in hydrometallurgical practices. Some of the more frequently used anions include OH-, SO4 2- and CO3 2-. The halogen elements, Cl-, Br- and I- , as well as CN-, are also used in hydrometallurgical applications, especially in relation to noble metal recoveries (Meng and Han, 1994; Zhang et al., 1996; Sun et al., 1996; Starovoytov et al., 2007)."

Jan 1, 2015

By N. B. Prenn

Case Studies in Open Pit Design Using Whittle 4D will be discussing the following four examples: 1) Optimization of process plant design using Whittle 4D as a sizing tool. In this case, the home office and project team had different views of the size of the final project. The home office wanted to reduce the size of the project to reduce capital expenditure, while the project team suggested a larger project. Whittle 4D was used to determine Net Present Value and Internal Rate of Return using the Whittle scheduler (FDAN) to schedule many alternatives, and calculate the NPV and return for each alternative. 2) Comparison of pit optimization methods: Lerchs-Grossman vs floating cone. Many pits are developed based on floating cone economics. A comparison of methods in one large gold mine resulted in the stripping ratio of the pit being reduced by 10% in the Whittle 4D optimization as compared to the floating cone optimization. 3) Pit optimization to determine if a new mill is justified. A large heap leach gold and silver mine needed to determine if a new mill was justified. The planning and scheduling would take several man months if completed without pit optimization and scheduling software. Whittle 4D software proved to be very effective in providing optimized pits and a mine schedule in a short time frame. 4) Pit optimization to determine how to mine and size. Open pit optimization determined where the underground operation should start, and whether the material should be processed by milling or heap leaching. One of our clients asked for help in determining where the pit should end and a new underground operation start. We used Whittle 4D to help determine where the pit should end and where to go underground. In addition the operation needed to determine how to process the material.

Jan 1, 1996

By T. M. Barczak, Zach Agioutantis

The Support Technology Optimization Program (STOP) has been developed by the National Institute of Occupational Safety and Health (NIOSH) to provide a comprehensive tool for evaluating the performance of available secondary roof support systems and designs based on support strength, convergence, cost, installation time and material handling considerations. This program uses engineering models developed from full-scale testing of numerous supports under controlled loading conditions in the unique NIOSH Mine Roof Simulator load frame to determine the support load-displacement performance as a function of the various sup- port design parameters. Both standing supports, such as wood cribs installed within the entry, and intrinsic supports such as cable bolts, can be evaluated. Support systems can be designed and compared for use in a new mining section or to evaluate potential replacements for an existing design. The support design requirements are defined from 1) estimating the failure height in the immediate roof, 2) in-mine measure- ments of support and strata interaction or development of ground reaction curves from numerical modeling, 3) the performance of an existing support system, or 4) from any arbitrary load and convergence criteria. All settings can be saved to permit resuming an evaluation in progress at a later time. This paper presents the updated MS Windows version of this tool, which can be easily installed and run on a personal computer. This Windows implementation will feature a project and application scenario setup where multiple scenarios can be assigned to one project. Scenarios can then be compared on support parameters and cost. Cost parameters can be provided either as individual cost items or using a price grouping feature. Additional support performance curves and/or additional ground reaction curves will be distributed with program updates as new information is developed. Users can also provide their own support and/or ground reaction curves.

Jun 25, 2024

By Jodie Robertson, Isabel Barton, Cristian Caro

Prior to mining, all rocks contained in a mineral deposit must be placed in one of several categories known as ore types. Each ore type describes a different way that the rock will behave in processing. The ore type determines which process stream a particular mined unit will be sent to, sets expectations for process engineers and metallurgists, and forms the basis for recovery predictions and production estimates. At first, ore types are assigned mainly by geologists based on visible geological criteria (dominant ore mineralogy, estimated grade, rock type, alteration type). As metallurgical tests are carried out and process data accumulate, the resulting criteria may be factored into the ore typing process, typically as assays and recoverable metal using the relevant process(es). Other variables may also be incorporated into ore typing at operations if they are particularly important, such as grindability, preg-robbing, gangue reagent consumption, and the presence of especially deleterious minerals such as talc or smectite. The mine operation department may also redefine ore types and/or process destinations based on economic criteria. A “best” ore type is designated that (in theory) best describes how the rock will behave in processing. Codes representing this ore type are entered into the drill-hole database, then interpreted on cross sections and level plans, and finally added to the block model. Ore types are iteratively reexamined and adjusted over the life of a project. Major difficulties in effective ore typing are (1) lack of communication between geologists and metallurgists, (2) the imprecision inherent in imposing clean categories on messy natural systems without discrete divisions, (3) the need for ore types to provide useful information for a wide range of processes, (4) difficulties in scaling up laboratory test results to mine production, and (5) conflict among geological, metallurgical, and economic criteria for ore typing.

Jul 1, 2024

By John L. Schroder

The design of an underground mining operation requires the integration of transportation, ventilation, ground control, and mining methods to form a system which provides the highest possible degree of safety for mine personnel; the lowest cost per ton of product; the most suitable quality of final product, considering its intended use; the maximum percent recovery of the reserve, consistent with cost and subsidence rights; and the minimum disturbance to the environment. The selection of mining method is generally the controlling factor in overall mine design. Some important considerations in the choice of mining method are : 1) Depth of cover over the coal seam. 2) Strength and thickness of rock strata in the overburden. 3) Characteristics of the coal seam, including thickness, nature and location of impurities or irregularities, dip (or pitch), nature and direction of cleavage, hardness, and structural strength of the coal. 4) Nature and strength of the immediate roof (top) over the coal seam. 5) Nature and strength of the floor (bottom) under the coal seam. 6) Presence and amount of explosive gas (methane). 7) Quantity of water likely to be encountered in the mining operation. 8) Any previous or concurrent mining in seams above or below the coal seam to be worked. 9) Any restrictions on recovery required for subsidence control. Topographical and geological reports and maps of the area are usually available from state and federal geological surveys, and contain excellent general information to guide more detailed investigations. Exploratory drilling, and the plotting and analysis of the information thus obtained, are essential steps in any development program, and aid in the correlation and evaluation of information gained by other means. Other good sources of information are operating coal mines that may be in the same seam in nearby properties. Such mines may indicate in advance much about roof control, drainage, ventilation, and other mining problems.

Jan 1, 1981

By Staff

Assembling consistent data for a 40-year time series is a challenge because of the many changes that have occurred in this period. In many cases, because of process improvements and consumer requirements, grades of materials change in a 40-year span. This can lead to a wide fluctuation in price data simply because of increased costs for premium grades. Sources for obtaining price data can also change during a 40-year period. Changing the data source also introduces anomalies in a time series. In some cases, published price data are not representative of the real value of the material and provide little basis for further analysis. For some industrial mineral commodities, there are only a few producers, and value and production data are considered proprietary. In the data presented in this chapter, every attempt was made to ensure that the data were as consistent and representative of the industry as possible. However, the factors mentioned above may have introduced inconsistencies in some of the data series. Many different factors influence trends for industrial minerals because of the diversity of end uses for these commodities. Trends in production and prices for most of the industrial minerals can be related to one or more of the following: changing demand or technologies in end-use industries, development of new end uses, the effect of environmental regulations, production and processing technology changes, energy costs, foreign competition and substitution by other materials, and production as a byproduct of other minerals. World economic conditions, such as recessions and inflation rates, also play a part in some production and price trends. Data for 40 years of production for industrial minerals in the United States are presented in [Table 1] in 10-year increments. Unit values for the industrial minerals, in actual and constant 1990 dollars, are shown in [Tables 2 and 3].

Jan 1, 1994

By Christopher Haycocks

Self-supporting, open stoping underground mining methods are some of the earliest, starting with the prehistoric flint mines of Europe and the Egyptian gold mines of Nubia in the time of the pharaohs. When the early miner created an opening, it had to be inherently stable and self-supporting to enable him to continue. Therefore, as mining progressed beyond simple gophering, self-supported open stoping methods were devised to meet the needs of individual ore bodies. Timber for support appears in mines as early as 1000 BC, but this was for no more than local ground control. The ingenuity of mining engineers through history has so refined open stoping methods that today they produce more tonnage than any other mining system. Self-supported open stoping finds its application in moderate to strong rocks that exhibit little jointing or fiagmentation. Good back, hanging wall, and footwall conditions that require no more than incidental support are an essential part of these mining methods. Open stoping methods find applications in all inclinations of ore body from flat to vertical, and where ore movement may be either mechanical or under the influence of gravity. Table 18.0.1 demonstrates the range of ore bodies suitable for open stoping, with particular reference to their dip and the deposit thickness. Most of the flat-dipping applications use ore pillars for overall support and ground control that may be recovered under some conditions after completion of normal stoping practices. In steeper ore bodies, rib, crown, and sill pillars used to block out the ore may be recovered subsequent to the completion of stoping. Open stopes are frequently filled to facilitate pillar recovery. Two of the methods, vertical crater retreat (VCR) and shrinkage stoping, utilize fragmented ore for temporary support in the stopes. In these two cases, the stope is not strictly open until after drawdown.

Jan 1, 2008

By Orville L. Lantto, Bruce K. Cantrell, Keith S. Olson

Approximately 10 to 20 pct of exhaust aerosol mass from diesel equipped vehicles operating without exhaust controls is attributable to particles in the 1.0-10 µm size range. Recent in-mine measurements suggest that, in addition to reducing overall exhaust aerosol mass concentrations, the use of waterbath exhaust conditioners (water scrubbers) may significantly reduce the concentration of diesel exhaust particles larger than approximately 0.5 µm aerodynamic diameter. This implies that a measurement system that discriminates between diesel and mineral dust aerosols on the basis of aerosol size would have different accuracies depending on use of a water scrubber. To test this hypothesis, laboratory measurements were made by the Bureau of Mines to determine diesel exhaust aerosol penetration efficiency through a water scrubber as a function of size and engine operating conditions. Size distributions were determined using a Micro-Orifice, Uniform Deposit Impactor. Concentrations of CO, CO2, NO x, O2 and gas-phase hydrocarbon concentrations were also measured. No significant effect was found for these gases other than a slight decrease in NOx emissions. The water scrubber served as a moderate aerosol control with reductions in mass concentrations of about 29±9 pct under most conditions. This reduction occurred for particles less than 0.2 µm and greater than 0.6 µm aerodynamic diameter. Coal dust aerosols are mostly larger than 0.8 µm aerodynamic diameter. With use of a water scrubber, separating the aerosol into two parts, greater and less than about 0.8 µm aerodynamic diameter during sampling, results in collection of diesel and coal dust aerosol accurate to within 10 pct. Without a water scrubber in operation, such a sampling protocol would only be accurate to approximately 15 pct.

Jan 1, 1991

By Charles G. Weakly, Alan M. Lander, Pamela B. Zohar, Robert W. Leonardson

The cluster of gold deposits at the Goldstrike property is located in the Northern Carlin Trend in northeastern Nevada (Fig. 1). At many mature mining camps like Goldstrike, exploration innovation and success may wane with time. In this regard, sufficient exploration drilling was conducted in the area immediately north of the ultimate pit crest and under the final planned pit wall to test for a major deposit. No indications of large deposits were found, however, the majority of drill holes did intercept mineralization. Mining along the northern pit wall exposed significant structurally controlled ore-grade mineralization that had been missed by exploration drilling. In 2002 and early 2003, the mine geology group added reserves to the pit and improved the definition of underground resources by targeting specific ore-controlling structures within favorable stratigraphic units by utilizing relatively inexpensive horizontal RC drilling from the pit wall. The widespread occurrence of mineralization to the north of the pit was re-interpreted to be an indication of continuity of these mineralized zones. In January 2003, 70 feet of 0.332 opt Au at a depth of 710 to 780 feet was intercepted in a dewatering well. This intercept was significantly thicker and higher-grade than average for the area, which confirmed the existence of high-grade structural continuity, and helped secure management support for additional exploration. A timely and cost effective solution was needed to confirm the potential of the area. Immediate follow-up drilling on tighter grid spacing was rejected in favor of improving the geology model. Exploration across the property would benefit from an improved geology model, but to provide results in a timely manner, the area to the north of the Betze pit was chosen as a starting point (Fig. 2).

Jan 1, 2005

By William A. Dupree, Mark J. Schultz, William J. Francart

Maintaining intake escapeways clear of contaminants from mine fires has been an issue for many years. Removing fire sources from these entries reduces the risk of contamination greatly. In recent years, concerns have been raised related to contamination of the intake escapeways by leakage when fires occur in adjacent airways. This paper discusses measured leakage rates for multiple stoppings at various pressure drops across stopping lines in a number of actual coal mines. Also addressed are methods for reducing the contamination risks for the intake escapeways, considering the ventilation system employed. The Effects of Air Leakage The effects of air leakage through ventilation controls has been well documented. While normal leakage is expected through stoppings, overcasts and other structures, abnormally high leakage results in an inefficient ventilation system and can increase operational costs for a mine. In addition, excessive leakage results in less control over mine ventilation airsplits. This may compromise the isolation of mine escapeways, and allow adjacent entries to be rapidly contaminated by products of combustion in the event of a mine fire. A study was conducted at ten mines in which air pressure and quantity measurements were taken near the active sections. The method used to construct the ventilation controls to separate airways was also noted. Stopping Construction Table 1 shows the type of stopping construction commonly used in the 10 mines included in this study. Main entry stoppings were normally built with mortared joints (wet stacked) with sealant applied to the high pressure side of the stopping. In most cases the entire face of the stoppings were sealed. In one case only the joints between blocks were sealed but the perimeters were sealed in all cases. Some mines also sealed the low pressure side of the stopping. Solid blocks were normally used for main entry construction and hollow blocks were sometimes used for longwall panel development. [ ]

Jan 1, 1993

By Gerald L. Finfinger, Michael A. Trevits

Since 1964, the US Bureau of Mines has been conducting a comprehensive program of research on the origin and control of coal bed gas. Several extensive databases have been developed during the work representing information from every major coal producing region in the United States. Gas content measurements have been made on 885 coal samples from 15 states representing 221 coal beds. Records indicate that the range of gas content varies from zero to over 640 ft3 per ton (20 cm3 per gm). During the conduct of the research program, several drilling and completion techniques have been developed or refined to divert and collect coal bed gas before it enters the mine environment. Such techniques include horizontal, cross-measure and vertical boreholes. Samples of gas collected from boreholes drilled under the program indicate that the primary gas constituent is methane which comprises 97 pct, the remaining hydrocarbons make up approximately 0.2 pct. Nitrogen and carbon dioxide account for the majority of the remaining gases at average concentrations of 1 pct each. Surveys have been made on the total daily gas emissions from underground coal mines. In fact, the most recent survey indicates that 180 mines had emissions in excess of 100 Mft3 per day (2.8 Mm3 per day). Furthermore, total mine emissions were in excess of 300 MMft3 per day (8493 Mm3 per day). An analysis of the data showed that there appears to be no relationship between in-place gas content and coal bed gas composition. Furthermore, no correlation was observed between in- place gas content and daily mine emissions. In a general sense, the higher the gas content the higher the relative gas emissions. Perhaps a more direct relationship would be observed if more complete emissions information (including gob gas borehole production) was available.

Jan 1, 1989

By Ben Ainsworth

I am going to discuss gold mining in Mexico during the last twenty centuries and try to suggest some of the futures that may develop in this industry in the next decade. Before I start this rapid review of such a long developing mining tradition, T must give credit to my sources. I an, talking here as an associate of Commoditi,c,s Reseflrch Unit Ltd. of London (better known as CRU) and as a principal of Ainsworth-Jenkins Holdings, an international mineral resource consultant group that has been exploring and deveJoping mining properties in Mexico and Central America for the last six years. CRU is the leading consultancy providing industry, governments and financial institu­tions with analysis and advice on the metals and the mining industries. They recently completed a three volume study entitled Gold Mining Opportunities in Mexico. This .<1tudy describes Mexico, the business environment, its geology and details its gold mining potential with property descriptions and maps. I have adopted and adapted the s11mmari.zed Mexican government statistics and rlrawn on some of the historical comments in these volumes. However, I do not have ti me (nor would it be appropriate) to be giving the details of the CRU work. Jorge Ordonez, whose shoes I am trying to fill today, was a major contributor to the study and it was an honour to be asked by Willem Beunderman, the Principal Editor, to step in when Jorge was unfortunately unable to be here. Jorge has been an enormous help to Ainsworth­Jenkins and a consistent friend of foreign investment in Mexico. Our own work in Mexico has allowed us to obtain a few useful in­sights into this most interesting slice 0f Mexican history.

Jan 1, 1992

By Craig Cina

Transporting goods from producers to consumers is a fundamental part of the US economy. The demand for transportation closely reflects the economic health and well-being of the nation&apos;s industrial, commercial, and agricultural sectors. The inland waterway barge industry has traditionally played a key role in transporting needed commodities to industries on an economical basis. Barges account for 10% of the domestic intercity tonnage carried by all transportation modes. Water carriers operating on inland waterways hauled more than 450 Mt (500 million st) of cargo in 1980-up from 428 Mt (472 million st) in 1970. Inland Waterway Shipments Metallic and nonmetallic minerals comprise one of the largest commodity groups moved on inland waterways. As illustrated in Table 1, metallic and nonmetallic mineral shipments accounted for more than 69 Mt (77 million st), or nearly 15% of all commodities shipped in 1980. Putting this tonnage in perspective, coal and petroleum represent the two largest commodities moved on inland waterways. In 1980, petroleum and petroleum products accounted for 138 Mt (152 million st), or 28% of all commodities shipped. Similarly, coal represented 120 Mt (132 million st), or 25% or all commodities shipped. Even though shipments of metallic and nonmetallic minerals are smaller than coal and petroleum, they still account for a sizeable portion transported. The inland barge industry moves a variety of metallic and nonmetallic minerals ranging from manganese ores to building cement. Minerals transported by barge include: • iron, copper, bauxite, and manganese ores for metal production; • lime and limestone for construction and electric utility scrubbers; • phosphate rock, natural fertilizers, and gypsum for agricultural needs; • sand, gravel and crushed rock, clay, asphalt, tar, and pitches, and building cement for construction purposes; and • a host of other minerals to meet industrial needs. A few mineral commodities are responsible for a large proportion of waterborne mineral shipments.

Jan 10, 1982

By James L. Vanderbeek

INTRODUCTION Semiautogenous (SAG) grinding technology first started to be seriously considered for comminution circuit design in the mid-1960s as an alternative to the three-stage crushing circuits and rod-ball mill circuits in vogue at the time. Since this time the application of SAG mill technology has become widespread with over 250 mills larger than 20 feet in diameter sold (Jones 2002). The vast majority of concentrators constructed over the past quarter century have utilized SAG mill technology, and a SAG-based comminution-based circuit is generally the first choice considered for new projects. Since the first generation of SAG milling technology was applied, many lessons have been learned, and new supporting technologies have emerged which make the design, installation and operation of SAG &-based grinding circuits more robust and efficient. This paper will discuss the current status of SAG and associated technologies. MILL MECHANICAL DESIGN ISSUES Table 1 summarizes the manner in which SAG design capability has evolved over the past 30 years. Increases in mill diameter came slowly as the industry took a wait-and-see attitude as the technology of larger mill diameters was advanced. The first 32-foot- diameter mill was ordered in 1969. It was not until 1979 that the technology jumped to the 34-foot-diameter mill and 1988 for the first 36-foot-diameter mill. The first 38- and 40-foot-diameter mills were both ordered in 1996. At least five 38-foot-diameter and two 40-foot-diameter mills will be in operation by mid-2004. Advances in structural analytical techniques and increased computational power, coupled with field measurements of large operating mills, have increased the confidence of the mill suppliers in their designs and the operating companies in the viability of larger diameter mills. No failures of a 36-foot-diameter mill or larger have been

Jan 1, 2004

By D. G. Krige

In the world mining scene, precious metals (particularly gold) have recently monopolized the attention in exploration as well as new production activities. As the major world producer of gold, South Africa plays and will continue to play a very significant role in this commodity. This role will, of course, have its consequential effects on the world&apos;s gold supply and, hence, also on the international price of gold. It is essential, therefore, and in the interests of all gold producers that the likely future pattern of gold production in South Africa be estimated such that some measure of production uncertainties can be predicted. For this purpose, some elementary geostatistical models will be used, and the main factors affecting the situation will be analyzed in broad outline only under the following headings: South Africa&apos;s basic gold ore potential in the three categories of: (1) ore remaining in existing mines; (2) extensions to existing mines via unexplored areas and/or reef horizons of lower grade; and (3) new mines, potentially deep or ultradeep. Cutoff grades - operating, marginal, and flotation, for the foregoing three ore categories, respectively. The availability of production, financial, and other resources required. Cutoff grades will, of course, depend on cost levels (operating and capital), future inflation and cost escalation trends in South Africa (SA), and/or gold price trends (in US$ and via the exchange rate in SA Rand terms). The differentials in inflation and escalation trends with those in overseas countries, mainly the USA, will, in general, set the scene for future exchange rate levels of the Rand. To predict the effect of various cutoff grades on future gold production from South Africa, it is essential to use geostatistical models. The applicability of such models is tested against past records of price, cost, and grade trends and then applied to the estimated ore potentials. These models were referred to briefly in a paper at a recent seminar on gold (Krige, 1986) but are discussed in detail herein. This lecture was intended originally to cover only technical aspects, but, in view of South Africa&apos;s position internationally and the possible effect of sanctions, internal unrest, and strikes on gold production, a brief reference to the third factor above (availability of production, financial, and other resources) seems appropriate.

Jan 1, 1988

By K. V. Raj, T. Bhowmick

"Computational fluid dynamics (CFD) modeling of an open-pit mine has several challenges. A good quality mesh of the model domain is a pre-requisite for convergence in solution. Good quality tetrahedral meshing and insertion of prism layers in faceted open-pit mine geometry is complicated due to presence of numerous vertices and ridges. Presence of vertices and ridges result in poor quality tetra elements and holes in prism layers, which cause instability in the model. Good quality pit bottom prism mesh elements significantly improve the formation of velocity boundary layer and thermal boundary layer at the pit bottom. This results in an expected temperature dependent micrometeorological buoyancy flow in the pit. Whereas, model geometry without a prism layer, results in abrupt temperature gradient near pit bottom. Buoyancy flow and recirculation have significant effect in lifting mechanism of fugitive dust particles generated near pit bottom. Representative initial and boundary conditions and selection of appropriate turbulence model have significant influences on the simulation results. In this paper, three-dimensional models of open-pit mines are developed using commercial CFD software (CRADLE-CFD). Important constraints and various consequences in modeling complex open-pit geometries are presented. INTRODUCTION Application of Computational fluid dynamics (CFD) to model an open-pit mine to predict fugitive dust and pollutant concentrations in and around a mine is relatively new. In former Soviet republic (USSR), Baklanov [1] and Belousov [2, 3] presented numerical models of pollutant transport in open-pit mines in Arctic regions. In recent years, advances in mathematical and computational fields have enabled better modeling of pollutant dispersion phenomenon around mines using computational fluid dynamics (CFD). Several CFD software packages have been used by various researchers to model the pollutant transport in open-pit mines. Works by Alvarez et al. [4]; Lowndes et al. [5]; Silvester et al. [6]; Torno et al. [7]; Choudhury and Bandopadhyay [8]; Collingwood et al. [9]; Raj et al. [10] and Flores et al. [11]; are examples of application of various commercial and open source CFD packages (COMSOL, ANSYS, OpenFOAM) to model pollutant transport in open-pit mines. These researches provided a strong background for additional research involving CFD modeling of open-pit mines. Due to the limitations of computational times and capacity, 3D modeling and various aspects of fugitive dust dispersion in open-pit mines remained unexplored."

Jan 1, 2015

By Gregory M. Rumbaugh, Christopher Mark

"The proliferation of unconventional gas well development in the Northern Appalachian coalfields has raised a number of mine safety concerns. Unconventional wells, which extract gas from deep shale formations, are characterized by gas volumes and pressures that are significantly higher than those observed at many conventional wells. The gas is composed largely of methane as well as other hydrocarbons. Hundreds of planned and actively producing wells penetrate protective coal pillars or barriers within active mine boundaries, including chain pillars located between longwall panels. Gas released from a well damaged by mining-induced ground movements could pose a risk to miners by flowing into the mine atmosphere. The mining-induced ground movements that may cause well damage include conventional subsidence, non-conventional subsidence (e.g., bedding plane slip), pillar failure, and floor instability. This paper describes the known risk factors for each of the four failure mechanisms. It includes a framework that can guide the risk assessment process when mining takes place near gas or oil wells.INTRODUCTIONSouthwestern Pennsylvania has served for centuries as a producer of coal, oil, and natural gas. Coal mines were first developed in the Pittsburgh Coal Seam in the 1760s, with the first recorded mine in 1761 being the Fort Pitt Mine (Edmunds, 2002). Natural gas was introduced to the region when the Haymaker Well in Murrysville, Pennsylvania experienced a blowout on November 3, 1878. Although it was originally drilled prospecting for oil, it produced an estimated 30 million cubic feet per day of natural gas. Without well controls or pipeline infrastructure in place, it burst into flames and lit the night sky for over a year. Soon after, the Penn Fuel Company continued development of the Murrysville gas field and is credited with making Pittsburgh, Pennsylvania the first major city in the U.S. to have natural gas supplied for home and industrial uses. “Natural gas is king in Pittsburgh,” proclaimed an October 18, 1885, article in the New York Times headlined “Value of Natural Gas, Its Use Clarifies the Atmosphere at Pittsburgh” (AOGHS, 2018). By the early 1900s, coal, oil, and natural gas production had reached the southern extents of Pennsylvania in Greene County. The county had first produced oil in 1886, and shortly after began producing natural gas in 1890. By 1906, over 1,300 wells had been drilled (Stone and Clapp, 1906)."

Jan 1, 2019