Search Documents

By Keith S. Olsen

Any discussion of barite mining is virtually impossible without considering, almost in the same breath, the many other variables, such as beneficiation, transportation, infrastructure and location that impact the economics of a particular barite orebody. My comments will be limited to just the mining aspects of barite orebodies and these observations must be considered in proper context with all other parameters affecting exploitation of a barite property prior to final decision on economic feasibility. In many instances low-cost mining situations are more than offset by high-cost beneficiation requirements. Likewise, there are cases of relatively high-cost mining associated with no or low-cost beneficiation requirements thus permitting economic exploitation. Barite orebodies and occurrences are almost as varied as the number of occurrences. As in most mining situations, the design of a mine plan and implementation of a proper exploitation program must be approached on a case by case basis. Proper and sufficient geologic control of a barite orebody must precede the development of a mine plan. Generally speaking, barite orebodies can and, at many times, do "strange" things; pinch with depth; almost never increase in size with depth, overturn, fault-off, change grade, etc. Due to the extreme variable nature of most barite orebodies, a mine plan and its implementation must be as flexible as possible. The ability to quickly adapt to variable conditions will greatly enhance the economics within a particular barite mining environment.

Jan 1, 1980

By William E. Mead

Although this paper has been entitled "Design of Embankments for Tailings Ponds and Waste Dumps," a better title might have been "Site Selection and Design Planning for Tailings Ponds and Waste Dumps," The procedures normally followed in performing the engineering calculations based upon the measured soil properties will not be detailed. Instead, we shall look at the subject in broader perspective which considers the environmental factors and the fundamental design principles. Disposal of waste materials left from the extraction and processing of mineral deposits has become, in recent years, an important engineering consideration for mine and mill operators. Land occupancy and land values almost everywhere are increasing at a rapid rate. So is the size of mine and mill waste dumps, thanks to our advancing technology in mineral beneficiation. Mining companies are being pressured to "clean up" their operations, both literally and in a safety hazard sense. Once, long ago, the miner could use any convenient location to get rid of his mine waste and mill tailings without being concerned over possible infringement on the rights and well-being of the public. Today, however, he must pay close attention to these matters if he wishes to avoid liability suits and the unfavorable reaction of civic groups as well as local and regional government. Most operators keenly sense their increased responsibility, not only because they are legally vulnerable, but because they realize the necessity for sharing space harmoniously with the rest of society.

Jan 1, 1967

By Graham Jones

During 1982 Ash Grove Cement Company contracted with Wagester, Walker, Thornton & Company to convert a Calcimatic rotating hearth kiln at its Portland, Oregon lime plant from gas/oil to coal firing. Increasing costs of natural gas and No. 6 fuel oil made this conversion mandatory if the plant was to remain competitive. Unique features of this project are 1) the distributors of pulverized coal to eighteen small burners around the inner and outer perimeters of the doughnut-shaped kiln, and 2) the attrition-dryer-pulverizer mill. Ash Grove, a century-old lime and cement producer headquartered in Overland Park, Kansas (Kansas City area), looked at various schemes for multiple burner firing before employing Wagester, Walker, Thornton & Company. W.W.T. of Pitts- burgh, Pennsylvania was formed in 1980 by acquiring the Solid Fuels Division of the Pullman Swindell Corporation. This division had been active in developing solid fuel firing for brick tunnel kilns, principally in the Southeast. They continue to supply such systems firing wood waste as well as coal to the brick industry and have also converted two Calcimatic lime kilns for Dixie Lime, Sumterville, Florida, a fluid bed calciner for Texas Gulf in Aurora, North Carolina, and a chrome concentrate kiln for Diamond Chemicals at Castle Hayne, North Carolina. Testing and a proposal have been made for INMETCO, Elwood City, Pennsylvania, on an annular hearth furnace for direct reduction of iron ore to metallic iron.

Jan 1, 1984

By W. E. Adkins

Ore and waste haulage usually constitute one of the major cost items in the operation of an open pit mine. This has always been true at the Chino mine, one of Kennecott Copper Corporation's western mines, located in southwestern New Mexico. The Chino Mines Division is an integrated operation. The open pit mine, dump leaching operation, and a precipitating plant are located at Santa Rita, New Mexico, and a power plant, mill, smelter, and fire refinery are located in Hurley, New Mexico, about nine miles south of the mine. Approximately 22,000 tons of ore are mined and processed each day; in addition, approximately 42,000 tons of waste are removed from the pit. Originally the pit was located along the side of Santa Rita Creek but it has deepened and expanded so that at the present time, it could better be described as a nearly circular pit about one to 1.5 miles in diameter. The bottom is at an elevation of 5750 feet, the west rim near the ore gathering yard is at an elevation of 6230 feet, and the highest bench on the east side of the pit is at an elevation of 6750 feet, giving a vertical mining depth of 1000 feet. The train waste dumps reach a maximum elevation of 6580 feet. At the present time, Chino uses electric locomotives, off-the- highway trucks, and a skip system for mine haulage.

Jan 1, 1963

By P. Mousset-Jones

The USBH Mining Systems and Ground Control Generic Research Center funded a project leading to the improved planning of airflow and refrigeration in underground mines experiencing hot working conditions. The project was broken down into approximately five phases: 1. Visiting some of the known hot mines in the U.S. and sending out a questionnaire to determine other mines which might be experiencing hot working conditions in the future. 2. Choosing two mines in which to conduct measurements to determine the insitu rock thermal conductivity and diffusivity, and comparing these values with laboratory determined values. 3. Investigating computer programs available for mine climate simulation, and testing these programs with case study examples from the two mines where the rock thermal properties were measured. 4. Considering improved ways of estimating the wetness factor and experimenting with the use of heat flux meters to more directly measure the rock heat inflow into the airstream. 5. Applying the results of these investigations to simulating the climate for future mine development and production areas at the two- mines. The project is approximately two thirds completed. At the time of writing, the insitu measurements at one of the mines are still being recorded. This paper will be describing the procedures and results of the insitu rock thermal property measurements, and give a brief description of some of the simulated results obtained for a mechanised cut and fill stope case study from the Homestake Gold Hine in South Dakota.

Jan 1, 1985

By L. Alan Weakly

Global warming is a political reality in the United States today. The actual impact of the continued utilization of fossil fuels is not known and is open to debate. However, the growing concern about the continued use of these fuels is problematic. There are technologies available to us today that would permit America to meet its future energy demand and not contribute to global warming. The inconvenient reality is that fossil fuels cannot completely be replaced with energy conservation and the use of wind, solar, hydroelectric, geothermal and nuclear energy. The connection between the greenhouse effect and global climate change was first identified by physicist John Tyndall in the 1860s. To prove the greenhouse hypothesis, Roger Revelle and his research team started measuring carbon dioxide concentrations in the earth?s atmosphere in the early 1960s (Becker, 2007). With the rising concern about global warming, scientists throughout the world have been obtaining surface and air temperature data. Scientists have also been comparing CO2 concentrations found in glaciers throughout the world with theorized annual temperatures. Data for both dates back 1,000 years. Armed with this data, many nations of the world and the United Nations Intergovernmental Panel on Climate Change (IPCC) have concluded that climate change is a reality. The IPCC also concluded that this climate change is the result of mankind?s consumption of fossil fuels and that catastrophic consequences will occur on a

Jan 1, 2008

By E. L. Smith

The development of larger size flotation machines, which started in the 1950's and 1960's, is still proceeding. The largest mechanical flotation machines avail-able today have unit volumes in the range of 38.2 to 43.9 M3 (1,350 to 1,550 cft). Designs for mechanical flotation cells of unit volumes as large as 70.8 m3 (2500 cft) are available. The development of these "giant" flotation machines has been and continues to be motivated by the increasingly critical forces of higher energy and capital costs, decreasing ore grades, and lower (real) prices for the products. As an expansion of a previous publication, this paper reports the further development of an improved mechanism for these large flotation machines. The starting point in the develop¬ment of a new mechanism was the identification of design and performance features of existing machines where improvements were desirable. The result of this work was a mechanism which s is an improvement over existing designs. This new design is characterized by an efficient overhung stator and a vortex profile rotor. The overhung design of the stator removes unnecessary and undesirable baffling in the lowest parts of the tank, while maintaining a quiescent froth layer. This results in lower power consumption and better pulp circulation and solids suspension. The vortex profile rotor acts as an efficient pump, with higher pulp recirculation rates as evidenced by the increased air dispersion capabilities, good solids suspension, and low power consumption.

Jan 1, 1983

By Stephen J. Swan

Feed-forward control (i.e. the forward flow of information in a control system ) is a simple and flexible approach to over-all control of a ball mill-rake classifier grinding circuit. The system described here uses mill power, feed rate, mill dilution water flow rate, and classifier dilution water flow rate only, to control mill load, throughput, mill solids, and classifier solids as well as classifier overflow grind. The mill load is controlled by a simple feed-back control loop, while mill solids and overflow grind are controlled within close limits by a unique non-linear feed-forward water control system. In effect the control follows a limited mathematical model of the grinding circuit. The configuration minimizes deviations in overflow grind due to changing ore conditions, but since it is open-loop, and the model is not complete, it is affected by changes in total ball charge and liner wear. These effects are fortunately gradual, and changes in the grind can usually be kept to within ± 1% of a target continuously for several days without adjustment. Control adjustments are made with feed-back from laboratory composites. This control system has won enthusiastic operator acceptance; it is easily started up and shut down; grind-outs have been eliminated along with flotation machine flooding and mill spills. Standard measurements and components are used throughout providing a trouble-free system design at a moderate cost.

Jan 1, 1970

By C. N. Bailey

Let's review the economic factors in mill design. Talking about engineering economics in mill design reminds me of an old' saying "There are several ways to handle women but nobody knows what they are." I'm sure that the same economic factors do not apply to each of our mill designs, but for the sake of emphasis, it is well to review a few of them. They are so simple that we are apt to feel they are not important. Other speakers today are talking about the engineering factors of mill design, but I'm asking the question "Have we checked the economic factors involved in these designs? Have we checked the obvious? Many great discoveries have resulted from analysis of the things that appeared to be obvious all the time. What are some of these economic factors? Perhaps the most basic one is simply the value of minerals in place if minerals have value in the ground, they have value only because many important and different considerations add up to the probability that the minerals can be mined, beneficiated and disposed of profitably. The size formation and quality of the ore body, the distance from market, and the transportation available to the market area, depreciation rates, taxes, and the ownership interests involved are but some of the factors that must dovetail into the answer of profit, an answer prerequisite to hoped-for financing.

Jan 1, 1961

By R. B. Pedersen, A. Filipa, A. Marques

"Shallow-depth (<200m) high temperature submarine hydrothermal systems are not commonly observed on mid-ocean ridge settings, as most known systems are located within arc - island-arc volcanic settings, often proximal to coastal areas (meteoric waters added). In 2013, during the Centre for Geobiology’s 2013 oceanographic cruise (R/V G.O. Sars), a dredge successfully sampled massive sulfides and hydrothermally altered rocks from the top of a volcanic edifice (~144-284 m depth) located on the northern sector of the Kolbeinsey Ridge. Dredged samples include vesicular altered basalts, and matrix supported anhydrite, barite and barite-pyrite-anhydrite volcaniclastic breccias. Sulfidesulfate breccias depict complex zoning where numerous small vent orifices are distributed with large blades of marcasite that appear intergrown with barite. Native sulfur was observed in one of these small vent orifices. On a microscopic scale, numerous spherical silica-rich structures appear associated to these breccias. These rocks are evidence for the presence of a relatively high temperature (minimum 150ºC) shallow-water seafloor hydrothermal system in the Arctic mid-ocean ridge, far from mainland influence. This new discovery presents a unique opportunity to study the dynamics of fluid chemistry, metal mobility and precipitation in shallow hydrothermal systems. The remote location of this system dictates that fluid recharge will be influenced solely by seawater and magmatic fluids, In July-August 2014 we expect to survey this new field and acquire new and comprehensive data."

Sep 1, 2014

By R. P. Stahura

Conveyors are the lifeline of mining and processing operations. They continuously supply materials to and carry finished products and waste away from operations. When conveyors are running at less than maximum efficiency, the operation&apos;s productivity is in trouble. Inefficient conveyors suffer carryback, leading to the buildup of materials on idlers and other expensive components. They spill and leak material at loading zones and conveyor transfer points. This causes accumulations that can damage the belt, wear moving components and increase both clean-up labor and maintenance expenses. They wander or mistrack, spilling material and endangering personnel and conveyor components. And through these inefficiencies, they multiply the operation&apos;s hidden costs. These added expenses occur in lost production, increased downtime for service and maintenance and replacement costs for prematurely worn equip¬ment. In addition, the attitude that inefficiency is "good enough" is recognized elsewhere in the organization. As a result, other departments relax or operate at less than peak performance. Quality goes down along with employee morale. Costs of inefficient conveyors A key to efficiency is handling a product as few times as is necessary. That means conveyors should be loaded, track true to the unloading point and then be unloaded, all without material escape or waste motion. An inefficient conveyor is often signaled by accumulations of DURT - material that has escaped from the system and is now at large in the plant. This DURT adds costs to an operation that are uncounted and, perhaps, uncountable.

Jan 1, 1992

By E. Snyders

Two alternative processes for the manufacturing of zircon products were investigated by the South African Nuclear Energy Corporation (Necsa). These were the synthesis of doped-zircon pigments from plasma dissociated zircon, and an upgrading process to convert an inferior grade of zircon into a superior opacifier for ceramic applications. A solid-state reaction process was used to produce doped-zircon pigments. The influence on the colour intensity of the pigments on a ceramic tile were compared for milled and un-milled plasma dissociated zircon (PDZ) material, prior to calcination. Pr-yellow and V-blue zircon pigments produced from PDZ with different degrees of dissociation, were also evaluated. The second zircon beneficiation process entails the upgrading of an inferior grade of zircon to a superior opacifier for ceramic applications. A calcining process in the presence of mineralizers was used to produce the product in a conventional muffle furnace. A pre-milled inferior grade zircon concentrate, with a relatively high iron and alumina content generally not suitable for high quality ceramic applications, was used in this study. The influence of the upgrading process on the opacifying properties of the zircon material was compared to that of commercially available prime grade zircon opacifier products currently being produced in South Africa. The influence of a calcining step in the presence of mineralizers with regard to the milling characteristics of the treated zircon material was also evaluated.

Jan 1, 2005

By John C. Wangaard

A method of raising by use of a cage has been used with considerable success at the Cary Mine, Hurley, Wisconsin, on the Gogebic Range. The Cary Mine is owned by the Odanah Iron Company, Pickands, Mather and Company, Managers. This method is known to have been used previously at Tennessee Copper and probably in other districts. The Cary Mine was engaged in deepening the main shaft from the 34th to the 37th level. To accomplish this, and to facilitate development, a ventilation shaft 930 feet north of the main shaft was sunk from the 34th to the 37th level and crosscuts were driven on the 35th, 36th and 37th levels to the main shaft location, These levels are at 200 foot verticalintervals. All of this work was done in granite and the majority of the workings were not timbered. The next step was to get a raise through from the 37th to the 36th level, and from the 36th to the 35th level, which would later be stripped to full shaft size. Several methods of raising were considered including double-compartment cribbed and two single raises interconnected at regular intervals, but the cage method was chosen as it appeared to be a safe and efficient operation. The method in general consists in the use of a drill hole between levels through which to pass a hoisting rope to raise and lower a light cage.

Jan 1, 1958

By Wallace W. Wilson

It scarcely 1s possible to read a new issue of any of the principal mining trade journals without noting some mention of a major new overseas mining venture with which one or more domestic companies are involved. These projects seem to follow a fairly consistant pattern; most are large open pit mines, with relatively low grade ore bodies but huge reserves; the total price tag will be in nine significant figures and around half or more of the total cost will be for infrastructure. West1 noted that nearly $7 billion is being invested annually for new mining projects and the expansion of existing mining ventures, in the free world. He anticipates that by the year 1975 such investments would reach the $10-11 billion annual figure, and that in the five year period 1972-76 total capital requirements can be expected to aggregate around $45 billion. The proportion of this total investment which will be financed in some manner is difficult to estimate because of the variety of available financing arrangements. West estimated that around 75% of the total capital would come from internally generated funds, with the balance from external sources. Conceivably a substantially higher percentage could be attributable to external financing, especially it the trend of recent mineral development in Canada and Australia is any criterion, and of course, including local sales of equity as a part of the package.

Jan 1, 1973

By M. E. Smith

Waste rock dumps, tailings impoundments and spent leach heaps are common features at mine sites. Many of these waste disposal facilities contain sulfide-rich minerals, which generate acid mine drainage (AMD) when they oxidize. Control of AMD can be achieved by removing one or more of the three essential components in the acid-generating process (sulfides, air, or water). One AMD prevention measure is to cap the waste with an engineered cover system to minimize the infiltration of water and diffusion of air into the system. Possible cover designs include: ? Earthen ?store-and-release? covers (also known as evapotranspiration (ET) or water balance covers), including capillary barriers. These covers require climate conditions where there is both modest to little rainfall and significantly greater evapotranspiration than precipitation, both annually and seasonally. ? Low-permeability compacted clay liners (CCLs) employing fine-grained soils typically compacted at wet of optimum water content. ? Geosynthetic composite covers, comprised of a geomembrane liner placed over and in direct contact with either a CCL or a geosynthetic clay liner (GCL). These systems have been commonly used for base liners for leach pads as well as covers and base liners at municipal solid waste (MSW) landfills in the US for more than 20 years. Comparisons of these different cover systems, both in terms of performance and cost, are presented later in this paper. First, however, a review of the typical mine closure approaches, and the possible rationale behind them, is warranted.

Feb 27, 2013

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada&apos;s Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By D. J. Boron, R. Kollrack

Introduction The US has the largest total (364 Gt, or 401 billion st) and recoverable (182 Gt, or 201 billion st) coal reserves in the world (Steam/Its Generation and Use, 1978). However, progressive coal use in this country has been hampered by environmental constraints, such as restrictions on sulfur dioxide, nitrogen oxides, and particulate emissions. Since the promulgation of the Environmental Protection Agency (EPA) New Source Performance Standards (NSPS), it has become apparent that all US coals will require sulfur removal to comply with the new sulfur dioxide emission levels. If more stringent emission standards are imposed, through pending acid rain legislation or amendments to the Clean Air Act, it is likely that the coal industry, and the industrial and utility sectors will look increasingly toward the development of new and advanced coal cleaning technologies to continue and promote the use of coal. Table 1 summarizes the EPA NSPS for sulfur dioxide emissions. Fortunately, the basis for determining sulfur dioxide removal is the Btu and total sulfur content of the run-of-mine coal. Thus, credit is given for any sulfur removed before combustion, e.g., through physical coal cleaning. Unfortunately, physical coal cleaning alone may be unable to produce a compliance coal with respect to the EPA NSPS. Therefore, all industrial and utility coal-fired boilers will require post combustion gas cleanup, e.g., flue gas desulfurization (FGD), to comply with the new regulations. Table 1 - EPA NSPS for SO2 Emission Category SO2 Content of ROM Coal SO2 Reduction Required 1 0-2 Ib/M Btu 70 2 2-6 Ib/M Btu To 0.6 Ib/M Btu (70% to 90%) 3 6-12 Ib/M Btu 90% 4 12 Ib/M Btu To 1.2 Ib/M Btu Alternatives to post combustion gas cleanup, such as staged combustion, lime and coal injection, fluidized-bed combustion, and chemical coal cleaning, are currently under development. Of these, chemical coal cleaning offers the advantages of providing (1) high quality fuel for combustion (low sulfur and ash, high Btu, and attenuated product variability); (2) a solid fuel product for use in retrofitted gas and oil boilers; and (3) a solid fuel product with multiple end uses, e.g., internal combustion engines, carbon electrodes, coke, or chemical feedstock. Gravimelt process In the Gravimelt process, 1.4 mm x 0 (14 mesh x 0), physically beneficiated coal of about 5% moisture is fed into a 3401 to 3901 C (644 ° to 734 ° F) molten bath of sodium hydroxide and potassium hydroxide. Organic and pyritic sulfur and most of the mineral matter in coal react with and are dissolved in a molten mixture of potassium hydroxide and sodium hydroxide. The reacted coal sulfur is converted to alkali sulfides and polysulfides, whereas the mineral matter is apparently converted to alkali-aluminum silicates and other byproducts that are water and acid soluble. Coal residence time is 60 to 240 minutes. The sodium hydroxide to potassium hydroxide ratio is about 9:1. The total hydroxide-to-coal ratio is no less than 10:1. The specific gravity of the caustic melt is about 1.8 g/cm3 . Thus, coal is buoyant in the bath, and insoluble mineral matter with density greater than 1.8 g/cm3 would sink. Reacted coal is first water-washed with an equal weight volume of water in a countercurrent mode of operation. The coal is filtered and the filtrate sent for regeneration. The coal is then acid-washed in 10% sulfuric acid. The acid-water/coal mixture is filtered to recover the final coal product, and the filtrate treated with lime to precipitate dissolved mineral matter. The precipitate is filtered from the mixture and discarded. The water is recycled to the plant. Filtrate from water-washing is treated to remove sodium sulfide. The sodium sulfide is chemically treated to produce hydrogen sulfide, calcium car¬bonate, sodium hydroxide, and other products. The hydrogen sulfide is treated to produce sulfuric acid for the acid-washing step. The calcium carbonate is treated further or discarded, and the sodium hydroxide is concentrated by evaporation and recycled to the reactor. The Gravimelt process is currently at the bench-scale stage of development. A 0.9-kg (2-lb) continuous reactor is in operation with a continuous water- and acid-washing section adjacent. The regeneration section is not developed to a meaningful extent, although key reactions have been tested in the laboratory. Process conditions, product recovery efficiencies, etc., however, have not been determined adequately and continuous regeneration operation has not been attempted. Work is now in progress to develop each section to a continuous bench-scale operation in processing coal at 9.1 kg/h (20 lb per hr). Microwave process The carbon and hydrogen in coal are relatively transparent to microwave radiation. However, water, caustic, pyrite, and other mineral matter components are strong conductors of microwave radiation energy at the proper frequency. As a result, selective heating of the noncarbonaceous materials occurs that enables the chemical reaction of sulfur and ash with caustic. Although very high temperatures may result from

Jan 1, 1987

By Lok W. Home

INTRODUCTION Today, raise drilling (boring) has become the stan¬dard method of raising throughout the Western world. It is estimated that the majority of raise footage com¬pleted in 1977 was accomplished by the raise drilling method. Extensive replacement of the drill and shoot system by mechanized raise drilling has occurred be¬cause users have recognized one or more of the follow¬ing advantages. In a direct cost per meter comparison, mechanized raise drilling is economically more attractive than drilling and shooting. Mechanized raise drilling offers improved personnel safety over the drilling and shooting method. Mechanized raise drilling is a faster and more predictable method for excavation than is drilling and shooting. The raise cost per unit of length becomes less as the raise length increases, offering greater flexibility in the overall design of mines. Mechanized raise drilling provides a substantial reduction in labor requirements. 6) Mechanized raise drilling improves the working environment. 7) Mechanized raise drilling improves the rock in¬tegrity of round, unblasted raises. Increasing acceptance of the raise drilling technique is enhanced by continuing equipment development. Today, a wide range of machines and capabilities is available, machine drive systems have been improved, reliability has been increased, and cutter technology is being improved quite rapidly. Table 1 lists the four major manufacturers of raise drilling machines, cutters, and reamers, together with the approximate number of machines that each company had in operation at the end of 1977. Although Reed and Smith are major suppliers of reamers and cutters, they do not manufacture the drilling machines. Four or five other relatively minor suppliers have built machines, reamers, and cutters for raise drilling, but they did not have a substantial num¬ber of machines in the field at the end of 1977. MACHINES FOR MECHANIZED RAISE DRILLING There are three principal arrangements for making machine-bored raises. They include the use of standard raise drills, reversible raise drills, and blind-hole or box¬hole raise drills. Standard Raise Drills The standard raise drilling machine is set up on one mine level or on the surface, and a pilot hole is drilled from that level or surface to a level below. When the pilot hole has been drilled, the raise then is reamed back from the lower level to the upper level or the surface. This arrangement is illustrated in Fig. 1. Reversible Raise Drills The reversible raise drilling machine is set up on a lower level of the mine, and a pilot hole is drilled up¬ward to a higher level. The raise then is reamed back from the upper level to the lower level. This arrange¬ment is illustrated in Fig. 2. The machines used for this type of raise drilling normally are capable of being used as standard raise drills. Blind-Hole or Boxhole Raise Drills The blind-hole or boxhole raise drilling machine is set up on a lower level of the mine, and a full diameter raise is bored to a higher level without the use of a pilot hole. This type of machine is illustrated in Fig. 3. Drill Recommendations If access to upper and lower mine levels is equal, the standard raise drill should be used. The complexity of the equipment and associated handling procedures for boxhole drilling far outweigh any mechanical advantage or reduction in the number of operations, provided that access to an upper level is available. If there is good access to a lower level but only limited access to an upper level, the use of a reversible raise drill should be considered. If there is no access to an upper level, a blind-hole or boxhole raise drill is the only practical mechanized system. MACHINE SELECTION There are six major considerations in selecting a machine for raise drilling. These include operating range, cutter loading, machine structure and main¬tenance, drill string configuration, drive system, and automated controls. Operating Range As shown in Table 2 and Figs. 4 through 6, raise drilling machines are available for a wide range of hole sizes. Standard raise drill ratings range from raises of 0.9 m (3 ft) diam and 120 m (400 ft) length up to raises of 3.5 m (12 ft) diam and 900 m (3000 ft) length. Boxhole raise drilling machines are available for raises ranging from 0.9 m (3 ft) to 2 m (6 ft) diam with lengths up to 90 m (300 ft). Table 2 lists, among other criteria, the torque and thrust of typical machines; torque and thrust are the main criteria for analyzing the capacity of a machine. The raise sizes shown in Table 2 assume a condition of medium hard to hard rock formations and optimum cutter loading. The recommended raise size for a given machine can be exceeded when raising softer rock formations or when the operator is willing to accept a lower level of performance; the performance is related directly to the 1093

Jan 1, 1982

By Daniel F. Meyer

INTRODUCTION The Metropolitan Boston Transit Authority (MIA) Red Line has served greater Boston commuter needs from its Harvard Square terminus since its opening in 1912. Since then, the population of Boston&apos;s northwest area has increased substantially, and with this growth have come dramatic changes in the type and distribution of urban land uses, including the number of households, jobs, shopping and recreational opportunities; these changes have created a need for improved transportation. Extension of the Red Line to better serve these communities has been a topic of discussion for nearly 40 years. In November of 1977, the Urban Mass Transportation Administration (UMTA) granted the MBTA approval and funding for the Red Line Extension Project. With funding in place, the MBTA proceeded with final engineering and construction of the section from Harvard Square to Alewife. When completed, the Red Line Extension will add 3.2 miles of new track to the MBTA system. Estimated final cost is about $400 million. On October 3, 1978, the MBTA received bids for the construction of Project 091-105, Harvard Square to Porter Square Tunnels and Shafts, which is part of the overall Red Line Extension. The Joint Venture of Morrison-Knudsen, White, Mergentime was the low bidder at $47,478,600. Official award of the contract came on October 25, 1978, with Notice to Proceed given October 31, 1978. The contract called for construction of two independent parallel concrete-lined tunnels in rock, soft-ground and mixed-face

Jan 1, 1983

By V. Bush, T. G. Killingsworth, E. P. Ruffel

Effective and sustained corporate energy management programs are becoming a necessity for energy users as electricity deregulation laws are enacted on a state-by-state basis across the country and environmental regulations pertaining to emissions become more prevalent. The “Total Energy Asset Management” (TEAM) approach has been developed to assist energy end users with the development and implementation of corporate energy management programs designed to reduce energy risks and increase energy cost savings. This paper highlights the ways in which this approach can be applied to the mining industry while minimizing initial program implementation costs and increasing potential program benefits. Nearly the entire mining industry has become very focused on identifying opportunities to reduce costs per unit of yield. The industry as a whole has increasingly faced pricing pressures in large part due to the effects of global competition. The result of international marketplace participation and the emergence of higher efficiency recovery techniques have resulted in flat or even decreased prices for mining products. Thus, opportunities to reduce costs are actually opportunities to gain a competitive edge over similar mining operations. Due to the energy-intensive nature of many mining operations, the costs associated with energy are often a substantial percentage of operating costs. Therefore, it becomes very important to develop the corporate mentality that energy is an asset whose use should be optimized globally. In addition to the cost benefits of the TEAM approach, other benefits may occur such as reductions in facility greenhouse gas emissions.

Jan 1, 2002