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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 G. A. Swanquist

INTRODUCTION This article describes the yellow cake drying, packaging and dust collection circuit of the United Nuclear Church Rock Mill. It also describes the layout and floor plan of this section as well as the operation and performance of the various equipment. In order to place the yellow cake section in perspective of the mill as a whole, the mill circuit is described briefly. CHURCH ROCK MILL GENERAL DESCRIPTION The Church Rock Mill contains six operating sections shown in Figure 1 in block diagram form. The grinding section consists of a semi-autogenous mill in closed circuit with a hydrocyclone. The slurry is then routed to the leaching tanks. Sulfuric acid and sodium chlorate is added to dissolve the uranium. The pulp is transferred from the leach tanks to the thickeners where the uranium bearing liquid is separated from the barren solids. The uranium bearing solution is then filtered and pumped to the solvent extraction section. [ ] In the solvent extraction section, the uranium bearing aqueous solution is brought into contact with an organic solution composed of kerosene, amine and isodecanol. The amine extracts the uranium from the aqueous solution. This organic solution is then routed to the stripping section, where the uranium is removed from the organic solution with an ammonium sulfate aqueous solution. The pregnant stripping solution now concentrated to 30 gpl U308 is routed to the precipitation tanks where ammonia is added to precipitate the uranium. The yellow cake is then washed in a counter current circuit of two thickeners and one centrifuge to remove impurities, mainly sulfates in order to meet specifications. The centrifuge centrate is routed to the second wash thickener for forward wash. The centrifuge cake at approximately 50% solids consistency is transferred to the yellow cake dryer via a screw conveyor. YELLOW CAKE DRYING, PACKAGING AND DUST COLLECTION Figure 2 shows the flowsheet of the integrated circuit of drying, packaging and dust collection. This section of course centers around the yellow cake dryer as the main equipment. The other main units are the dust collectors for the dryer and packaging respectively, a yellow cake storage bin and the yellow cake packaging facilities. Physically, all of this equipment occupies four floors, beginning at the top floor: 1. Both dust collector fans, the centrifuge and the screw conveyor to transport the yellow cake to the dryer. 2. The dryer, fire box and both dust collectors occupy the third floor. Figure 3 shows the floor plan of the third and fourth floors. 3. The yellow cake storage bin and dryer dust collector sump are located on the second floor. 4. The ground floor houses the loading facilities consisting of two load out ports from the bin to barrels, a roller conveyor loop, drum weighing scale, and change room facilities. Figure 4 shows the ground floor plan. Figure 5 is an elevation view of this section showing the overall equipment arrangement.

Jan 1, 1981

By Michael T. Ballard

The key to productive computer usage is system software integration with full compatibility among individual components. The majority of software developers in our industry are still in the process of perfecting individual components. As a result, exchange and validation of data between disciplines becomes an arduous process. We at Morrison Knudsen Engineers, Inc. (MKE) have integrated an entire system, allowing exploration, geology, modeling, mine planning, and operation modules to work as a single, complete system with direct exchange and validation of data. The importance of integration to both developers and users of today's mining soft- ware is becoming increasingly apparent as we, an industry, move toward growth and enhancement of the technology of ore reserve acquisition, evaluation, and management. It is also important to understand the present state of integration and where it is leading US. Not long ago, another industry passed through the same stages we are now setting up for the mining industry. The stereo industry began with a few individual components, the record player, tape recorder, and radio. People were able to play records, listen to the radio and record conversations and music only through microphones on their tape recorders. Soon, one enterprising individual discovered he could tape sound directly from the radio or record player by capturing audio output from the speaker wires. Interfaces and patch cords were developed, providing a means to transfer audio information between components with not only little loss of sound quality, but also the elimination of back- ground noise. The industry took another step forward with the introduction of "integrated" stereo, where the components were packaged in one cabinet with shared resources. This reduced overall cost which paved the way for better quality. Today's stereos are now so well integrated that many of the components share data and resources from the same manufacturer, designed to work together for almost perfect data transfer between any of the components. The result is better sound, convenience, and reduced cost. This is the ultimate goal for integration. Any system development passes through these same steps on the road to total interaction. Initially, the components are designed, tested, and marketed individually, often by separate companies. As the users become familiar with the components, the need arises for an automatic interface between the components. This is necessary to cut down on error during data transfer and to eliminate the redundant work required to enter results from one component into the next component in the system. As a result, interfaces must then be developed to aid in this data transfer. After data transfer is accomplished, the next step is resource sharing. Resource sharing is careful design and engineering by either cooperative, individual manufacturers or one firm taking on the task to produce a single package product. Design and resource sharing leads to lower overall cost, while still providing component capability with options to grow as the user grows. The mining software industry has been working to develop some very high quality component software and has indeed accomplished this task in every aspect of the mining industry. However, realizing the importance of integration, some companies with limited individual resources have started working together to interface their individual components. Other firms have undertaken the project of developing complete, integrated systems.

Jan 1, 1987

By C. E. Capes

INTRODUCTION In the coal industry, ever greater amounts of fine coal must be processed due to the need to crush to finer sizes for impurity liberation in low cost physical cleaning processes. Fine coal processing is a major problem for the industry, both through the cost of disposal of reject fines and through the mining cost of lost coal values. Of the available disposal methods, impoundment has always been the most popular. In many cases, when the coal fines were of sufficient quality, froth flotation and expensive filtration methods were used to recover the coal in these waste streams. The coal fines are not always of a marketable quality, however, and flotation, due to poor yields and difficult dewatering, has always been an expensive and not necessarily profitable process. The oil agglomeration method is increasingly important as a technology for the treatment of fine coal-in-water suspensions. These fines can be preferentially wetted and agglomerated by many different oils mixed with the suspensions. Ash-forming impurities remain in suspension and are rejected when the agglomerates are recovered, for example, by screening. Agglomeration-based methods provide perhaps the only practical means for upgrading extremely fine coal to useful products on a large scale. As with froth flotation, oil agglomeration relies on differences in the surface properties of coal and impurities to effect separation. Whereas flotation be- comes less effective with particles finer than 200 mesh (0.075 mm) or coarser than 50 mesh (0.3 mm), there appears to be no lower limit on particle size for oil agglomeration while particles % in. (3.175 mm) diameter or larger can also be treated, if desirable. In addition to this ability to clean a broader size range of coal particles, oil agglomeration can produce a dense, coarse granular product of acceptable strength and low water content without the need for vacuum or thermal drying. The preferential wetting of solid surfaces by immiscible liquids has been utilized, if not fully understood, for a long time. Indeed Gaudin,229 in tracing the ancestry of froth flotation, refers to a 15th century oil-pulp kneading procedure in which water-wetted particles were freed under water from an oiled mass. An understanding of the physical chemistry involved in such

Jan 1, 1991

By D. T. Iseley, David Bennett, Christopher Norris

The microtunneling test program is one of three main elements of the research project, "Trenchless Construction: Evaluation of Methods and Materials to Install and Rehabilitate Underground Utilities." The other elements of this project include an evaluation and demonstration of horizontal directional drilling technology and pipeline rehabilitation systems. The research is a Corps and Industry cost-shared project funded under the Construction Productivity Advancement Research (CPAR) program. The laboratory partner is the U.S. Army Engineer Waterways Experiment Station (WES), Geotechnical Laboratory. The industry partner is the Louisiana Tech University's Trenchless Technology Center. Industry participants are contributing over half the total cost of this research. The overall objective of the CPAR program is to improve productivity in the U.S. construction industry, thereby enhancing the competitiveness of the U.S. industry in domestic and overseas markets. The Corps and other federal and state agencies will benefit by realizing reduced costs, enhanced safety and better products and services as the products of this research program are developed, tested, and pushed into the market place. The microtunneling demonstrations and evaluations were performed during September and October 1992 at a specially constructed test facility built at WES in Vicksburg, MS. The test facility is 104 m (340-ft) long, 5 m (16-ft) wide, and 4 m (13-ft) deep. It was constructed, ironically, by first excavating a trench into which 6 different types of soils were placed and compacted in approximately 18.3 m (60-ft) long sections. The soil profiles include a lean silty clay (loess), a dry sand, highly plastic buckshot clay, wet sand, clay gravel and silt. The interfaces between soil profiles were sloped to simulate mixed face conditions as the machines exit one zone and enter another, and provided challenges to ground control and alignment and grade control during these transitions. Moisture and density tests were performed on all materials in each lift to assure uniformity. Horizontal inclinometers and 3 levels of settlement plates were installed at key locations in the test bed. The settlement plates and rods and the horizontal inclinometers, together with surface survey points, provided data to evaluate settlement and heave along the test section. Penetration resistance was measured in each section of the completed test bed. Strain gages were installed in 8 of the 0.6 m (24 in.) diam. Hobas pipe sections to measure strains in the pipe as it was being jacked into place. Steering jack loads, jacking thrust loads, and cutterhead torque were also measured during the drives. The results of these tests will be used together with other information to develop guidelines for conducting feasibility studies and selecting proper methods compatible with anticipated conditions.

Jan 1, 1993

By Geoff Beale

In terms of their hydrogeology, most open pit mines can be divided into three categories. There are mines that occur above the water table, mines that extend below the water table in high permeability rocks and mines that extend below the water table in low permeability, tight formations. Mines that occur above the water table do not experience major ground water seepage. It may some- times be necessary to construct in-pit trenches and sumps to remove water from precipitation. But dewatering or enhanced drainage measures are typically not required. Open pits that extend below the water table in high permeability rocks find it cost-effective to pump from dewatering wells. This is done to lower the water table below the working pit floors, creating dry or depressurized conditions in advance of mining. If the rocks in the mining area are homogenous, complete drainage of the entire mining area may be achievable by pumping from wells. In this case, it is not necessary to consider drain- age enhancement. However, there are many open pits and excava- [ ] tions that can lower the water table using conventional wells, but for which drainage enhancement is also an integral part of the dewatering system. The Sleeper Mine in Nevada is an example of this. Hydraulic layering at Sleeper has prevented downward drainage of water as the main water table is lowered. In addition, there are many mines that have areas of low permeability wall rocks that do not freely drain as the surrounding water table is lowered. Ground water seepage also creates difficulties in mines that occur below the water table but for which lowering of the water table using conventional wells is not feasible because of the overall low permeability of the rocks. The volumes of water involved may be small. However, it is necessary to deal with all of the water inside the pit as mining proceeds. It is not possible to dissipate pore pressures in the wall rocks in advance of excavation.

Jan 1, 1997

By Michael J. Martinson

Mine ventilation has been evolving as a separate field of study since the middle of the nineteenth century, and some practitioners might assert that mine ventilation has come of age as a discrete field in the last decade or two. Certainly the extensive literature now available on many facets of the subject, international congresses, the phenomenal upsurge of interest in mine ventilation in the United States of America since the passage of the 1969 Federal Coal Mine Health and Safety Act, and the existence of a virile mine ventilation society in South Africa with its own monthly journal all point in this general direction; on the other hand the lack of a rational design philosophy for ventilating mines suggests that mine ventilation may still be on the threshold of maturity. To generalise, many mines are still ventilated simply to comply with arbitrary statutory regulations, or with some ill- defined notion of reducing the risk of injury associated with the atmospheric environment in mines. Hazards in this category include oxygen deficiency, toxic gases and particulates, pathogenic gases and aerosols, explosive gases and dusts, and extreme climatic conditions. Although many mines are ventilated to abate the risk of a specific injury, there is inevitably a spectrum of latent and patent atmospheric hazards in every underground mine. The writer (Martinson, 1976) has previously drawn attention to the dearth of material on mine ventilation design in textbooks and technical journals, and to the overly simplistic approach sometimes adopted in the limited work published on the subject. The available material undoubtedly contains useful technical data and persuasive views, but seemingly could not be used to demonstrate the rational engineering approach to mine ventilation design advocated in the present paper. The range of papers presented at the 1975 and 1979 Congresses, and the diverse interests and affiliations of delegates, emphasize the multidisciplinary nature of mine ventilation. Easily recognisable topics include: Fluid mechanics Thermodynamics Heat transfer Flames and combustion Aerosol physics Radiation physics Geophysics and geomechanics Occupational diseases Epidemiology Industrial physiology Despite these heterogeneous connections there can be little doubt that mine ventilation is fundamentally an engineering field and not an off-shoot of some other branch of applied science. The distinction is important because mine ventilation has essentially developed as a form of ad hoc trouble- shooting, and in the process mining engineers have relied extensively on the assistance of experts drawn from other fields. Thus historically there has been limited scope for the application of rational design techniques, and personnel enlisted from scientific disciplines may not have had any previous experience in the intricacies of engineering design.

Jan 1, 1980

By E. J. Jacques

A typical problem in ventilation networks with multiple fan systems is the possibility of them interacting. Multiple fan systems are no exceptions in complex tunnels, and may also be met in mines where underground fans are installed. When aiming at controlling a multiple fan system in a ventilation network, e.g. with a view to maintaining an acceptable air quality level, yet at minimum power consumption, transient flow conditions frequently occur. Although in most applications these transients are short compared to the time needed to achieve the planned actions on the ventilation, one should, however, keep in mind that it is important to avoid any bouncing effect between different (adjacent or distant) fan systems. Therefore a simulation of the unsteady-state flows could allow us to learn more about the propagation of transients within the network, and their possible secondary effects Among others, one question that often is raised, is to find out in which sequence and at what time schedule each of the fans in a network should e.g. be started up? Another typical problem, into which an unsteady-state simulation could provide some insight is the following: how should the facilities of variable pitch angle and/or changeable RPM of several fans operating in parallel be handled and combined in order to obtain a flow rate that changes gradually from zero to its maximum value i.e. the sum of the individual throughput flows of all the fans. The literature is not very abundant on the topic of simulating transient flows in networks. There is no doubt that numerical models have been developed (Pollak, Christensen, 1987), most of the time for particular applications (Wala, Kim, 1985), (Tanaskovic et al., 1984), (Schultz, Sockel, 1991), etc. The aim of this paper is to present a model that also simulates one-dimensional compressible flows in unsteady-state conditions with, however, the specific intent to handle networks with more than one fan. PROBLEM STATEMENT The mathematical model chosen to describe the unsteady flows in a ventilation network is based on the dynamic and continuity equations applied to each of its airways. The following basic assumptions are made : - the flow in the airway is considered to be one- dimensional, - - the flow is assumed to be fully turbulent, and hence the ATKINSON equation for calculating the friction losses will be applied. Whether this equation remains valid for transient state flows remains an open question. However, this may reasonably be assumed since the interesting flow frequencies are low. A discussion may be found in (Kämpf, 1987) ; - the airways are assumed to have a constant cross- sectional area along their length. If obviously this would not be the case, it is always possible to subdivide the concerned airway into sections for which the previous assumption holds. Applying the basic equations of mass and energy conservation, the flow in each branch may be described by the Euler equation for Newtonian flows.

Jan 1, 1995

The US economy grew at a rate of 4% in 1994 and the gross domestic product rose to $6.74 trillion. With 261 million people - 605,000 in mining, including petroleum -the United States has l 10 million dwelling units and 200 million licensed motor vehicles. New private housing units authorized in 1994 rose 14% to 1.4 million units. New car production rose 11 % to 6.6 million units, while truck and bus production rose 30% to 5.2 million units. The inflation rate was less than 3% and the prime rate of interest averaged 7.15%. The producer price index (1982 = 100) for fuels and related products and power was down 4% to 78, while metals and metal products rose 5% to 125. Chemicals and allied products rose 3% to 132, and non-metallic mineral products rose 3% to 124. The merchandise trade deficit rose 25% to $166 billion. Disorder in the former Soviet Union led to substantial dumping of mineral materials. With vast mineral resources, the former Soviet Union land area totaled 22.4 million km2 (8.6 million sq miles). In comparison, the North Ameri¬can Free Trade Agreement (NAFTA) area (United States, Canada and Mexico) totals 21.4 million km2 (8.3 million sq miles).

Jan 1, 1995

By R. Svensson

Mill linings play an important role in all comminution processes with tumbling mills. The selection of the lining profile and the lining material has a great impact on overall grinding cost and mill throughput. The cost of the lining itself is minor in comparison with media cost, energy cost and the value of the material produced in the mill. The total scope must be considered when mill linings are selected. This paper presents case stories and explains the technology behind: •Different lining profiles in rubber and steel. •Orebed magnetic lining. •Steel-cap rubber lining. Introduction Grinding mill liner life has been a subject of extensive studies. Performance of mill liners in terms of wear life and how it relates to lining material and profile is abundant in the literature. The most economical liner material and design has been selected based on "cents per ton milled". The evaluation of the effect of liner design on power and grinding media consumption which is a far more significant part of grinding costs, has been less comprehensive (Fig. 1). Lining materials Steel. The most commonly used lining material is steel, calculated on a volume basis. Steel linings have a long tradition and a number of alloys is being developed. The main families are: •High chromium white iron. •Cr and Ni white iron (NiHard). •High chromium martensitic steel. •High carbon Cr-Mo perlitic steel. •High carbon Cr-Mor martensitic steel. •Manganese austenitic steel. Steel linings can be successfully used in most applications. Rubber. Rubber linings were introduced 25-30 years ago and are today commonly accepted to be the most economical linings for secondary and regrind applications. Rubber linings offer long service life in these applications and generally [ ] outlast steel in wet applications, where corrosion is a factor on the steel. The cost relation between rubber and steel has changed dramatically over the years. In the past rubber was more expensive than steel, but today, the cost of rubber is close to half the cost of steel, calculated on a volume basis (Fig. 2). The present cost situation has made rubber linings an economical alternative in several primary grinding applications as well. The wear rate of rubber in primary applications is higher than the wear rate of steel, thus giving a shorter wear life. This can be compensated for by making the rubber lining thicker. The cost advantage of the rubber allows for this, but care must be taken not to jeopardize the grinding performance. Rubber, as a lining material, offers some distinct advantages over steel lining, such as: •Ease of installation and removal. •Safer to handle, less risk of injuries. •Faster replacement. •Reduces noise.

Jan 1, 1996

By James D. Reilly

The time-worn adage "an ounce of prevention is worth a pound of cure" was never more applicable than when used in reference to coal mine safety. Many individuals have dedicated many hours to developing effective accident prevention programs, and much has been accomplished. Each production escalation, however, creates new hazards and safety problems which need special study and research. It is a never ending cyle. To study the progress of coal-mine safety is to raise memories of many individuals who have been dedicated to removing the hazards that have developed with increased productivity and deeper penetration into the earth for coal. Some vivid examples of this dedication are the works of the men who pioneered the use of the safety lamp and the use of rock dust in mining.

Jan 1, 1968

By David Williamson

In the post-World War II years, metals analysts and statisticians have tracked the demand for metals in line with changes in Organization of Economic Cooperation and Development (OECD) industrial production (IP) growth rates, Williamson said. This was because the correlation between IP growth and metal demand was reasonable. "However, the correlation was exaggerated," he said. During times of expanding IP growth, metal demand increased at a faster proportional rate than IP growth. This was probably due to industry restocking in anticipation of growing demand at the start of each economic cycle. Conversely, metal demand would slacken at a greater rate than would be suggested by negative IP growth as industry destocked to reduce costs. Nonetheless, he said, a reasonable correlation could be painted between IP growth over the longer term and metal demand. "Yet, although this correlation has stood well since the last world war, it has been based on the fact that the bulk of the world's IP growth stemmed from the industrialized countries of the western world, lately including Japan," Williamson said. The so-called "developing world" had been of little consequence. The Soviet Union and other communist bloc countries were deemed to be closed, self-contained entities that, for each metal, would either be net exporters or importers, he said. These quantities would be used by analysts to calculate the overall supply/demand equation for the western world only and to deduce whether a commodity was in shortage or surplus. It is now believed that this method of projecting total world supply and demand for raw materials is outdated, Williamson said. It does not take into account the worldwide political changes that have occurred since the collapse of communism and the revitalization of China. "It is no longer appropriate to assess global balances for raw materials on the basis of three economic blocs," he said, including the western world, developing world and communist bloc. Rather, global balances need to be projected on a global basis. In this regard, since the collapse of communism and the removal of many other totalitarian regimes, capitalist forces are attempting to satisfy the expectations of the great mass of the world's poorer people, Williamson said. Such people, usually located in Asia, Africa and Latin America, have aspired to the same standard of living as people living in western Europe, North America and Japan. "Now, with free market forces at work, their aspirations will start to be met." Consequently, the demand for metals and other raw materials will now rise at a much faster rate. The production of raw materials has to keep pace with rising IP in the developed world. It will also have to keep pace with the demands associated with the rising expectations of increasing populations in the developing world, as well as from within the old communist bloc countries.

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 Graham J. Morgan

The following material is excerpted from a presentation Mr. Morgan gave recently to representatives of the financial community. Introduction 1982 was a period of challenge for United States Gypsum Company because of reduced market opportunity. Housing starts for the first 10 months averaged a seasonally adjusted annual rate of 1 million units, 9% below the comparable 1981 period when housing starts were the lowest in 35 years. Industry gypsum board shipments declined 9% from the already low levels of the same period in 1981. US Gypsum met the challenge with stringent cost reduction programs and aggressive action in the marketplace. In late 1981, anticipating market deterioration, the company's management adopted strategies to focus on two shortterm objectives: • Adjust operations to perform profitably in view of limited opportunities through more efficient operations and asset management. • Continue to strengthen U.S.G.'s resources and ability to take advantage of important market opportunities as the economy improves through selective expansion of production capacity in key growth markets. Concentration on these two objectives resulted in a nine month performance in 1982 that, while certainly not satisfactory by company standards, compared very favorably with that of principal competitors in terms of net earnings, margin on sales, and return on average stockholders' equity. To counteract limited market opportunities, U.S.G. focused on productivity improvement, cost reduction, expense control, and maximization of market penetration. Energy Use Improving performance encompassed all segments of operations. For example, efficient use of energy was a significant factor despite lower operating rates at most plants. Overall, in the third quarter of 1982 versus 1972, the company used 27% less energy per unit of production, a savings of $15 million in that quarter alone. These continuing improvements helped to partially offset constantly rising energy costs. A number of operations achieved superior results in the first nine months of the year compared with the same period in 1981. For example, gypsum board plants in Jacksonville, FL, and Sweetwater, TX, recorded 8% and 7% reductions, respectively, in energy use per unit of production. Paper mills in Oakfield, NY, Galena Park, TX, and Kansas City, MO, each reduced energy use more than 12%, while the Greenville, MS, ceiling products facility achieved a 16%o reduction. Safety Safety is another important element that received increased attention. U.S.G.'s safety performance over the past five years has consistently improved. In 1981 it was five times better than that of the gypsum industry; nearly 10 times better than all industry as measured by the National Safety Council; and almost 20 times better than all US manufacturing as recorded by the Bureau of Labor Statistics. In addition to human concerns, this excellent record results in significant dollar savings that directly affect company results. In 1981, U.S.G.'s pre-tax savings in claims, premiums, and related expenses were about $20 million compared to the industries in which it competes. For the first half of 1982 performance improved further, with a cumulative lost workday incidence rate per 200,000 man-hours worked of only 0.27, an all-time record. Expense Control Careful control of inventories and receivables also received special attention. As of Sept. 30, 1982, inventories were 29%, or $44 million, below the level of Sept. 30,

Jan 2, 1983

By Eduardo Poltronieri Trés, Anderson Denadai Thomazini, Rodrigo Boyer Fernandes, Francisco de Assis Dutra Macedo, Gilvandro Bueno, Maycon Athayde, Rossano Augusto Pilon Nunes

The earliest industrial application of high pressure grinding rolls (HPGR) at comminution was in 1984 in the cement industry. Since then, the equipment has been widely applied in mining activities. Despite the rapid spread across the industry, several challenges are still present in the equipment’s application, especially considering the complex ore breakage behaviour reported for this process. The pellet feed’s HPGR comminution was recognized as fundamental to efficiently increase the particle surface area (e.g., blaine specific area, BSA) at a lower energy and water consumption level (Chapman et al. in J Southern Afr Instit Mining Metall 113:407–413, 2013). The present study considers a novel milling process applying only HPGR as a re-regrind stage after concentration. The amount of successive ground product recirculation into the machine to achieve the required particle size for the pelletizing process was investigated. The study compared the ball wet milling process with the innovative processing technology for hematite–goethite ores. Green pellet balling and induration processes were simulated in bench and pilot scale. The milling results showed a considerably steep increase in BSA after each recirculation step into the HPGR, although compared with the ball-milled product the size fraction <45 μm% increased a modest five percentage points after three recirculation cycles, remaining steady until the last recirculation step. As a result, the green pellet generated presented higher concentration of pellets between 10 and 16 mm (%10-16 mm), with a higher drop number than typically observed when produced with ball-milled feedstock. Consequently, on the induration process, a higher pellet deformation was observed during the pot grate loading. Despite that, the fired pellets still presented an outstanding performance in cold crush strength (CSS) considered suitable to an industrial application.

By C. V. Trang, V. T. Le, K. Enns, J. J. Byerley

The precipitation characteristics of copper and nickel cyanides from solutions continuing copper zinc and nickel cyano complexes are examined. Equilibrium precipitation calculations are presented as well as comparative experimental data. The practical significance of the precipitation data is discussed in the co·ntext of an initial step m an overall treatment strategy for mixed cyanide effluents.

Jan 1, 1998

By Tom F. Deems

The term ball clay is applied to plastic sedimentary clays that have a high content of kaolinite and that possess white firing properties and high green strength (Mitchell and Stentiford, 1973). However, the name ball clay is not derived from any specific property of the clay. Instead, it comes from the original method of production in southwestern England (Scott, 1929). From open pits, the clay was cut into cubes with hand tools. The sides of the cubes were about 220 to 250 mm (9 to 10 in.) and weighed 14 to 16 kg (30 to 35 lb.). The eventual rounding of the corners with handling before sale gave rise to the term ball clay. Ball clays provide the workability and green strength required by the whiteware section of the world&apos;s ceramic industry. The main applications are in the sanitaryware, tile and tableware sectors. Each, though, requires particular technical properties in the ball clay component. The clay has to be as white firing as possible for all applications. However, for sanitary-ware, the rheological properties are most important. For tableware, fired whiteness and freedom from specking are the main criteria. Tiles need a reasonable green strength but the clay also needs to be cost-effective. The technical properties of the individual clays in any deposit depend on the clay mineral assemblage. This, in turn, depends on the geological formation of the deposits. Hence, it is the latter that largely determines the market sector for which the clay blends from a particular deposit are suited. Formation of ball clays Ball clays are formed from three main components - kaolinite, micaceous minerals and quartz - while carbonaceous material is also often present. The crystallinity of the kaolinite and mica varies. There are also occasional minor occurrences of feldspar, chlorite, montmorillonite andmixed. layer minerals. Marcasite, siderite and anatase also occur in small amounts. The properties of the clay depend on its mineralogical composition and its particle size distribution, particularly in the lowest size ranges. The original source rocks and the environment of deposition of the clay are the controlling factors. The importance of weathering processes in the formation of ball clays has long been recognized (Bristow, 1968). The diversity of source rocks produces considerable variation in the clay minerals formed. Therefore, each deposit is unique, with time-related variation within it. The greatest development of a weathering mantle takes place in a tropical climatic environment. Plant remains confirm that most ball clay deposits were formed under hot and wet conditions. Sedimentation of the clays had to be in fresh or low salinity water or clay minerals detrimental to whiteware manufacture would have been formed. Fresh water traps for the ball clay, sand and lignite sequences depended -on tectonic features, such as basin development along fault zones or lakes and lagoons on the edge of a large scale embayment. These factors for Devon ball clays were discussed by Vincent (1983). The combination of suitable source rocks, tropical climate, significant fresh water sedimentary environment

Jan 1, 1997

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984

By John W. Amos, John A. Christophersen

The Deister No. 7, "77" and "88" Coal Washing Tables have been used by the coal industry in the cleaning of finer sizes of coal (-1/2") since washing of coals became necessary. This paper will deal with the in-plant use of the CONCENCO "88" Deister Coal Washing Table, along with a brief description of the Deister Refuse Superscalper. Efficient in-plant operation of the Deister Coal Washing Table can be accomplished by the adjustment of six basic variables. These variables are: 1. Headmotion 2. Feed 3. Riffle Pattern 4. Dressing Water 5. End Elevation Adjustment 6. Side Tilt Adjustment VARIABLE NO. 1 HEADMOTION The reciprocating movement of the "88" Table is developed within the headmotion entirely by rotating eccentric weights. Four parallel shafts each carry pairs of eccentric weights (Figure 1-A). The shafts are geared together with gears of 2 to 1 and 1 to 1 ratios, so that the two shafts in vertical alignment which carry the small eccentric weights operate at twice the speed of the other two shafts. The other two shafts are also in vertical alignment with each other and carry large eccentric weights. As power is supplied to one high-speed shaft, once each revolution the forces produced by the large eccentric weights are reinforced and then opposed by the small eccentric weights in a horizontal plane. The direction of movement of the table decks is reversed much more sharply at the forward end than at the backboard end of the stroke, causing material to move toward the refuse end of the table. Because in the vertical plane all centrifugal forces oppose each other and cancel out, there is no vertical movement and virtually no impact to the support steel. The "88" Coal Washing Table is suspended by wire-rope cables from three overhead steel beams, and operates with a perfectly uniform and exact differential motion curve in every stroke (Figure 1-B). The decks, being attached to the headmotion, are reciprocated in a parallel line from the head- motion to the refuse corner at speeds varying from 260-300 strokes per minute. Both length of stroke and speed are set at the factory for the coal characteristic desired and should not be changed. A 3 HP motor is mounted on the top of the headmotion and uses 3 V-belts to transmit driving power to the headmotion. The headmotion must be kept running on a horizontal plane. Punch marks are provided on the side and end of the main frame for use in leveling the motion (Figure 1-C). In leveling the motion, the work must always be done with two cables. To level the long side of the headmotion both front or both rear cables are raised or lowered to bring motion back to level. To level side to side, both right side cables or left side cables are raised or lowered to bring motion back to level VARlABLE NO. 2 FEED When tabling different size coals, each deck of a "88" DIAGONAL-DECK Coal Washing Table will handle feeds as described in figure 2-A.

Jan 1, 1988

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