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By Thomas R. Kuesel, Harvey W. Parker

INTRODUCTION This paper describes the use of a shield and compressed air for construction of a short section of a pair of twin single-track tunnels for the Metropolitan Atlanta Rapid Transit Authority (MARTA) transit system. Although only two blocks long, this section posed the greatest difficulties to both engineers and contractors, and is the most expensive section of line construction on the MARTA system. All this derives directly from its setting--these tunnels had to be constructed through mixed-face conditions. SETTING The Broad Street tunnels form a link between the Peachtree Center and Five Points Stations, in the heart of Atlanta's central business district (Figure 1). Peachtree Center sits atop a solid rock ridge, which inspired the design of a deep rock-mined station, as described in the 1979 RETC. Five Points Station is MARTA's hub--the intersection of the system's two main branches. It is located in the old railroad gulch, a filled-in erosional valley at the base of Peachtree Ridge, and was constructed by open-cut methods. The alignment and profile of the Broad Street tunnels were closely constrained by these two adjacent anchors, and in the six-block stretch between them there was very little room for maneuvering. Since the rock ridge petered out about in the middle of the section, it was evident at the outset that there would be about two blocks of rock tunneling at the north end, two blocks of soft ground construction at the south, and two blocks of mixed-face in between. It is this central two-block section that is the focus of this paper.

Jan 1, 1981

By J. R. Muhm

Environmental due diligence investigations, also known as Phase I Environmental Audits, are increasingly important to major investors, venture partners, or prospective purchasers or sellers of mineral properties. Knowledge of environmental risks associated with a property will allow better appraisal of a property's overall desirability prior to completing the transaction. It is possible that the cost of environmental cleanup could exceed the value of the property. Topics appropriately investigated include permit status, water quality, past use of the property, hazardous materials management, storage tanks, public records pertaining to the environmental character of the property, and environmental costs.

Jan 1, 1991

By C. H. L. Bastos

A physical and chemical characterization study was carried out on 483 kaolin samples from 2.9 km2area at the Para Pigmentos mine along the Capim river, northern Brazil. The samples obtained from cores, drilled with regular spacing of 100 m, were evaluated using grit percentage, particle size distribution (psd), brightness, leachability, magnetic response and viscosity. These parameters were correlated with each process step in the plant and also the physical and chemical properties of the ore and finished product. Cluster analysis was performed and four main groups of kaolin were identified. These were then related to their performance at the various process steps and the quality of the initial product. These groups are well defined by plotting grit X “psd” and TiO2x “psd” in scatter plots.

Jan 1, 2004

By J. Kenneth Wittle

The DC Graphite Melter has been shown to be capable of processing radioactive waste materials. The process is characterized by its high temperature capability and operation in a closed system. From an environmental perspective, the control of emissions coupled with destruction for hazardous organic materials is mandatory. These same characteristics may be applicable to the minerals processing industry. Under normal operation, the system reaches temperatures for nominal operation of 1600°C, with an arc temperature in the order of 5000°C. The system temperature is determined by the melt temperature of the material being processed. In testing of a 750 lb/hour unit, temperatures of 2200°C were reached. Current focus is on finding waste treatment and metals recovery opportunities in the metals processing industry for his innovative technology.

Jan 1, 1999

By C. S. Harman, C. A. Enright, J. F. Brune

"The mining environment presents challenging and complex situations for the application of medical care. The combination of remote locations, difficult access and longer response times may create an increase in morbidity and mortality in the event of serious accidents. The U.S Mine Safety and Health Administration (MSHA) has only minimal requirements of first aid training for mine workers. For mine rescue team members, MSHA requires medical training at the Emergency Medical Responder (EMR) level, as only prescribed by mine rescue contest rules. This paper examines, based on mine accident statistical information, whether a higher level of medical training, for example at the Emergency Medical Technician (EMT) or at the Advanced Life Support (ALS) level can better address injuries most commonly disabling or leading to the death of miners in the United States. BACKGROUND Prehospital emergency medical care for the sick and injured has evolved considerably from first inception. Ambulances used to be staffed by persons with little training and equipment and prehospital care was focused largely on the rapid transport of the injured person to a hospital and. The modern prehospital care model was initially developed in military applications, where medics were responsible for some treatment on the battlefield and rapid transport to the care of a physician. In 1965, the National Academy of Sciences and National Research Council Committees on Trauma and Shock published Accidental Death and Disability: The Neglected Disease of Modern Society (National Research Council Committees on Trauma and Shock, 1966). This white paper laid out the consequences of the insufficiency of the Emergency Medical System (EMS) in the United States at that time. According to this paper, in 1965, 107,000 people were killed from accidental injury, with another 400,000 injured with permanent disability. The authors estimated that approximately 50% of ambulance services were provided by local morticians that were not capable of rendering advanced life support. The paper states that, based on experiences in the Vietnam war, seriously wounded soldiers had a better chance of survival in the battle field compared to traffic accident victims with similar traumatic injuries because the military was capable of providing first aid and efficient transportation to emergency care facilities. The white paper states that death rates from battle injuries declined significantly from 8% in World War I to 2% in Vietnam. The publication of this white paper is credited for the broad evolution and rapid advance of the civilian EMS system after 1965. It eventually lead to the creation of the National Highway Transportation Safety Administration (NHTSA) National Standard Curricula (NSC) for emergency medical care in 1985. Training, and therefore required skill and knowledge for prehospital emergency care has continued to grow to reach the current standard."

Jan 1, 2016

By T. E. Bernard

"Refuge alternatives provide shelter to miners trapped underground during a disaster. Manufacturers must demonstrate that their refuge alternatives meet the U.S. Mine Safety and Health Administration (MSHA) requirements for oxygen supply, carbon dioxide removal, and management of heat from the occupants and mechanical/chemical systems. In this study, miner size and activity level were used to determine the metabolic heat rate, oxygen requirements and carbon dioxide generation that are representative of miners in a refuge situation. A convenience sample of 198 male miners was used for the distribution of current U.S. coal miners, and the composite 95th percentile height and weight were determined to be 193 cm (76 in.) and 133 kg (293 lb). The resting metabolic rate (RMR) was determined to be representative of activity level in a refuge alternative. The highest likely metabolic heat generation ranged from 113 to 134 W, depending on occupancy. The highest required oxygen supply and carbon dioxide removal were estimated to be 23 L (0.81 cu ft) of oxygen per hour per person and 20 L (0.71 cu ft) of carbon dioxide per hour per person, which means the margin of safety is 50 percent or more compared with the MSHA requirements. The information on metabolic heat generation can be used to assess refuge alternative thermal environments by testing or simulation. The required oxygen supply and carbon dioxide removal can be used to assess refuge alternative requirements. IntroductionRefuge alternatives provide shelter to miners trapped underground during a disaster. Manufacturers must demonstrate that their refuge alternatives meet U.S. Mine Safety and Health Administration (MSHA) requirements for the supply of oxygen (O2), removal of carbon dioxide (CO2) and management of heat resulting from the refuge alternatives’ occupants and mechanical/ chemical systems (MSHA, 2008)."

Jan 8, 2018

By Joseph E. Worthington

The Haile Gold Mine is one of the earliest gold discoveries in the eastern United States as well as one of the largest gold producers from that region. Gold was first discovered in the gravels of what is now known as Haile Gold Mine Branch in 1828 on lands then owned by Colonel Benjamin Hale (spelling not the same). Small placer workings were immediately started and it was not long before the surrounding residual soils and saprolites were also found to contain gold. Following the presence of gold in the residuum eventually lead to the discovery of the underlying gold-bearing lodes that actually outcropped in saprolite form where not masked by heavy vegetation or a thin veneer of Coastal Plain sand. The earliest attempts at production, immediately after discovery, utilized slave labor, washing gravel and residuum from leased tracts that were 15m (50 ft) square. Several at- tempts at small milling operations were made during the en- suing decades, but there was really no substantial or successful production until the 1880s.

Jan 1, 1990

By Robert M. Dreyer

In an industrialized nation, as the process of industrialization continues, the ratio of the value of production of non-metallic to metallic minerals normally increases gradually over a period of years. For example, in the United States, the ratio of the value of production of non-metals to metals now is approximately two-to-one. Those of us who come from California like to think that we inherit the mantle of the old hard-rock miner, but the truth of the matter is that, in California, of the total value of mineral production exclusive of fuels, 97% comes from non-metallic - industrial - minerals. Since the industrial minerals furnish a principal source of raw materials for the construction and chemical processing industries and since these two industries, in industrialized nations, generally increase directly with increasing population, the future of industrial mineral production both in the United States and throughout the world is indeed bright. Because of the importance of industrial mineral production, it often surprises those of us in the industrial mineral industry to find that industrial minerals are poorly understood by many of our colleagues in the mining industry. I think that one reason for this comprehension of the industrial mineral industry is that the education of both geologists and mining engineers still markedly emphasizes exploitation of the metals.

Jan 1, 1964

By J. J. Whittaker

Zaldivar is an open pit copper mine operation located in the Atacama Desert, in northern Chile. The plant produces Cu cathodes using 4 solvent extraction trains. The plant was started up May 1995. Since 1996 the plant has been using on-stream analyzers based on X-ray fluorescence and titration for process monitoring and control. Elemental analyzer alone is not able to provide the critical information for plant operation, it has to be complemented by on-stream titrators integrated to one sampling system. The analyzers are located in a prefabricated shelter. The analyzer system project, operation and performance issues are the main topic of this paper.

Jan 1, 1999

By Wallace W. Roepke, James C. Church

Introduction The history of mining machine development has been and is one of "bigger is better." While this may not always be true, it is currently the only approach open to machine designers without specific in situ cutting force information. The US Bureau of Mines has tried to modify this approach by supporting research, both in the laboratory and in field tests, that demonstrates that bigger is only better if the equipment is used correctly. In-house research has shown conclusively that deeper cutting reduces both dust and energy (Roepke, Lindroth, and Myren, 1976). This approach has been applied to a large Lee-Norse (LN456) continuous mining machine, which confirmed that slow deep cutting is beneficial and does not have a negative effect on production (Black and Rounds, 1977; Black et al., 1978). Specific in situ engineering data needed on coal cutting forces have not, however, been obtained by field tests with the LN456. Although the machine was instrumented to obtain data on one seam, it would be impractical to move the machine from mine to mine. It became apparent that a portable, easy to use device was needed for obtaining coal cutting forces in situ. Such a device could provide the engineering data needed by manufacturers, the site-specific cutting information needed by operators, and a field research tool to establish correlation with laboratory cutting tests. Design and Development According to a set of design guidelines established by the Bureau, the proposed device should be able to measure cutting forces directly with standard coal cutting bits, be transportable by two miners, and be self-contained. These guidelines were given to Ingersoll-Rand Research Inc. (IRRI). The contractor performed a rigorous review of 16 existing in situ portable tester designs, none of which could use a standard coal cutting bit to measure cutting force directly. These included a wide range of devices, all of which are discussed in the final report (Banerjee and Wisner, 1982). A series of design tradeoffs were established to satisfy initial design guidelines. The final deci¬sion was heavily weighted by an industry survey to determine the preferred concept from four man¬ufacturers and four operators. To assure ease of use with maximum information, all respondents wanted a linear cutting device that could work in seams down to 1 m (40 in.). The final design choice was between a linear and a rotary tester. The rotary tester would cut on an arc in the same manner as a rotary drum machine. It had several faults, however, that could not be resolved within the scope of this contract. The worst was related to arc cutting (Roepke, Lindroth, and Myren, 1976). Since any bit cutting on arc enters the coal at zero depth, goes to maximum depth only at the horizontal centerline, and exits at zero, maximum cutting force is obtained only at the instant the bit crosses the horizontal centerline. This centerline peak response would require that many replications be made with an arc-type cutter to get reliable force information. The arc-type tester would require extensive face preparation to establish the cutting path and extensive wasted clearing cuts to advance the device for additional layers of cuts. The linear tester could be mounted directly on any face or rib with only minor preparation. By its nature, a linear cut at a constant depth provides multiple peak force data over the length of the cut, thus reducing the required number of replications. The final review of these two major design concepts plus indus-

Jan 9, 1983

General. Data for the operating information given in the 15 tables in this chapter was obtained from voluminous questionnaires sent to over 100 operating companies in 1972. There has been a considera¬ble change since then in supply and power costs and many companies have had expansions or revisions since that time, so it must be realized that information is valid for the year 1972. This is true except for the three new properties of Sacaton, Metcalf, and Pinto Valley. These three operations started up since 1972 and all information for them was obtained in 1975. The questionnaires were sent to all major copper properties throughout the world as well as to many smaller units and to some mines that had complex or unusual conditions. Properties for which no information is given did not submit completed questionnaires. Several operations were unable to supply information requested be¬cause of company policy prohibiting it. However, data is presented for 75 different operating properties. The reference number of the left-hand column is the same for all 15 tables of data. If the company had several operating properties as of 1972, then subsequent properties use the same number but a different letter. For example, Anaconda's property at Butte, Mont., is IA, whereas Anaconda's (now Anamax) Twin Buttes property in Arizona is IB. Table I gives general data for the properties including exact loca¬tion, milling rate, pertinent assay values and recoveries, and a brief listing of ore and gangue minerals. It was not possible to obtain any more information from the properties in Chile except that given in Table I. Crushing. Crushing plant operating data is given in Table 2. This table gives data on the crushers, feeders, and screens used at each property. The data obtained on feeders was confusing, but, rather than completely eliminate this information, it was decided to report it as given. The confusion lies only in the description of where the feeder is located. Where a primary feeder is listed it was not possible to determine if this feeder was on primary feed or primary discharge. For secondary feeders it is known that the feeder listed usually is feeding the secondary crushers, but at some properties it is handling the secondary crusher discharge. The same reasoning holds true for tertiary feeders-they handle either tertiary feed or discharge. This table contains a voluminous amount of data. A column giving tons per day per square foot of screen area would have been very informative, but some properties included circulating load in the feed rate and others listed new feed only, so calculations were meaningless. Autogenous Mills. The use of autogenous or semiautogenous mills for any property must be given serious consideration in the future. Table 3 lists data for four properties that are fully autogenous and four more semiautogenous mills that use balls to supplement the chunks of ore. This method of grinding is especially applicable when the ore is of such a nature that crushing problems would be severe, such as when ore is wet or sticky or contains a lot of fines. Paradoxically, it should also be considered when ore is tough and hard and blocky because then the competency of the ore means it can serve well as the grinding media. Where the ore varies greatly in hardness, as at Cyprus Pima and Lornex, semiautogenous grinding has been especially beneficial with capital costs and operating costs both 10-20% lower than could be obtained with conventional crushing and grinding. Rod Mills. For a great many years the standard milling operation used two- or three-stage crushing followed by open circuit rod mills and then ball mills in closed circuit with classifiers of some kind. A surprising number of properties crush to the fine size of about 3/8 in. and then use ball mills only, in closed circuit, with no rod mills. It is difficult to say whether rod mill-ball mill circuits or ball mill only is the best. One can be certain of two things: (1) the argument will go on for a long time and (2) for some mills one method obviously was superior over the other. As ball mills of 14, 15 ft, and larger came into use, more properties like Duval Sierrita (16b ft diam by 19 ft long) and Bougainville (18 x 21 ft) were built with single-stage ball mill grinding. Although there are larger rod mills used, especially in the iron ore industry, the largest rod mills on copper ore are at Twin Buttes (14 x 18.5 ft) and at Gibraltar (13'A x 20 ft). Table 4 lists operating data for rod mills. If a company is listed in Table I

Jan 1, 1985

By A. C. Beaulieu

CITY OF EDMONTON Rate of Growth The City of Edmonton has been classed by some economists as being the most progressive City in Canada since the early 50's. The high rise buildings that are changing Edmonton's skyline, the rapidly expanding sewer and water network, and the proposed rapid transit system all indicate the requirements of an ever expanding City. Edmonton, the core of Canada's Energy Empire is the country's seventh largest metropolitan area. Bisected by the 22 mile long park of its magnificent river valley, it sits in the midst of the 50,000 sq miles of riches that have created it. Underneath the City, however, extending down for thousands of feet and encircling it for a radius of scores of miles, lies a fabulous store of petroleum wealth. By the early 1940's, based upon the needs of the farming areas, accompanied by some small exploitation of coal and timber, Edmonton grew to 100,000 people. Then in a short space of 20 years Edmonton shot up to be a city of 450,000. Figure 1 (Slide No. 1) indicates the population growth in the City of Edmonton and the Census Metropolitan Area from 1906 to 1971.

Jan 1, 1997

By Steven Hart, Graeme Hill

Boddington Gold Mine is situated 130km South-east of Perth, WA. The operation is managed by Worsley Alumina Pty Ltd. The Direct Leach plant was commissioned in mid 1986 at a rate of 3 million tonnes per annum. The plant was upgraded to 4.5 mtpa in 1988 and almost immediately to 6mtpa in 1989. Subsequent debottlenecking resulted in further increases in throughput up to 7mtpa by 1991. The SupergenelBasement Plant was constructed in 1990 and commissioned in early 1991. It was designed to process both copper-gold-silver anomalies within the laterite gold orebody and future hard rock ores. The Supergene plant incorporated control potential s­ulphidisation (CPS) and flotation ahead of CIL to remove a host of copper minerals. It operated in that configuration for 2 years at rates up to 350,000 tpa and average grades of 6.0 git gold and 1.2% copper. Gold recovery averaged 90.7% and copper recovery averaged 80.0% before the flotation circuit was decommissioned in March 1993. Quartz hosted high grade, free milling oxide and basement ores were discovered at the northern end of the orebody in 1990 and now comprise the major feed source to the basement plant. Oxide ores are currently milled at a rate of 550,000 tpa, while basement ores are milled at a 420,000 tpa rate. Metallurgical testwork during the feasibility stages indicated that up to 30% of gold in plant feed could be recovered by gravity separation methods. It was also highlighted that signifi­cant improvements in overall plant recovery and reduced operating costs would be achieved by maximising gravity gold recovery ahead of cyanidation. To achieve this, rec;iesign and upgrade of the gravity circuit was necessary.

Jan 1, 1995

By Roger Holmberg

INTRODUCTION The driving of drifts is a very important aspect of underground mining. It is not unusual for the percent¬age of rock broken during development in a mine using sublevel caving, for example, to be as much as 25% of the total. If one also considers the amount broken for transport, ventilation, and exploration drifts, one can easily understand that the planning and excavation of drifts play a major part in the total economics of the mine (Fig. 1). Increasing mechanization in mining demands larger tunnel areas for transport and mining equipment. With modern machines the hard work involved in using hand¬held pushers is gone, and a better environment is achieved. More rational methods could be used, but much of the experience the working man acquired by working close to the rock face (such as utilizing the natural weak planes in the rock when he placed the drill¬ing holes) has unfortunately been lost. By having sepa¬rate shifts for drilling, loading, and hauling, more atten¬tion has to be placed upon a well-designed drilling pattern. Some reduction in the number of holes required can be achieved with mechanized drilling because of the larger holes that can be produced. On the other hand, it is probably not possible to achieve the same precision as with pneumatic pushers, and it is difficult to utilize the larger holes because these cause more damage to the re¬maining rock. Recently, however, the precision has be¬come very good with the parallel booms and automatic devices for setting the lookout angle (Fig. 2). A larger arch of the drift roof requires a more carefully executed blasting procedure than before in order to prevent rock fall and to insure a sufficiently long stand-up time. In this chapter, empirical relationships that can be used to design an economic and optimal drift blasting design will be presented. The principles of the calcula¬tion method are based upon the earlier work of Langefors and Kihlström (1963) and Gustafsson (1973). COMPARISON OF EXPLOSIVES To provide for the use of various explosives it is necessary to have a basis of comparison. Several meth¬ods have been developed to characterize the strength of an explosive. Some examples are comparison of values given by (1) calculated explosion energies; (2) the ballistic mortar test; (3) the Trauzl lead block test; (4) the brisance test; (5) the weight strength concept; and (6) the underwater test. However, most of these meth¬ods should be used carefully when stating the breaking

Jan 1, 1982

By Ricky Chipka, Roshan Thapa, Richard Yovichin, Greg Rogoff, Patrick Niemuth

A 4.5-mile-long, 96-inch-diameter outfall pipeline was constructed between North Texas Municipal Water District’s proposed Sister Grove Regional Water Resource Recovery Facility and the Stiff Creek Discharge Structure. Two TBM-mined tunnel segments were excavated for the outfall pipeline through rock and mixed-face conditions within the Austin Chalk Formation. The TBM mining performance was tracked in detail and evaluated during mining. This paper discusses the TBM configuration, TBM utilization, advance rates, and challenges of both tunnel segments.

Jun 13, 2023

By P. K. Hopke, A. Hubbard, K. H. Leong, J. J. Stukel, K. Nourmohammadi

INTRODUCTION In order to understand the transport and deposition of radon daughters in mine atmospheres, it is necessary to know the variation in the attachment of the daughter atoms to particles as a function of particle size, composition, number density, relative humidity, temperature, and radon concentration, the free gaseous diffusion coefficients of the daughters, and the variation in the mass transfer of the activity, both free and attached to particles, to mine surfaces as a function of particle size distribution, surface roughness of the mine walls, and the flow conditions. If all of these parameters are known in a model system, it should be possible to understand the transport and fate of the airborne radioactivity in real mines under certain well-defined flow conditions. There have been a number of recent investigations of the attachment of radon decay products to particles 1-4, but there are still a number of unanswered questions regarding the process. However, it is clear that for most real mine atmospheres, the vast majority of the activity is attached to particles. The size distributions for the activity-bearing airborne particles have been studied 5,6, and it has been found that most of the activity resides on particles with diameters in the range of 0.05 µm to 0.3 µm with an average mass median diameter between 0.1 and 0.2 µm. The behavior of the unattached radon daughter species has also been recently studied[ 7], and many of the previous problems regarding the value of the diffusion coefficient for Po-218 have been resolved. A major problem in the understanding of the airborne transport of radioactivity in mines is the lack of detailed knowledge of mass tranfer to and fluid flow over rough walls under fully developed turbulent flow conditions. This paper will report the progress on a project that is designed to obtained that information. MATHEMATICAL MODEL DEVELOPMENT Deposition of particles on smooth surfaces in turbulent flow has been extensively studied. A comprehensive review of these results has been prepared by Sehmel 8. There has not been such a comprehensive study of particle deposition on rough walls under such flow conditions. In recent years, only a single model has been proposed to explain such deposition 9,10 and in both of these papers the flow structure in the rough walled pipe was not taken fully into account. As part of the work being conducted on this project, a more complete model was outlined in a previous report [11]. The basic theory will be reviewed to provide a context for the flow measurements to be reported. The flux of particle deposited on the walls of a pipe in a turbulent flow is derived from the one dimensional form of Fick's law as given by [N = Dpdpp/dr (1) where N is the flux of particles deposited per unit area per unit time, D is the total eddy diffusivity of the particles, p is the airborne concentration of particles, and pr is the distance measured from the center of the pipe. The rate of deposition is best expressed by a deposition velocity Vp = NIP pb (2) where P b is the mean particle concentration in the sulk flow. The shear radius, V/ut and the shear velocity, u , are used to calculate a nondimensional distance, and velocity, respectively, where v is the kinematic viscosity of the fluid. The nondimensional form of equation 1 is given by Vd = DP dpp(3) V dr+ where Pp = Pp/ Ppb(4) By integrating equation 3 from the rough wall stopping distance, S , to the center of the pipe, the deposition velocity can be obtained. In order to make this calculation, it is necessary to have accurate descriptions for the particle eddy diffusivity, stopping distance, and shear velocity in order to insure that the influence of the flow structure has been properly accounted for. The shear velocity can be determined experimentally from the shear stress evaluated at the wall, Tw, and the fluid density, ut =VT w/p = ub V f/2 (5) where ub is the mean bulk axial velocity. The wall shear stress for a given pressure drop, dP/dL, and hydraulic diameter, Dh, is]

Jan 1, 1981

By L. H. E. Weyher

The Bethlehem Mines Corp. is a captive mining subsidiary of Bethlehem Steel Corp. Its coal production is tied to the demands for and production of steel, a situation that has to balance between the following two extreme strategies: (1) either the mines are designed and operated for some optimum production level and the demand fluctuations within the steel corporation are met by supplementary corporate coal sales and purchases, or (2) the mines absorb all the demand fluctuations, regardless of the effects on mine costs. A rational definition of the best strategy between these extremes is, of course, needed but is very difficult to formulate in view of the complexity of the many factors that would have to be included in the analysis. Consequently, rather than setting profit targets for each given mining division and/or individual mine, the practice evolved to set production targets for producing specified tonnages of coking coal for shipments to the ovens located at the corporation's iron and steelmaking plants. This in turn leaves the mines with the dilemma of having to explain demand-related cost-per-ton fluctuations while essentially depriving them of the ability to control profit through tonnage optimization. In the early 1970's, the coal prices rose significantly, while other factors led to sagging productivity in underground coal mining. Thus, the role of Bethlehem's coal operations in the corporate profit picture gained in relative importance. This, in turn, sparked a closer attention to the coal mining cost structure and factors controlling productivity. At the same time, the drastic increase in the capital needed to open a new mine also increased the amount of planning required to minimize risks. All of this called for more well-defined, quantitative operating data. Bethlehem therefore launched, in the early 1970's, a number of innovative steps that, as the future may show, represented pioneering contributions to the growing sophistication of North American underground coal mine planning and management control practices. This case study describes the changes that led to the current implementation of a novel production reporting system at Bethlehem's coal mines. These changes took place in two distinctly different phases. The current system that resulted from the Phase II developments utilizes, of course, experience gained earlier and is expected to be amplified to incorporate more advanced concepts and technology as future experience dictates.

Jan 1, 1979

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

"The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud.INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1]."

Jan 1, 2018

By M. Li, W. R. Reed, J. P. Rider, J. D. Potts

The National Institute for Occupational Safety and Health (NIOSH) has recently developed a series of validated models utilizing computational fluid dynamics (CFD) to study the effects of air-blocking shelves on airflows and respirable dust distribution associated with medium-sized surface blasthole drill shrouds as part of a dry dust collector system. Using validated CFD models, three different air-blocking shelves were included in this study: a 15.2-cm (6-in.) wide shelf; a 7.6-cm (3-in.) wide shelf; and a 7.6-cm (3-in.) wide shelf at four different shelf heights. In addition, the dust-collector-to-bailing airflow ratios of 1.75:1, 1.5:1, 1.25:1, and 1:1 were evaluated for the 15.2-cm (6-in.) wide air-blocking shelf. This paper describes the methodology used to develop the CFD models. The effects of air-blocking shelves and dust-collector-to-bailing airflow ratios were identified by the study, and problem regions were revealed under certain conditions. The simulation results could be used for future development of dust control methods, particularly to address respirable silica dust, for a surface mine blasthole drill shroud. INTRODUCTION Surface blasthole drilling can generate considerable amounts of respirable silica dust. These high dust concentrations can be exacerbated by the high silica content as drilling cuts through silica-bearing materials such as sandstone and shale, causing overexposures for miners to respirable silica dust. These overexposures can lead to silicosis, an occupational lung disease that has no cure and is often fatal. A review of the Mine Safety and Health Administration (MSHA) respirable silica dust sample database from 2010 to 2016 for metal/nonmetal mining shows the following overexposure rates for occupations related to blasthole drilling: • 14.1% of rotary air drillers were overexposed to respirable silica dust (22 out of 156 samples) • 5.8% of rotary drillers were overexposed to respirable silica dust (9 out of 155 samples) • 7.3% of drill helpers were overexposed to respirable silica dust (3 out of 41 samples) • 7.5% of blaster/shotfirer were overexposed to respirable silica dust (3 out of 40 samples) Of significant interest is an x-ray surveillance study recently completed under the NIOSH Enhanced Coal Workers’ Health Surveillance Program (ECWHSP). This program offered free chest radiographs to surface coal miners in 16 states. Pneumoconiosis was found among 2.0% of the 2,328 screened miners with at least 1 year of mining tenure. Twelve miners had radiographic changes consistent with progressive massive fibrosis, and 9 of these 12 miners reported no underground mining tenure. All but 1 of the 9 miners worked at occupations that were in the vicinity of blasthole drilling, either as a driller, blaster, or blast crew member [1].

Jan 1, 2019

By George M. Meisel

This paper is devoted to a discussion of the factors governing design of concentrators for large grinding circuits. Emphasis is placed on the manner in which use of ever larger grinding mills has affected the size of grinding bays, service cranes, gear reducers and electric motors, and the selection of ancillary equipment for cushioning of the starting load of large mills. Information is presented on the effect of using large grinding mills on the capital and operating costs of concentrators, e. g. : (Fig. 1) ? Benefits in lower cost of mills, drives and materials handling equipment, plant space, power, grinding media, operating labor and maintenance, and in ease of instrumentation and automation. ? Penalties in distribution of materials from the grinding mills to multiple lines of equipment which have not been developed to commensurately large sizes, and higher cost of ancillary equipment.

Jan 1, 1968