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By Irwin H. Reiss

Gentlemen, the title of my paper is Surface Mining An Interim Land Use. Surface mining has become one of the major controversies of our time. This has happened in spite of the fact that all surface mining to date has involved only 6 million acres out of a total of 2300 million or approximately one-fourth of one percent of the total land area in the United States. I'm not going to dwell on whether peoples' perspectives are out of focus in light of the area involved, the energy crisis, national security, or the international balance of payments problem. While these things are all critically important, it is not the function of this paper to relate to them. Maybe the amount of attention attracted to surface mining is out of proportion to a lot of other things, but I don't feel that we should use that as a basis for procrastination. However, let me categorically state that anything that affects the surface of the earth is a form of land use --- be it an interstate highway, a power plant site, a golf course, a corn field, a strip mine, or the plot of ground on which your house is located.

Jan 1, 1974

By K. C. Kocsis

Underground ventilation systems, which require electricity to operate the fans and fossil fuel for heating in winter, can account for a significant portion of the energy consumption of mines. Recent studies have shown that the ventilation system in deep and highly mechanized metal mines could be responsible for 40 to 60 percent of their overall energy consumption. Presently, underground ventilation systems are designed for peak air volume requirements, which usually occur well into a future stage of the mines’ operating life. As a result, during their early operating stages, the intake air volume of the mines could be well in excess of their true air volume needs. Such ventilation systems could be considered as uneconomic, and to remain competitive the ventilation design needs to improve while the mines are trying to reduce energy consumption and their carbon footprints. This paper introduces a new mine ventilation design method where an underground base-metal mine’s ventilation system is designed by integrating discrete-event mining process simulation (DES) with mine ventilation modeling techniques. Furthermore, based upon the output data generated by means of discrete-event mining process simulations, the air volume requirement for the operating life of the mine is determined for “traditional” versus “activity-based” ventilation requirements. The output data generated through ventilation simulations are then used to quantify the economic and environmental benefits of an activity-based ventilation delivery system.

Jan 1, 2018

By John D. Rockaway

This report and accompanying engineering geologic map describe the generalized engineering geology of St. Louis County, Missouri for use in determining the most efficient, economical, and environmentally sound program for land development and urban growth. It includes an evaluation of such factors as the physical and engineering properties of soil and bedrock, surface stability, and subsurface geologic structure. The map and text are designed primarily for property owners, engineers, architects, real estate developers, and city planners involved in land-use planning. The geologic information and engineering interpretation must be considered as a general guide and not as a substitute for a complete on site investigation prior to initiation of any land-use project. The map scale, 1:63,360 (1 inch to a mile) is too small to include the detailed information normally provided on the 7 1/2-minute quadrangle series (1:24,000). The smaller scale is most useful at regional planning levels and was chosen to present information on the more significant engineering factors of the geologic environment for county-wide planning and development. A conventional geologic map deals primarily with the distribution and variation of bedrock formations, and a pedologic map presents information on the soil or natural medium for plant growth. Basic geologic maps and pedological maps are valuable tools that can be of assistance in studying the geologic environment, however, these maps do not provide all the information required for land-use planning. The engineering geologic maps are designed to provide this data. The engineering geologic map is related to both of these basic maps but emphasizes the engineering properties of the rock formations and soils rather than their age or origin. Some of the unit boundaries shown on the county map will closely follow some geologic formation and/or pedologic boundaries; others will transgress these conventional boundaries. Earth materials, bedrock and soil, that have similar engineering characteristics have been mapped as one unit without regard to their geologic or pedologic classifications. This is to assure that geological factors are recognized, adequately interpreted, and presented for use in engineering practice.

Jan 1, 1970

By Ali Sogue, Steven Schurman, R. Jensen

This paper presents the results of the remediation of the Gregory Gulch #1 Superfund site, an EPA Clear Creek Superfund project located in central Colorado. The Gregory Gulch #1 site was an historic gold mine and mill located within the city limits of Central City, Colorado. Remediation efforts included the use of the Envirobond™ additive to chemically bind the metals in 3,000 tons of acid generating mine waste that contained high levels of arsenic, lead, zinc and mercury. The toxicity of the mine waste amended with Envirobond™ was not only reduced to RCRA TCLP standards but also to the more stringent UTS standards. The final cost of the treatment was much less than standard industry solidification processes and significantly less than offsite disposal options.

Jan 1, 1999

By Jozef Stachulak

Prof. H. Bystron: "An Application of the First Law of Thermodynamics to Winter Dry Air Flow in Mr. Stachulak's Mine Model". [ ]

Jan 1, 1980

By N. R. Sainath

This paper discusses the results of a laboratory investigation carried out to study the microscale flow characteristics of copper ores and the variation in flow behavior resulting from flow of leach solution. Fractures in 1.5 and 4 inch diameter cores were mapped by visual inspection. The cores were stressed in a triaxial stress field representative of in situ conditions. The permeability of the samples to water was then estimated. The samples were exposed to 20 g/l sulfuric acid and the experiment repeated. The Permeability of unleached samples ranges from 10-1 to 10-19 m2 and it increases as leaching progresses. The porosity variation and the rate of recovery of copper from the specimen during leaching were also studied.

Jan 1, 1991

By John H. Bassarear

Large scale equipment can reduce both capital and operating costs for concentrators. Large semiautogenous mills can be effectively utilized in big mills but they must be kept fully loaded to take advantage of their great potential. With varying hardness ores, controlling throughput is difficult and careful design must take this into consideration. A variable speed drive on large primary semi- autogenous mills can overcome this problem to a considerable degree. This paper provides a base case using standard accepted technology of two wound rotor single speed 4,476 kW (6,000 hp) motors to drive the primary SAG mills of a hypothetical 36,364 mtpd (40,000 stpd) typical porphyry copper mill. One other fixed speed drive will be described and five variable speed drive configurations will be reviewed. Capital costs for the entire grinding circuit, from crusher product to flotation feed, were developed for the base case and each of the alternatives. A technical evaluation comparing the advantages and disadvantages of each drive is presented.

Jan 1, 1985

By Werner A. Ruemmele, Steven R. Wardwell

INTRODUCTION The first phase of the Los Angeles Metro Rail Project consists of 7.1 km (4.4 miles) of subway and stations running through downtown Los Angeles from Union Station to MacArthur Park. Shank-Ohbayashi has two of the tunnelling contracts on this initial phase: A-171 and A-146. Contract A-171, the Wilshire/Alvarado Line, involves tunnel construction beneath 7th Street from the 7th/Flower Station to the Wilshire/Alvarado Station. The project consists of the excavation and lining of 3000 m (9900 ft) of twin-bore, 6.7 m (22 ft) excavated diameter, 5.5 m (18 ft) finished diameter subway tunnels. Contract A-146, the 5th/Hill to 7th/Flower Tunnels, comprises 1220 m (4000 L.F.) of twin-bore tunnel identical in cross section to the tunnels in Contract A-171. The majority of the excavation on Contract A-171 is beneath city streets in a weak siltstone/claystone bedrock with cover varying between 10 m and 25 m (32 ft and 82 ft) to the street surface. Approaching the 7th/Flower Station some fine-grained alluvium is encountered. The tunnels on both contracts are excavated with an open face digger equipped shield. On Contract A-146 the tunnel alignment is a 305 m radius curve of approximately 550 m (1800 L.F.) length underneath several multi-storied buildings of various construction. Excavation is still ongoing on Contract A-146. The contractor's choice for initial tunnel support on both contracts is a system of precast concrete segments similar to what the contractor had used on 7 miles of tunnel in Phoenix, Arizona, for the Arizona Department of Highways.

Jan 1, 1989

By Larry D. Wayland

Almost 100 sedimentation ponds were conceptually designed for a large surface mining study area in northeast Texas. An approximate procedure was developed to economically estimate construction quantities in order to predict surface water control costs. This procedure utilized site-specific empirical relationships developed from detailed analyses on a representative number of proposed sedimentation ponds. Use of these equations provided earthwork volumes, and spillway pipe lengths. The procedure developed for this study is presented along with the results of a verification analysis.

Jan 1, 1983

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today's DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine's ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982

By Gregg J. Hodges

The Stillwater pt-pd mine and flotation mill are located about 130 km (80 miles) southwest of Billings in Stillwater county, M6ntana~ The pt-pd ore zone stretches about 43 km (27 miles), but varies in thickness from 0 to 6 meters (0 to 20 feet, but generally 3-6 feet) and dips 55-90,degrees from horizontal. The ore is hosted in basic, layered igneous rocks containing about 1% sulfide minerals which include chalcopyrite, pentlandite, pyrrhotite, and minor pyrite. The pt-Pd minerals are intimately associated with these sulfide minerals and are found as precious metal sulfides, tellurides, bismuthnides, and arsenides as well as native Pt metal. Approximately 90 different precious metal minerals have been found, but about 80% of the pd is in solid solution in the pentlandite. The ore assays about 0.1% Cu, 0.1% Ni, 6.2 g Pt/tonne (0.18 Troy ounce Pt/short ton (opt)), 21.3 g Pd/tonne (0.62 opt Pd), 0.21 g Au/tonne (0.006 opt Au), and 0.21 g Rh/tonne (0.006 opt Rh). The ratio of pd to Pt in the Stillwater ore and concentrates is about 3.5.

Jan 1, 1988

By Richard L. Bullock

OPEN STOPING An open stope is an underground cavity from which the initial ore has been mined. Caving of the opening is prevented (at least temporarily) by support from the unmined ore or waste left in the stope, in the form of pillars, and the stope walls (also called ribs or abutments). In addition to this primary support system of open stoping, some secondary support may also be required using rockbolts, reinforcing rods, split pipes, or shotcrete to stabilize the rock surface immediately adjacent to the opening. The secondary reinforcement procedure does not preclude the method classified as open stoping. There are many forms of open-stope mining used to extract the initial material from a mine. Having once established that the mineral and waste rock are competent enough to use an open-stoping method, and assuming that the reserve is not classified as gassy, the Form which the method will take is primarily determined by the dip and thickness of the reserve. How these two factors affect the selection of the open-stope mining is discussed in a later chapter. At this point it will suffice to say that the classification of the open-stopes mining system which follows is based on whether dry' broken materials flows by gravity or whether it must be moved by nongravity methods where energy must be supplied to move the material. Room-and Pillar-Mining Room-and-pillar mining is an open-stoping method where mining progresses in a nearly horizontal or low angle direction by opening multiple stopes or rooms, leaving solid material to act as pillars to support the vertical load. Since the direction of excavation (angle of dip) is below that which would cause the dry material to flow by gravity to a drawpoint or gathering point, the material must be loaded in the room where it was extracted and transported to a point where it will flow, either by gravity or mechanical means, to a central gathering point to be taken out of the mine. This is an important aspect of room-and-pillar mining which differentiates the system from other open-stope mining methods which rely heavily upon gravity to transport ore from where it was broken to a lower elevation, usually through a drawpoint. There are many variations of the method which go by a number of names in local districts: breast stoping, breast-and-bench stoping, board-and-pillar, stall-and-pillar, and panel-and-pillar are all basically open-stope room-and-pillar mining. In some instances detailed stope planning is almost nonexistent; i.e., the operator simply follows the visual pay values, leaves pillars only where necessary, and tries to locate them in the zones of lower value. This method of mining is as old as the beginning of underground mining itself, dating back thousands of years. Early in the history of mining in this country, the term "gophering" was used to describe this method (Peek 1941). The term is appropriate, for it brings to mind the exact results of this type of system-a random and irregular room-and-pillar mine. In other instances where the mineral values are consistent both in physical dimensions and quality, the mine layout can be planned to the last detail, resulting in a uniform room-and-pillar mine. Coal, trona, gilsonite, potash, oil shale, salt, limestone, and sandstone mines can usually follow such a system. Today, most metal mines using a room-and-pillar operation try to mine as regular a pattern as possible but deviation in height, width, thickness, dip, and grade of the ore results in comparable deviation in the mine plan. Variations of the Room-and-pillar System It is necessary to briefly describe some of the many variations of the room-and-pillar system of mining, enabling the reader to fully explore the concepts and become familiar with the terminology used before going on to the details of mine design. Full-Face Slicing: If in the process of opening the rooms the total vertical extent of the mineral values of the particular seam or strata are extracted from the advance of one vertical face, the term used to describe this is full-face slicing. This face is also known as the "breast." There is no mineral of economic value intentionally left either in the floor or the roof (back) to be mined later. To be able to extract the full-face height in one pass, the mining equipment must obviously be designed to reach as high as the back. In an Appalachian coal mine, this may be all of 660.4 mm (26 in.); for a future oil shale property it might be 15.24 m (50 ft). Normally, however, in a majority of mines where the mining face gets over 6.09 to 6.7 m (20 to 22 ft), the tendency is to divide the face into more than one pass. Over this height, it becomes difficult to properly see and remove loose rock from the back with a hand "mining bar." Where the process of taking down loose rock has become mechanized, higher full-face mining can be safely practiced. Most eastern and midwestern coal seams and western uranium, trona, and potash seams in the United States are easily reached in a single face; many limestone, lead, and zinc mines must resort, at least in part, to "multiple-slicing" to remove all the minerals of value. Multiple Slicing (also known as multiple-pass mining): In many cases it is not practical to carry the full vertical height of the mining horizon as a full face. The face is divided into parts known as the breast, bench, and/or brow. Ideally, if the operator knows the vertical extent of the mineralized zone, he will drill and blast the first pass at the top of the zone, thereby

Jan 1, 1982

By F. Wolf, R. Ellmies

"Fluorspar is one of the most important industrial minerals. It has a wide range of applications, for example it is used in the steel and aluminum industry, and is also the raw material to produce hydrofluoric acid, Teflon or Gore-Tex. There are only a few major fluorspar mines worldwide, with China being the most important fluorspar producer (U.S. Geological Survey National Minerals Information Center, 2017). At Okorusu, Namibia, fluorspar has been mined with conventional open pit mining methods from 1955 until 2014 from a large scale metasomatic fluorspar deposit with seven identified orebodies. There are about 7 million tons of measured resources left with an average grade of 27.5 % CaF2. This paper is based at an MSc thesis at Technical University Freiberg and describes the design of an underground mine for the A and D orebodies with an annual production rate of 100,000 t acid-grade fluorspar product. By comparing different mining methods and adapting variants of access to the underground orebodies the paper will emphasize on the selection of a suitable mining technology, with a special focus on the economic considerations to decide about the most profitable operation. INTRODUCTION Mining holds an important part in the economy of Namibia. In 2016 11.1 % of the GDP was made through mining activities with 22 active mines (The Chamber of Mines of Namibia, 2016). Only one of them consists an underground mining operation. The global annual production of fluorspar is approximately 6 to 7 million t (U.S. Geological Survey National Minerals Information Center, 2017). Okorusu produced on average about 50.000 t annually in recent years (U.S. Geological Survey National Minerals In-formation Center, 2015). Okorusu Mine is Namibia’s only fluorspar mine. Fluorspar was mined in an open pit mine with four individual pits (A-D) since 1980. In 2014 the mine closed due to the increased stripping ratio and beneficiation problems. The mine assets were bought by Gecko Namibia. Currently Gecko uses parts of the processing plant for beneficiation of graphite from a neighboring mine. Figure 1 shows a panoramic view with the four pits and the mine village. The intent is to develop a concept for the further underground development of the mine. There are still about 7 million t of measured resources left in the deposit, with an average grade of 27.5 % CaF2. This study focusses on the underground mine design. The above-ground plant and facilities are considered as remaining unchanged."

Jan 1, 2018

By A. G. Provost

On April 3, 1981, Harrison Western Corporation was awarded a contract by Amax Alloy - A Division of Amax Inc., to sink the Henderson #5 ventilation shaft. The project was located on the Henderson Mine site 88 kilometers west of Denver near the town of Empire, Colorado (Figure I). The Henderson No. 5 Shaft, located in an Alpine environment at an elevation of 3145 meters, was a 9.8 meter diameter concrete lined shaft with a total depth of 793 meters. Work started and continued with Harrison Western employees until January 15, 1982. At that time, Amax made the decision to continue the shaft-sinking with their own labor, and with Harrison Western providing supervision under a management service contract. Shaft sinking continued under this arrangement to October 1, 1982, at which time Amax put the Henderson No. 5 project on standby status until further notice.

Jan 1, 1986

By J. K. Szymanski, Val Srajer

In this decade mining companies in western Canada have searched for reliable and inexpensive ways to improve mine productivity. Truck dispatching systems, equipment vital sign monitoring and sensing, vehicle guidance systems, have all been considered. In this paper the off -highway vehicle monitoring systems used or planned to be installed at surface coal mines in western Canada are described. The monitoring systems are divided according to their function and application; installation and testing procedures are described, as well as the economic benefits to the mining operation.

Jan 1, 1989

By Thomas R. Yanske, Jonathan E. Carmack, Lyndon M. Clark

In order to maximize the extraction of high grade pillars at The Doe Run Company, artificial pillars are created using cemented rock fill (CRF). CRF pillars aid extraction of pillars in two ways. One is to offer back support for the immediate area where pillars are being extracted. The second is to increase the post failure residual strength of pillars surrounded by backfill. The CRF is produced using a combination of minus 1.0 m run of mine (ROM) rock and minus 5 cm crushed rock. Depending on the operation the CRF slurry is either produced in an underground or a surface batch plant. The slurry and rock are transported to the fill area by truck where they are mixed and placed in 1 m lifts. To reduce the amount of CRF needed, fill fences are constructed around the perimeter of the CRF pillar. Within these fences ROM CRF is placed in lifts until within 0.6 m of the back after which minus five centimeter CRF is slung tight to the back utilizing slinger trucks. In constructing these artificial pillars a number of set procedures have been established to ensure that the work is done both properly and safely.

Jan 1, 2001

By David E. Sims

Asarco has been using computer-aided open-pit mine design since 1966. The in-house developed program package includes drill-hole analysis programs, interpolation routines, and three-dimensional open-pit design programs. The financial analysis step after pit design required that a reasonably accurate schedule be produced so that a good cash flow model could assess the feasibility of a project. The manual method of scheduling was too time-consuming for feasibility work, so a computerized scheduler was required. Since no computerized mine scheduling programs existed, we requested proposals from consulting firms to develop such a program. The selection of a consulting firm, the formulation of the contract and design specifications required about three years to accomplish. The consulting firm, Systems Control Inc. (SCI), of Palo Alto, spent over one year doing the programming and testing alone. The details of the program development were reported in "Computerized Year-by-Year Open-Pit Mine Scheduling," SME-AIME Transactions, December 1974. In June of 1973 the program, which SCI dubbed Optimum Resource Extraction or ORE, was installed. Version 1 of ORE was utilized successfully on three separate properties over a period of two years. During this time it became obvious that some of the original design specifications were restrictive, and certain modifications were required. The ORE program was submitted for modification, and after one year version 2 was operational in May of 1976. The program now consists of approximately 20,000 lines of FORTRAN source statements grouped into 190 subprograms, and runs in a 450K region of Asarco's 370/148 in Cranford, NJ. The Tucson com¬puter (370/115) is tied to the Cranford computer via a 9600 baud leased T.P. line, for running mine scheduling and other large mine planning programs.

Jan 1, 1979

By D. M. Bodily, V. J. Hucka

INTRODUCTION Interest in the recovery of methane from coalbeds has increased rapidly over recent years, because of its potential as an energy resource and to alleviate pollution and safety problems. The resource base of coalbed methane in the United States has been estimated at 11.2 Gm3, roughly equivalent to 20 years consumption (Mroz et al., 1983; Kuuskraa and Brandenburg, 1989). Methane contributes to the adverse environmental impacts of global warming, tropospheric ozone formation (smog), and potentially stratospheric ozone destruction. Association of methane with coal has always presented a safety problem in mines due to the potential for explosions. The rate of release of methane depends on the amount adsorbed by the coal, the adsorption isotherm for the coal, and the size of the coal particles. Mining operations can affect the particle size distribution and the production of fines and therefore the rate of release. The time the mined coal remains in the mine may also be important in determining total emissions. The methane content of some Utah coals was determined in this study. The methane content of the coals is compared with the chemical and physical properties of the coals and the properties of the seam and neighboring strata. Correlations are developed to better understand the formation and retention of methane by these coal seams. METHANE FORMATION IN COALBEDS Three mechanisms for methane formation may be considered: (1) nonbiogenic formation, (2) formation by anaerobic bacteria during plant material deposition and the early stages of diagenesis, and (3) formation from pyrolysis of coal material during coal maturation. Nonbiogenic formation by thermal processes involving carbon- and hydrogen-containing compounds has been proposed as a mechanism for methane formation in the earth. It is not likely to be of importance where other fossil fuel materials are present. Coal would not survive the temperatures required and hence the methane would need to migrate and be trapped by the coal seam. This mechanism is unlikely for coalbed methane. Formation of methane by anaerobic bacteria undoubtedly occurs during coalification, but the methane may not be trapped until the pore system of the coal is developed, and would likely be displaced by methane formed during further coalification. Methane formed by maturation of the coal material as it progresses from peat to coals of various rank is the most probable mechanism for the formation of coalbed methane. Most of the methane is adsorbed onto the walls of microvoids in coal. It has been estimated that the surface area of one kilogram of coal is between 20,000 m2 to 200,000 m2. Surface area measurements depend on the coal and also on the method of measurement. At 20 atmospheres, in some coals, the adsorbed methane is 10 times greater than methane as a free gas. At 100 atmospheres the adsorbed gas can be as densely packed as a liquid. The volume of methane gas held in coal depends on the volume of gas that exists as a free gas and the adsorbed gas capacity. The volume of free gas in the coal can be calculated from the kinetic theory of gases (methane can be assumed to be an ideal gas because errors would not be greater than 1 % for normal mining pressures). Langmuir (1916) derived an equation for the adsorption of gases on uniform surfaces and this is applied to the adsorption of methane on coal:

Jan 1, 1993

By L. M. Yokum

Tremendous strides have been made in front end loaders in the last fifteen years. Fifteen years ago a 2; cubic yard front end loader was the largest standard machine you could purchase. Today there are two commercially available 15 yard front end loaders and many 10 yard front end loaders. The estimated total worldwide market for all front end loaders over one cubic yard has increased to over 30,000 units yearly. The trend for front end loaders, in general, and for the mining market, in particular, is to larger and larger loaders. Today there are 18 cubic yard and a 24 cubic yard front end loader in prototype form. It is not inconceivable that within another fifteen years that there will be front end loaders in the 50 cubic yard class with design features quite different from today's. These larger loaders are being designed at the demand of the mining industry and the design criteria are being directed at this market area. The costs of these larger loaders will be greater per unit of capacity, and it is the intent of the various manufacturers that they will be more reliable, be easier to service, give greater availability, live longer, and cost less to maintain, than previous loaders. It is also the intent of the manufacturers to compete directly with the large mining shovel as the front end loaders become appropriately larger and more reliable. There has been some trend in this direction in the past with smaller shovels and loaders. In Figure No. 1 are data for U.S. production of front end loaders (5-6 cubic yard class and 10 cubic yard class) and U.S. production of cable operation, crawler mounted power shovels, draglines and cranes (1; to 5 cubic yard class) for the years 1965 and 1970. (Source: Bureau of Census, U.S. Department of Commerce.)

Jan 1, 1972

By Adrian Smith, Debra W. Struhsacker

13.1 INTRODUCTION Cyanide is, a generic term indicating the presence of the cyanide ion (CN-). Cyanide is a very common, naturally occurring compound produced by many biochemical reactions. Many plant species synthesize organic compounds which contain cyanide in the form of cyanogenic glycosides (Knowles, 1976). For example, trace amounts of cyanide are present in the seeds and leaves of many members of the Rosaceae (rose) family (Kingsbury, 1964). Many common items such as lettuce, maize, sweet potatoes, kidney beans, (Oke, 1969) almonds, and cigarette smoke contain cyanide. As examples, a recent chemical analysis of chocolate-covered almonds measured a total cyanide content of about two ppm (Steffen Robertson and Kirsten, 1987); and the U.S. Surgeon General (U.S. Department of Health, 1964) has shown that cigarette smoke contains up to 1,600 ppm total cyanide. The chemistry of cyanide solutions is complicated because the cyanide ion forms compounds and complexes with many elements. Some cyanide species are highly toxic whereas others are relatively inert and harmless. Molecular hydrogen cyanide (HCN) is the most toxic form of cyanide. Under most conditions, HCN exists as a gas which readily dissipates or reacts with the environment to form less toxic or nontoxic forms of cyanide. Thus HCN is an ephemeral toxin, and many naturally occurring geochemical processes reduce the HCN concentration of a heap system with time. As discussed below, free cyanide includes the two species, ionic cyanide (CN-) and molecular hydrogen cyanide. Free-cyanide toxicity in man, mammal s, and aquatic species is we1 1 documented (Doudoroff, 1976; Ecological Analysts, 1979; Towill et al., 1978). The lethal doses reported for human adults vary with the type of exposure as follows: • One to three mg/kg body weight if ingested; • One hundred to 300 ppm if inhaled; and • One hundred mg/kg of body weight if absorbed. Acute toxicity of free cyanide to freshwater invertebrates ranges from 0.028 to 2.295 mg/l, depending on species and test conditions. Generally, free-cyanide concentrations greater than 0.1 mg/l are expected to kill sensitive species in freshwater or marine environments (Doudoroff, 1976;

Jan 1, 1988