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By T. Stuczynski

In the past ten years about 3000 wastewater treatment plants have been built in Poland. The result of this development has been a massive increase in overall sludge production. According to data supplied by the Ministry of Environmental Protection, 400,000 dry tons of sludge are generated every year in Poland (Koblak-Kalinska 1996). Still it is difficult to estimate the amount of sludge, which is used in agriculture or for reclamation purposes. However, it can be assumed that this comes to no more than 10 percent of all sludge produced in Poland on a yearly basis. Thus it became obvious that an urgent solution to this problem was needed. Local authorities in Silesia realized in the early 90?s that a demonstration was necessary to encourage the proper utilization of sludge. This was the basis for establishing the Biosolids Subpart of the Silesia Project, which addressed the use of biosolids in the reclamation of mining and smelting dumps (Stuczynski et al. 1996). Metal waste sites themselves are known to contain more than 87 million tons of waste (Pistelok et al. 1995). Each year this amount increases by approximately 400,000 tons. We realized that a simple solution for stabilizing these sites would be to cover them with vegetation in order to reduce leaching of toxic elements, as well as to keep metallic fugitive dust from entering the environment. The reclamation of smelter waste sites, which was carried out within the framework of the Silesia Project was a joint effort of local government, industry and international research institutions/agencies, including the US Environmental Protection Agency (EPA), the Center for Research and Control of the Environment (OBIKS), Virginia Polytechnic Institute and IUNG. One of the goals was to evaluate the possibility of adapting the rules of sludge management currently used in the United States to conditions commonly found here in industrial regions of Poland. However, the main objective of the Silesia Project was the development of guidelines concerning all aspects of sludge use for the reclamation of degraded lands and waste sites. Biological aspects related to reclamation of metal waste with the use of biosolids will be discussed in this paper.

Jan 1, 2002

By Andrew S. Lain

Before you begin mining aggregates in Oklahoma, you must first look at the market and ask these questions: (1) Is there a need for additional aggregate production? (2) Can we produce the aggregate efficiently enough to compete? (3) Is the deposit close enough to the market? If you can answer "yes" to each of these three questions, then you are ready to start the process for mining aggregates in Oklahoma. The first consideration to be made is whether the aggregates produced will need to meet state specifications. If you do not want to meet state durability standards, there are numerous formations you can mine and produce aggregates. We will deal only with the formations that will meet state durability standards. These formations are located in the southern half of Oklahoma and in the Tulsa area. Some of the formations that contain limestone, dolomite, and sandstone that will meet state durability tests are West Springs Creek, Kindblade, McKenzie Hill, Butterly-Royer, and Rush Springs Formations. Granite and rhyolite also are found in Oklahoma and can be used as construction aggregates. Once an area has been identified as having durable material, and drilling has verified its properties, a permit must be secured from the Oklahoma Department of Mines, Oklahoma Department of Environmental Quality, and any other applicable agency before construction of a plant occurs. There are many different brands of crushers available, but all work on one of two crushing principles-either compression breakage or impact breakage. Each type has its application in different industrial-mineral types and the desired end product. A typical small crushing plant would start with a jaw crusher for primary reduction, and then employ one standard cone and one shorthead cone for secondary reduction. One grizzly would be in the circuit with one finished-product screen and three to four finished-product stackers. Natural sands and gravels are found in rivers and old river channels. These too need to be tested to see if spec product can be produced. All of our sand operations have very little gravel; thus, we do not process the gravel. The sand is processed through classifying tanks for gradation.

Jan 1, 1999

By Mark W. DeLawyer

The base of the stone quarry still stands today, as imposing as any Mavan temple or Egyptian pyramid. For 80 years it has been a reminder to those who must sweat to earn a wage. There are no plaques to indicate its impressive history. Rather, it stands as a towering monument to the masons and stone cutters who built it-a testimony in stone to that most demanding and back-breaking of industries. Take your imagination back to the limestone quarries at Split Rock, near Syracuse, NY. The sound of an occasional steam drill and the mighty roar of a dynamite explosion herald progress. The ring and crash of a hundred sledge hammers are music in the calloused hands of the quarry men. It is a symphony of man against nature. Hundreds of sweat-soaked men break up the blue Onondaga limestone for the "sodi ash," the popular name for the quarry owner, the Solvay Process Co. Nearby, expert stone masons fit mammoth 454-kg (1,000-1b) blocks of cut stone into place, building the foundation for what may have been, back in 1902, the world's largest stone crusher. Consider for a moment, a modern stone quarry. The business is now essentially one of machine against nature. The shovels, drills, conveyors, and locomotives help make the quarry of today a profitable venture. Now consider the quarry of the early 20th Century. Anyway you look at it, a tougher, more demanding life would be difficult to imagine-12 hours a day, all kinds of weather, breaking up tons of rock that had been drilled, blasted, or pounded from the bowels of an uncooperative earth. Before proceeding, it might be prudent to briefly describe the history of the Split Rock quarries. Before the purchase of the quarry lands by the fledgling Solvay Process Co. in 1889, the quarries already had a reputation literally chiseled in stone. Most of the municipal buildings in Syracuse were made of it, as well as over half of the Erie Canal, including the engineering marvel of the time, the Montezuma Aqueduct. This was where the Canal passed not only through, but in some

Jan 11, 1982

By Albert W. Evans

Introduction Engineering Health and Safety is concerned with the understanding of the complex interplay between the risk of injury with technological ,environmental, individual, organisational and legal factors. With this understanding, the design and decision making processes can be balanced with the other objectives of industry and employment to arrive at solutions which place appropriate controls on technology and other aspects of human activities to prevent disease and injury to personnel in the coal mining industry and the design of safe production machinery and ancillary equipment. Advances in the detection and diagnosis of causes of diseases and the need for engineering health and safety in the coal mining industry has increased the need for appropriate education and training. Hazards arise in many forms; exposure to toxic chemicals, escape of radio - active materials, explosion of pressure vessels, contact with moving machinery, exposure to dust, environmental dangers such as excessive noise, high temperatures and manual handling etc. As technology advances, these dangers become more complex and less clearly defined and less easily perceived. Consequently, the ways of controlling the dangers also become more complex and wide ranging. The need to educate and train people in coal mining applications of engineering health and safety has been enhanced over the last five years and courses should be established to fulfil the need for competent people to deal with the problems which modern coal mining technology creates together with the legislation that governs engineering health and safety and to carry out research and development work in the coal mining industry. Industrial Involvement The engineering health and safety experts must be prepared to take a more active part in monitoring conditions that exist in the mining industry, and not design from information obtained from the mechanical, electrical and hydraulic expertise and off-the-shelf components from other industries. Surveying and control of the operating environment must be essentially the concern of the engineering health and safety experts and, whilst it is not possible to create a perfect environment for each machine, every effort must be made to minimise failures and hence plan for designs which will maximise

Jan 1, 1986

By Paulo Roberto Pereira

This paper is motivated by the belief that increasing drilling and blasting costs will not necessarily bring about better digging or crushing productivity. As a matter of fact it can be counter productive as the ensuing discussion will show. In the two case studies presented herein when powder factor increased, productivity at a rope shovel and at a primary crusher decreased. Experience has shown that equipment (shovel and primary crusher) productivity increases as the powder factor increases and concurrently operating costs decrease. This is illustrated by Figure 1where the graph shows (assuming drilling costs constant) that for increases in powder factor initially operating costs might fall till a certain point and then they start to go up again. By the same token we need to understand what is the behavior of the operation’s productivity. It seems intuitive productivity goes up for increases in powder factor but when an operation reaches the point of minimum operating cost one could say that productivity reached its maximum. For continuing increases in powder factor can productivity still go up or does it go down? This paper presents two examples where the productivities of a surface coal mine and of a quarry are reduced for increases in powder factor. This indicates that these two operations could be operating to the right of their minimum operating costs and maximum productivity and therefore wasting precious drilling and blasting dollars and resources. Moreover, their productivity is following the red dotted line in Figure 1 and the question becomes can they follow the blue dotted line for increased drilling and blasting costs? The industry owes to itself to determine this minimum operating cost point and what can be done in case that the productivity goes down afterwards despite an increase in drilling and blasting costs. In other words, utilize your drilling and blasting resources correctly. In the two examples that follow single explosive loading strategies make powder factor a valid means of comparing the results of productivity as a function of powder factor. Also, external factors such as equipment operator influence are eliminated due to the long time span considered in the analysis

Jan 1, 2005

By John A. Cordes, John J. Schanz

INTRODUCTION The economic structures of industrialized countries require a continuous supply of reasonably priced and appropriately processed nonfuel minerals. As a result, national policy makers are sensitive to real or perceived threats to the effective operation of these commodity markets. Their most persistent concerns can be subsumed into the related political economy issues of supply availability and marketplace stability. These concerns are not new. For over a century the "global struggle" for mineral resources has influenced national policies and international relations (Eckes, 1979). During the past two decades the complexity of these issues, if not their place on national political agendas, has increased. Dramatic changes in the structure of mineral industries and in the nature of international economic relations have redefined policy needs and response options. This paper focuses on how selected OECD countries have responded to nonfuel mineral industry issues over the past twenty-five years. Particular emphasis is placed on the impact these policy decisions have had on the size and location of productive capacity and its utilization rates. The results, perhaps, are not surprising for this group of market- oriented economics. While individual cases of significant impact are easy to identify at various times in most countries, the overall importance of policy decisions has not been great. This is especially true for the past decade when excess capacity reached unusual highs of magnitude and duration. Macroeconomic forces and market structure changes played much more important roles. PERSPECTIVES ON THE ISSUES OECD nations dominate the international nonfuel minerals marketplace as producers, consumers and traders. In 1986, these twenty-four countries accounted for 75 percent of the Western world's export trade value in minerals, ores and metals. They imported 88 percent of the total value. During 1966, the respective percentages were 70 and 94. Japan was the world's largest importer in 1986 followed by the US, FRG, UK and France. The same five countries led the 1966 list except that, then, Japan was ranked fourth. Table 1 ranks the fifteen largest Western exporting nations in the two specified years. It also indicates the growth in current export dollars earned in 1986 compared to 1966.

Jan 1, 1989

The day after I reached New York, on January 30th, John Ryan of Anaconda told me where they wanted me to go. Late in 1916 the Russian Czar, desperate for money and supplies with which to carry on his unhappy share In the first World War, sent emissaries to the United States to see if our great mining companies might be interested in trying to find mines and to bring them to production in the millions of acres of the Imperial Estates in Russia and Siberia. It looked like a long chance as no one knew whether or where there might be important ore bodies. Even if they found them it might be a hopeless task to equip them under wartime conditions But the American copper companies had plenty of money and they were used to taking risks. They made a supersecret agreement with the Czar's deputies to send an exploring expedition to Russia at once I never learned the names of the other companies but Anaconda was to man- age the venture. Horace Winchell, then the leading mine geologist In the United States, was to be in charge of operations. Charles Janin, an eminent placer and dredge expert, was to study chances for producing gold and platinum. Though at 33 years of age I was much younger than the others, I was chosen to hunt for copper and other valuable metals. Charles Boynton, a business man who had lived in Russia and had married an attractive Russian woman, was to join the party and carry on any necessary financial or diplomatic affairs. He had gone on ahead to arrange for our reception Finally a graduate student from Ann Arbor named Steinberg was to go along as Interpreter. We soon found out that if his Russian was no better than his English his interpretlng couldn't be trusted too much I knew nothing about the millions of square miles In Russia and Siberia, and even less about what mineral deposits might be found, but I was delighted at the prospect. I was treading on air for the few

Jan 1, 1976

By Marini David, Pigorini Andrea, Fuoco Stefano, Zurlo Raffaele

"The excavation of a tunnel in highly weathered and strongly fractured rock masses due to tectonic actions is still a challenge. In order to maintain reasonable productivity, excavation through a wide zone of highly tectonized rock mass requires a detailed, integrated study of the construction solutions and a monitoring system to minimize the process time and to optimize the adopted consolidation measures in accordance with a predefined matrix of possible support systems. This paper summarizes some interesting data acquired during the excavation of the Brenner Railway Base Tunnel system between Italy and Austria. The Brenner Base Tunnel (BBT) runs through the fault line between the European and the African tectonic plate. This big fault zone is called “Periadriatic Line.” In particular, data acquired in the exploratory-service tunnel have been analyzed. Furthermore, the base concepts for defining the consolidation and support solutions adopted for the excavation of the exploratory tunnel are described. INTRODUCTION The Brenner Railway Base Tunnel (BBT) is a system of underground structures and tunnels that is designed to connect Italy and Austria using new railway standards. It is based on high speed (HS) concepts and applies the European Standards for Interoperability (TSI). The railway includes twin tunnels for a total length of about 55.0 Km without considering the connection tunnels to the existing underground railway in Innsbruck. Including the latter, the total length of the tunnel system reaches 64 Km, and is among the longest tunnels in the world. The Italian section passes through the so-called ""Periadriatic Line,"" which is an important tectonic structure generated by the collision between the European and the African tectonic plates. Despite the extensive investigations carried out (boreholes down to 700 meter depth), significant geological and mechanical uncertainties inside the Periadriatic formations remained unsolved. In order to elevate the level of confidence and observe the behavior of the rock mass at a representative scale, the BBT Authority decided to excavate in advance the service tunnel, which has taken the function of an exploratory tunnel. The data and observations collected during its excavation would provide important insights for the design of the main tunnels."

Jan 1, 2016

Up to the middle of 19 14, when I had reached what seemed like the advanced age of 30, I had enjoyed far more than my share of success and a lot of minor adventure. But the exciting experiences in distant lands that Soldiers of Fortune had made me long for had not yet materialized. And I had not accomplished the two goals that Mrs. Hughes' geography class at Belmont had set up for me. They were to see In its tropical home the enormous yellow bill of the toucan bird and to ride on a straw "bolsa" on Lake Titicaca. These aims were gratified In the next five years that were the epitome of "romance" in its German connotation of facts seen through the rosy glasses of imagination. Dr. Louis Ricketts was to a large extent responsible for my opportunity. I had been lucky enough to do a lot of work for and with the Doctor, as he was affectionately called by the mining men of the southwest. My association was in connection with his work with John Greenway on the complex metallurgical problems at Ajo. Soon after the Ajo investigation Ricketts had been made chief consultant for the Anaconda Copper Co. When Anaconda decided to undertake a cam- paign of exploration of the west coast of South America Dr. Ricketts thought of me and arranged with Greenway to give me a leave of absence from my Calumet and Arizona job. It took three days in New York in June 1914 to arrange the expedition. Ben Thayer, executive vice president of Anaconda, did most of the planning and I was to report to him. T.H. France, a 45-year-old mining engineer who had been driven out of Mexico by the Madera Revolution, was to accompany and help me. He spoke excellent Spanish and his years of Mexican experience had taught him how to get along with Latin Americans. He went ahead to Bolivia, our first target, to get an idea of the country and the problems. This was lucky as I hardly knew where Bolivia was.

Jan 1, 1976

By F. Rosenblum, P. Cancilla

The manual moisture determination methods employed today by mineral processing plants falls short in providing timely information required for automatic control. The costs associated with transporting and handling concentrates still represent a major portion of the overall treatment price. When considering the cash flow of a mining operation, that is governed by both the smelter contract, with moisture penalties and the quantity and quality of the concentrates shipped, an efficient method of on-line moisture content would be a welcome tool. A novel on-line determination system for ore concentrate moisture content would replace the normally tedious classic manual procedure. Since the introduction of microelectronic¬based control systems, operators have strived to reduce the treatment costs to the minimum therefore a representative and timely determination of on-line moisture content becomes vital for control set points and timely feedback. Reliable sensors have long been on the "wish list" of mineral processors since the problem has always been that you can only control what you can measure. Today, the task of moisture determination is still done in the classical technique of loss in weight utilizing uncontrolled procedures. These same methods were introduced in the earliest base metal concentrators. Generally, it is acceptable to have ore concentrate moisture content vary within a range of 7 to 9%. Many times, delays in manually achieving reliable feedback on the moisture content results in the moisture varying from 5 to 12% before corrective actions can be made. This paper first reviews the traditional and widely available methods for determining moisture content in granular materials by applying physical principals and properties to measure moisture content. All methods are in some form affected when employed on mineral ore concentrates. This paper describes a unique and promising on-line moisture sensor in two mineral processing applications, which not only automates the tedious tasks but also results in reliable moisture feedback that can be used in the optimization of the de-watering process equipment such as pressure or vacuum filters and fuel-fired driers. Finally, two measurement applications will be presented to indicate the usefulness and to summarize the measurement requirements for the proposed method of employing drag force and mechanical properties of the material itself to determine the moisture content.

Jan 1, 2002

By G. Robert Guillet

Nepheline syenite is a light-colored feldspathic rock that is used primarily as an addition to the glass batch. It is also used in ceramics and as an industrial filler. Composed essentially of the soda and potash feldspars and nepheline, nepheline syenite contributes needed alkalis and alumina to the glass composition. And because it contains a higher proportion of these oxides per unit weight, it is a formidable competitor for feldspar in northeastern North America and parts of western Europe. Nepheline-bearing rocks are not uncommon on a world scale, but few are able to be beneficiated to a composition sufficiently low in iron to be used in the making of glass. Nepheline can only form in an environment deficient in silica. Hence free silica, as quartz, is always absent in the primary mineral assemblage. Biotite, hornblende and magnetite are usually present as minor accessory minerals, and zircon, garnet, muscovite, and apatite may be present as trace minerals. Nepheline syenite is produced for glass, ceramics, and filler markets by Unimin Canada Limited at Blue Mountain, in south- eastern Ontario. Unimin acquired the former Indusmin and IMC properties from Falconbridge Ltd. in 1990. Elkem Nefelin AIS commenced production of similar products from Stjernoy Island in northern Norway in 1961. Unimin acquired this company in 1992. Nepheline has also been a byproduct of apatite mining on the Kola Peninsula for many years, and in 1951 the production of alumina from nepheline was commenced in a plant near St. Petersburg (Leningrad). There are now four plants in the Russian Republic, two smaller ones based on nepheline tailings from apatite mining and two larger and more recent plants in central Siberia using other nepheline resources. All four of these plants are integrated complexes that produce alumina for the subsequent production of aluminum metal, sodium and potassium carbonates, and portland cement. Roofing granules are produced by the 3M Company from a nepheline rock in Arkansas. In 1980 and 1989, trial shipments of nepheline syenite were recorded from new sources in Brazil and western Turkey respectively. Recent interest is also recorded for deposits in China, Saudi Arabia, and British Columbia.

Jan 1, 1994

By Stephen Krickovic

"Bleeding", as used in mine ventilation language, has a number of meanings among mining men and is regarded with different degrees of importance. This lack of uniformity stems from a number of specific situations. A method that is adequate in a mildly gaseous mine may fall short in a very gaseous mine. Projected bleeders could be dependable under a certain type of roof, and totally unsatisfactory under a different type. Similarly a plan of bleeders can be rendered useless due to undulations of seam and water in one mine and function well in another. Finally, what is effective in a mining system with partial recovery will not suffice with full recovery. You may wonder, then, if there is a common denominator for the essential function of "bleeding". In my opinion, there is, and it can be defined as controlled movement of air across an active pillar line toward the back and outby ends of the adjacent gob area into the headings provided for the purpose. The objective is to maintain as methane-free as practicable a belt adjacent to the working places, so that, if face ventilation is interrupted temporarily, the gas will not accumulate in dangerous volumes at the working faces before ventilation is restored. It's possible for these temporary interruptions to be unnoticed in time or at all. In such cases, ineffective bleeding could cause a disaster. This has happened. To control, there must be power. That is to say, there must be adequate air, pressure differential between the front and back ends of a pillar line to facilitate proper regulation for diverting of air to the various portions of the gob and as far away from the active section as is practicable. As an indication of the importance of pressure, it is noteworthy that one may provide and maintain in a reasonably open condition a heading or two for a bleeder system and fail to obtain effective bleeding .because air pressure differential is inadequate. On the other hand, you are all familiar with the fact that pillar lines in some seams can be ventilated adequately by air flow straight across the gob when the section is near a main return shaft and thus obtains a heavy pressure differential.

Jan 1, 1962

By William H. Dresher

Introduction A subject long debated but rarely resolved involves the interaction between the academic community and industry in the sponsorship and conduct of research. The subject is, unfortunately, too often treated as a controversy involving ethics and the profit motive. But, in fact, the two branches of society - industry and academia - are necessarily part of the same system. In the last 25 years, the realization has developed that research alone is not sufficient to bring about use of those results to benefit society through productivity gains and economic growth. The flow of events shown in Fig. 1 was generally believed to be the steps required for reaping the benefits of research. Today, research still has the very important role of creating new information or new ideas. However, it is recognized to be but one of many steps in a process that we now refer to as technological innovation. Charpie (1967) said, "Technological innovation is the driving force behind economic growth in the US. In this context, innovation means the process by which a new idea is successfully translated into economic impact within our own society, by providing better products, and simultaneously creating new jobs in the manufacturing of those products. Only after an investigation is put into sufficient use to have an economic effect, is it termed an innovation." One of the best depictions of the technological innovation process is shown in Figs. 2 and 3. These are adapted from Haeffner (1975), and modified by the Committee on Mineral Technology of the National Research Council (1981). Figure 2 identifies the many steps involved in the process. It designates a generalized form for interaction between the academic community and industry. Figure 3 identifies the steps in the "innovation activity" box of Fig. 2. Academia is generally responsible for developing of new knowledge and ideas, as well as maintaining and enlarging the collected body of scientific knowledge. Industry is responsible for developing new products, methods, and processes. Both institutions operate for the benefit of society, but under different driving forces. In academia, the driving force is the achievement of academic excellence. In industry, the driving

Jan 10, 1984

By M. A. Trevits, G. L. Finfinger

Various methods have been developed through US Bureau of Mines research to remove gas in advance of, during, and subsequent to mining. These include drilling horizontal or cross-measure boreholes from a mine opening, and vertical boreholes stimulated for drainage in advance of mining or for gob gas drainage. Techniques and equipment now exist that allow drilling a borehole at any angle to the vertical, and even along an arc. Stimulation treatments have also been developed that provide for increased gas production capability. This paper describes the long-term production history of five select gas drainage installations. The Bureau of Mines conducted or participated in four studies. The remaining study was conducted by the Equitable Gas Co. Principles of storage, release, and migration of coal bed gases The amount of gas contained within any given area of a coalbed depends on the coal's inherent storage capacity and the physical conditions of the coalbed. The gas storage capacity of coal is primarily a function of rank. As such, it cannot be affected by the presence of drainage holes (Kim, 1977). Certain physical conditions of the coalbed can, however, be changed that will affect the release and migration of gas. By pumping water out of the coalbed through the vertical boreholes, two important changes occur - pressure is reduced and water saturation is lowered. Reducing reservoir pressure releases gas as the coalbed adjusts and equilibrates gas to the change. Increasingly large amounts of gas are released for each incremental decrease in pressure. Lowering the water saturation level raises the permeability of the coalbed to gas and allows the gas to more readily migrate to the wellbore (Kisse and Edwards, 1975). For horizontal boreholes completed near mine openings, the coalbed has already undergone a pressure or water saturation reduction phase. This condition thus allows such boreholes to be highly productive with minimal water production. Case studies of coalbed gas production Pittsburgh Coalbed - Marianna mine 58 A methane drainage study was begun in July 1978 at Bethlehem Mines Corp.'s Marianna mine 58, Marianna, PA (Fig. 1). The work was conducted in a section of the mine that was advanced 1.5 km

Jan 8, 1986

By Syd S. Peng

Introduction Modern longwall mining, introduced to the US coal industry in the mid-1960s, is the latest coal mining technique. Today, longwall mining produces more than 15% of all underground coal production. The growth of longwall mining in the US is slow. However, nearly two decades of longwall mining have demonstrated that its benefits outweigh the disadvantages. Longwall Advantages Production and Economics - A typical longwall production ranges mostly from 0.9 to 1.8 kt (1000 to 2000 st) of clean coal per shift. This is about three to six times the production of comparable continuous mining units. Furthermore, there are much less rejects in longwall mining, typically 15% to 25%. It must be noted, however, that a new longwall with an inexperienced crew may produce less, depending on mining conditions. But as experience gains, production increases rapidly. Due to its high production, the cost per ton of coal mined over the mine's lifetime is cheaper with longwall mining than with continuous mining. A study on a conceptual mine of 25 year life using actual operational data (Dangerfield, 1981) concluded that the cost per ton of coal mined is 32% cheaper by longwall than by continuous mining. Another benefit from its high production potential is that it enables the production to be concentrated in a few longwall faces versus many units of continuous mining required to achieve a similar production. As a result, mine organization and management simplify considerably. Safety and Savings - Most US longwalls use shield supports that cover the roof with solid canopy and isolate a gob completely from the face area. Thus, if the supports are properly selected and operated, the potential for roof fall accident is almost zero. Furthermore, it is not unusual that a well-run longwall face is much more orderly and cleaner that most well-run machine shops. The ventilation system for longwall mining is the most ideal type. It is simple and unique. No auxiliary fans are required. The fresh air sweeps through the whole face. Its volume and velocity can be adjusted as demanded. Recovery - In spite of multiple entries layout for US longwall panels, the longwall method recovers up to 40% more coal than the continuous mining method (Dangerfield, 1981). The typical recovery for longwall mining is 70% to 85% versus 45% to 80% for continuous mining. Versatility - Longwall mining has been employed successfully in seam heights ranging from 0.8 to 3.7 m (2.7 to 12 ft); in seam inclinations from horizontal to 35°; in overburden depths from 46 to 914 m (150 to 3000 ft); in single or multiple seam mining; and in uniform or irregular seam characteristics. For seam height less than 1.2 m (4 ft), the plow is used instead of a shearer. However, in spite of recent developments, the plow still is unsuitable for hard coal seams. Slow Acceptance How come longwall mining, with so many advantages, has not spread more rapidly? Undoubtedly, the major reason is the large capital investment required. Under normal market conditions, the capital required for a longwall panel ranges from $7 to $11 million. Another factor is the uncertainty of its applicability to any specific coal seam of interest where longwall has not been used or operated successfully. Although longwall mining elsewhere has proven its applicability to wide-ranging seam conditions, there have been several failures in the past two decades. Analyses of those cases show that most failures can be attributed to inexperience, both operational and technical, including equipment selection. Therefore, personnel training on the longwall mining technique is absolutely necessary before its initial operation. Some earlier statements about adverse surface subsidence associated with longwall mining were misleading. Research has proven that surface subsidence under longwall mining is controllable and predictable. And, if longwall mining was mandatory since the onset of coal mining in the US, there would be no surface subsidence problems over abandoned mined land today. It is a blessing that surface subsidence occurs

Jan 3, 1985

By G. W. Johnston

INTRODUCTION For many years mining at the Stobie mine of the Inco Metals Co. consisted mainly of fill-type mining for higher grade ores and block caving or blasthole open stope mining for lower grade ores. Open stope mining, although quite favorable in the initial stoping phase, has always presented problems in the pillar recovery phase, especially at depth. In 1966, sublevel caving was introduced. This method is ideally suited for low-grade ore and can be applied very successfully to intermediate or higher grade ores providing proper controls are enforced. Sublevel caving has three distinct advantages: it is highly adaptable to mechanization; it recovers the ore in one operation; and it provides adequate ground sup¬port and safety regardless of mining depth. This method does, however, make extra demands, particularly for new types of supporting services. The constant presence of rock makes quality control of the grade of the ore being extracted essential. Increased de¬velopment, different types of ventilation, roadways, and maintenance are only a few of the many changes re¬quired. The care and maintenance of the fleets of diesel equipment and the modification of equipment is a con¬stant challenge. CHOICE OF SUBLEVEL CAVING The question arises, why mine by sublevel caving? Fig. 1 shows a typical problem encountered by the open stope or blasthole mining method. Near the surface there is little or no ground support problem. As mining progresses to an intermediate depth, however, ground failure starts to occur in the supporting pillars. If open mining is carried beyond this point, the critical depth is soon reached, accompanied by hanging wall caving, ex¬tensive pillar failure, and the danger of losing control of the mining cycle. The rock caved from the hanging wall mixes with and dilutes the ore being extracted from the blasthole stopes. Ore losses occur when pillar failure pre¬vents drilling and development of part of the ore. The interruption to the mining cycle causes serious curtail¬ment in production until ground failure is controlled and the mining method changed to adapt to the new conditions. Having recognized the limitations of open stope min¬ing, alternatives were sought. The first alternative (Fig. 2) was to fill the open stopes with sand fill be¬fore extensive pillar deterioration occurred, recover the pillar ore by fill mining, and then convert to total fill mining below. If the ore grade is high enough it can support the costly fill method. Recovery will be at a much slower rate and will therefore extend over a longer period of time. On the other hand, sublevel caving gives complete primary mining recovery in a short time and is very adaptable to mechanization, which is now a necessity to combat spiralling labor costs and give greater produc¬tivity per man. Sublevel caving is ideally suited as a mining method for low-grade ore bodies which eventually reach a criti¬cal depth for open stope mining. DEVELOPMENT The development phase comes first and of course its main purpose is to gain access to the ore and provide a route for ore removal from the mine. Initially a 20% inclined service ramp is driven in the footwall rock (Figs. 3, 4, and 5). It is preferable to start at the top of a mining block and drive the ramp down¬ grade so gravity can assist in mucking the blasted de¬velopment round, and also access can be gained more quickly to the initial production area. The ramp nor¬mally turns at 3.1 rad (180°) about every 152.4 m (500 ft) to keep it reasonably close to the rock disposal

Jan 1, 1982

By Earl G. Hoover, V. S. Znamensky

For most geologists and miners the term asbestos refers to a group of highly fibrous silicate minerals that readily separate into long, thin, strong fibers of sufficient flexibility to be woven; these fibers are heat resistant and chemically inert, and possess a high electric insulation and therefore are suitable for uses where incombustible, nonconductive or chemically resistant material is required. The US Mining Safety and Health Administration (MSHA) specifically refers to the naturally occurring minerals chrysotile, amosite, crocidolite, termolite asbestos, anthophyllite asbestos, and actinolite asbestos as asbestos if the individual crystal fragments have the following dimensions: length-greater than 5 µm; maximum diameter-less than 5 µm; and a length to diameter ratio of 3 or greater. Since asbestos has been identified as one of the ten most carcinogenic substances (a cancer-causing agent), it has become, for many government regulatory agencies, a major issue concerning banning emissions to the environment. Very few recent issues have caused as much concern for the crushed stone industry as asbestos. Complicating the problem is the fact that several federal and state bureaus have greatly compounded the confusion because each has become involved in the asbestos dilemma. Simple solutions to a very complex problem have been proposed by these various governmental groups. In addition to governmental bungling is the fact that public interest groups have been able to use federal and state laws that have been enacted with a strong environmental bias to foment even greater complexities. Such is now the state of the asbestos problem. No con- firmed health risk has been associated with low-level exposure, namely crushed stone in an unbound form; a methodology for sampling and analyzing for asbestos has not been developed; a precise definition of what is asbestos does not exist, nor does a concensus among the various government groups as to what course of action to pursue. Pure asbestos as it is mined for commercial purposes does not exist in the properties mined by the crushed stone industry. Where the six asbestiform minerals do appear (to date we've been referring mainly to chrysotile in serpentinite formations in Maryland) the amount of the asbestiform mineral present is generally quite small. They generally appear in scattered, miniscule amounts. In the case of Rockville Crushed Stone in Montgomery county, Mary- land, the chrysotile is less than 1/2 of 1% of the rock deposit. No health risk has been identified with such minute quantities of asbestos minerals. MSHA standards for the occupational health and safety of quarry workers are being met; and there is every indication that normal background levels of asbestos, throughout the surrounding countryside, are as high or higher than levels measured on roads where the rock is used in pavement or base. The problem and the controversy lie with improper terminology on the part of so-called public interest environmental groups, aided and abetted by some governmental agencies who, to be charitable, we can best describe as uninformed. The news media frequently has fanned the issue, unintentionally we'd like to think, by quoting these misinformed authorities and not bothering to question the facts or adequately report balancing information handed to them. It is much better copy to talk about an asbestos cancer scare, than to deal with rather mundane measurements and descriptions that are hard to understand. At the heart of the issue is the improper use of the word asbestos. Instead of being confined to the geological description of the six asbestiform minerals-chrysotile, amosite, crocidolite, termolite asbestos, anthophyllite asbestos, and actinolite asbestos-the word asbestos is also wrongly used, but too-frequently to identify many other minerals that happen to have elon-

Jan 1, 1986

By Richard L. Breithaupt

INTRODUCTION A shuttle car is a self-propelled rubber-tired haul¬age vehicle designed specifically for underground min¬ing, primarily in coal mines (see Figs. 1 and 2). The drive mechanism is located on the sides of the vehicle, and the load is carried in the center. A chain-and-flight conveyor distributes the load when the car is being loaded by a loading machine or continuous miner and also discharges the load onto a conveyor belt or into a mine car. As the name implies, the shuttle car shuttles back and forth between the working face and the un¬loading point and is not required to turn around. Shuttle cars have served as primary haulage in this country since 1940 and still perform this function in most coal mines today. TYPES OF SHUTTLE CARS Shuttle cars may be battery-powered, diesel-powered, or cable-reel electric powered. Battery-powered cars were used between 1940 and 1950, but are almost out of existence today because of their limited power. Diesel cars have been used in coal mines, but are used mostly in noncoal mines where rules and regulations are not as stringent. Also, objections by organized labor have limited the use of diesel-powered shuttle cars in coal mines. Today, the largest number of shuttle cars are cable-reel cars. Electric power is fed through a cable stored on a cable reel that may have a capacity from 152 to 244 m (500 to 800 ft), depending upon the size of the cable and the cable reel (see Table 1). This, of course, limits the travel of a shuttle car. However, time studies have shown that for the most efficient haulage cycle, the length of haulage should not exceed 91 m (300 ft). As haulage distance increases, the waiting time for the loader or miner becomes longer, and haul¬age efficiency suffers. Cable-reel cars are manufactured in standard hand and opposite standard hand models. A standard car has the operator's control on the left front side and the opposite standard car has the opera¬tor's controls on the right front. Generally a loader or miner operates with two shuttle cars, one of each type. This allows the trailing cable to lie on either side of the entry and prevents the cables from fouling each other. Shuttle cars are designed to operate with either dc or ac power. Cars using dc power are designed for 250-v supply and ac cars are designed for either 440, 550, or 950 v, three-phase voltage. Most coal mines that are relatively level can use dc cars where dynamic brak¬ing is not required. For mines having steep grades (in excess of 15%) ac cars are used because ac traction motors have inherent dynamic braking and are safer when tramming downgrade. A recent innovation in shuttle car design in the last five years has been the introduction of an ac/dc car. This design utilizes ac input power through the cable, and the power is rectified on the car to 250-v do for traction motors. The pump motor and conveyor motor are powered by ac voltage. The main advantage of this design is efficiency-a much smaller cable may be used and the smooth operating characteristics of the dc traction motor is obtained. This prevents shock loading on the traction drive train, which is an undesirable feature of ac traction motors. SHUTTLE CAR FEATURES Elevating Discharge Today, all shuttle cars are equipped with an elevating front conveyor that allows the shuttle car to discharge into a mine car or belt feeder without the use of ramps. Hydraulic Power System Today, all shuttle cars are equipped with an hy¬draulic power system. An hydraulic pump is driven by an electric motor which provides hydraulic power for steering, elevating the front conveyor, and powering the

Jan 1, 1982

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader's orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature's effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By B. P. Boisen

INTRODUCTION Application Measurement of rock deformation is essential to stability and behavioral monitoring of underground openings. The difficulties of obtaining adequate sample distribution, density, and orientation can frequently be overcome by the use of single and multiple position borehole extensometers. In its simplest form, a borehole extensometer consists of an instrument head, usually anchored in the collar of a borehole, and a number of in-hole anchors, each of which is fixed in place at a known distance from the collar. As rock adjacent to the borehole deforms, the component parallel to the borehole direction causes a distance change between adjacent in-hole anchors, as well as a distance change between individual anchors and the instrument head. These changes are measured by means of a mechanical link (usually wire or rod) extending to the instrument head from each anchor. The resulting data can be used to compute the distribution, magnitude, rate, and acceleration of deformation in the material intersected by the extensometer's borehole. Wire Vs. Rod Extensometers Since the early 1960s myriad varieties of borehole extensometers have become available through the combination of various special features of two basic instruments, the wire and rod versions of borehole extensometers. Recent publications have recommended shallow depth limitations for wire extensometers, and favor rod extensometers for a broad range of borehole inclinations and depths. There are three main reasons for using a deep wire system: its cost effectiveness; the fact that tremendous depth means tremendous rod weight which will cause the rod to buckle or rub against the borehole or casing wall; and users' attitude and proper understanding of how to install and apply a wire-type extensometer. Here experience plays a big role. Over the years, many have learned the hard way that untensioned rod extensometers are simple and inherently inexpensive, but limited to vertical up holes where sensitivity is not influenced by friction between the rod elements and the borehole wall or protective conduit, the inertia of the rod elements, or the direction of the movement to be measured. In deep, horizontal, or inclined holes, and in instances where compressive as well as tensile strains or displacements are anticipated, tensioned rod or wire systems are preferred. System Components Borehole extensometers are made up of at least the following five principal components: the instrument head; the collar anchor assembly; the downhole system of anchors and rods or wires; the signal cable if the head is equipped for electronic readout; and the readout apparatus, whether mechanical or electronic. These components are shown in Fig. 1. Borehole Requirements Most extensometers are designed for installation in nominal 75 to 57-mm (3.0 to 2.25-in.) diam drill holes. Either a diamond drill or a percussion drill hole is satisfactory, provided the variation in drill-hole diameter is within the tolerance of the standard anchor design. The basic system may be modified for other hole diameters by changing the dimensions of the collar anchor and downhole anchor assemblies. Some instrument heads and collar anchor assemblies require enlarged borehole collars, though usually not over 150 mm (6 in.) for a depth of 1 m (3 ft). It is desirable, although not essential, that holes be free-draining. Holes must be clean and free of debris before installation. Depth/Orientation of Extensometer Boreholes Extensometers are designed to measure axial strain components in the instrumented drill hole. As the drill hole is axially deformed, the distances between individual downhole anchors and the instrument head change, as do the distances between adjacent anchors. The distance changes measured by a multiple position borehole extensometer, for example, are reduced to functions of time, hole depth, anchor position, or other parameters of significance in terms of the structure under investigation.

Jan 1, 1982