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By Curtis A. Pledger, H. Lee Wimmer, I. Michael Gowring

The Upper Diamond Fork Project, as part of the Central Utah Project Completion Act, conveys water six miles from the existing Sixth Water Aqueduct to the Diamond Fork Pipeline. During design, a Value Engineering team proposed a single-tunnel alignment that would be less destructive on the environment and might have a lower life cycle cost than the proposed design of two tunnels connected by two pipelines on steep hillsides. A decision was made to include this option in the bid process as a design-build alternative. The bid for this option was the lowest and a contract was awarded. During the TBM excavation a segment of tunnel wall caved in containing large quantities of water saturated with extremely high levels of hydrogen sul?de (H2S)gas (measured at 380 to 520 ppm in the air). The Central Utah Water Conservancy District (District) in consultation with Obayashi/Clyde (Contractor) quickly developed anew plan to finish the system by incorporating parts of the original tunnel/pipeline concept and re-bid the remainder of the project as a Phase 2 Design-Build contract.

Jan 1, 2003

By Biao Qiu

Longwall and/or room-and-pillar mining operations in multiple coal seams could not only induce subsurface and surface subsidence but also cause interactions between these mined coal seams. The interaction might destabilize mine structures and subsequently induce additional strata movements. Mechanical models to utilize the predicted subsurface deformations to assess the interactions and the consequences have been developed. The models are incorporated into a computer program, CISPM-MS, for predicting the final surface movements and deformations as well as the mining interactions associated with multi-seam coal mining operations. It predicts final surface movements and deformations caused by the individual mining operations and by the interactive effects. The program can also be used to assess mine structural stability in multi-seam mining operations. The paper presents the developed interaction models and their application in surface subsidence prediction caused by multi-seam mining operations. A case study involving mining in two coal seams, using longwall and room-and-pillar mining methods, respectively, will be used to validate the program.

Jan 1, 2013

By Yi Luo

Bolting the mine roof is one of the noisiest underground mining operations. The main contribution to the high noise is the drilling of the roof bolt holes, especially in hard rock. A previous research shows that the specific energy consumed in drilling the bolt holes is inversely proportional to the bite depth (i.e., penetration per drill revolution). Less specific energy at a reasonably high bite depth means less energy is wasted in the drilling process to produce heat, bit wear and noise. Therefore, control of drilling operation to achieve a higher bite depth can provide an opportunity to reduce drilling noise from its sources as a proactive noise control approach. In order to explore the feasibility and to develop the proactive technology for reducing drilling noise, drilling tests have been conducted in laboratory settings. The preliminary results show significant reduction in noise dose can be achieved by drilling in medium hard rock at a reasonably high bite depth. This paper presents the findings from the laboratory experimental studies conducted so far.

Jan 1, 2013

By Frederick K. A1lgaier

The Bureau of Mines' Denver Research Center initiated a subsidence research program in 1978. The short ten goals of the program were to determine practical methods and equipment for measuring subsidence and demonstrate these methods by monitoring subsidence at several mines having mining, geologic and topographic variables typical of those encountered in the western United States. Data from the monitoring program would be supplied to cooperating mining companies and used by the Bureau to accomplish the long term objective of developing and validating subsidence prediction techniques applicable to mines in the West. Subsidence monitoring is being conducted at five mines in Colorado and Utah with substantially different mine plans, mining methods, depths of cover, and topography. The monitoring procedures used at each site are similar and include designing the monitoring network layout, installing the subsidence monuments, initial and periodic surveying of the monuments, and processing the survey data. The network layout, surveying methods, and types of monuments vary according to the conditions at the individual sites and the type of data being collected.

Jan 1, 1982

By Paul D. Tomlingson

Maintenance costs in the mining industry often account for as much as 18-20% of total operating costs and with the cost of labor and materials rising, they threaten to become an even larger percentage of the total. Yet, without an effective maintenance program costs can soar beyond what they would be without a program. In other words, when production equipment is down, there is no production and no product to sell in the marketplace. This cost of downtime in lost product far exceeds the actual cost of maintenance, which is defined here as labor and materials. Thus, mine management must have an effective maintenance program to avoid downtime and meet production targets. Attaining an Effective Maintenance Program Many mine management and operations supervisors believe that an effective program is a function of how hard they push maintenance. Yet, when vital production equipment is inoperable and production is lagging, no amount of pushing helps. In actuality, maintenance has little control over its own destiny and needs support from other departments to deliver good service. Success of the maintenance program depends on gaining control of those factors that directly influence maintenance. Consider the following. Organization-The maintenance organization must be structured for maximum responsiveness to production needs. Unfortunately, some needs are immediate and many are unknown in advance. The maintenance department has little control over them. These outside influences limit the department manager's ability to control his own organizational structure. Preventive maintenance procedures-Routine lubrication, equipment inspection, minor adjustments and the like are controlled mostly by maintenance. Yet, scheduling preventive maintenance depends also on the availability of equipment controlled by other departments. Planning-The purpose of planning maintenance jobs is to ensure that the best use is made of maintenance resources. If the production department creates an environment of running repairs and emergencies, there can be little planning. Even well-planned jobs will suffer if there is less than full cooperation in carrying them out. As a consequence, the maintenance department is not blocked from planning, but the value of its planning is reduced. Scheduling-Maintenance scheduling should be a joint effort. Although maintenance does the planning, production must agree to the scheduled maintenance of its equipment. Workload-As a service organization, the maintenance department has limited control over its workload because it must consistently respond to other department's demands. At best, maintenance controls only the fixed portion of its workload-utility work, janitorial service, maintenance shop work, and assigned repetitive and preventive maintenance duties. Unscheduled and emergency work are initiated by operations. Major maintenance actions require full cooperation. Labor productivity-The effective use of labor is tied to other key functions of maintenance. These include organization, preventive maintenance, planning, scheduling, and emergency work. An area type organization, for example, can often respond more quickly to immediate needs than a centrally controlled maintenance group. Yet, area maintenance can result in lower labor productivity unless tight control is exercised. One of the advantages of a good "detection-oriented" preventive maintenance program is early discovery of problems and better opportunity to schedule corrective work in advance. Under planned work conditions, use of labor can be more productive. Complete planning and scheduling can minimize time spent on each job, increase effectiveness of supervision, and improve labor productivity. Finally, if emergency work is minimized, there are fewer delays, less lost time, better coordination, and more productive use of labor. While labor productivity is an important factor in carrying out maintenance, it also is influenced by outside factors. However, few realize the impact of being able to improve maintenance productivity. By improving the productivity of 30 men from 30% to 40%, maintenance could accomplish the same work value as if it had added 7.5 men working at 40% productivity. 30 (1800) (40%) = 21,600 man-hr/yr 30 (1800) (30%) = 16,200 man-hr/yr Difference = 5,400 man-hr/yr 1800 (40%) 7.5 men Leadership-The maintenance department can have significant control over its own leaders, the front-line foremen, but they must

Jan 12, 1982

By E. J. Garrison

Discussion by E.J. Garrison A.C. Noble presents a clear and concise summary of the factors pertinent to a competent reserve evaluation of a mineral occurrence. However, I believe that he confuses the difference between a resource and a reserve. By definition, a reserve estimate includes all known economic, legal, mining, metallurgical and environmental factors impacting the recovery of the product(s) found in the deposit. A resource estimate, however, is an estimate of what is there - what is available for exploitation and should not include economic and other considerations. Only after defining the resource, its extent, location, physical and chemical properties, can an intelligent selection process begin of the extraction and processing technologies most appropriate to the circumstances encountered. The early and almost certainly inappropriate application of these often cookbook choices has probably lead many companies to walk away from good properties. (How many mines have been found in others rejects?) An example of the process is the early selection of metallurgical samples, after encouraging values are found in a few holes or surface samples, in an attempt to determine the economics of the prospect. Unfortunately, the sampler does not know how representative of the occurrence the sample is because he does not know what the deposit looks like. As a consequence, the property becomes saddled with a bad metallurgical sample that is often inappropriately used in evaluating the deposit which it does not represent. Reply by A.C. Noble E.J. Garrison asserts that I confuse the difference between a resource and an ore reserve. In fact, this paper was motivated by the observation that resources and reserves are almost universally confused throughout the industry. While Garrison is concerned that too-early consideration of engineering, metallurgical and economic factors will result in dropping properties too soon, some issues that must be considered include: •The purpose of exploration is to discover ore reserves, not to discover resources. •Ore reserves are always smaller than resources. • Decisions on exploration expenditures must be made based on the expectation of the quantity of ore reserves that will ultimately be defined. Noble's Fig. 2 can be used to give guidance as to when it is appropriate to begin reserve calculations. When exploration enters the "flat" portion of the curve, representative samples can be selected for metallurgical studies and appropriate mining methods chosen. It is at this stage that we know enough about the deposit to make intelligent choices of representative samples and recognize constraints on mining caused by the deposits shape, environment and physical properties. Once the appropriate technologies have been chosen, then their cost can be estimated with reasonable accuracy. Early entry of economic and process factors gives the false appearance of reserve status to the resource estimation. It would be more informative and less subject to abuse and misunderstanding if resource calculations were made at a number of cutoff grades. The cutoff grades should be chosen to cover the range of grades found in practice in existing operations of similar type. This would emphasize the open nature of resource estimates and encourage creative evaluation of mining and process technologies. It does not preclude the use of economic, environmental and recovery technology factors in deciding the advisability of continuing exploration expenditures, but would highlight their speculative nature. It is thus hoped that decisions on the future of the deposit would not be made on the presumed suitability of a given technology but on the actually known reality. I agree that resource estimates should be reported as a range of cutoff grades. Since reporting at a cutoff implies application of underlying economics (even by analogy to similar properties), however, it should be clearly stated that resources do not consider economic, mining or cost factors, and that ore reserves will be substantially less. The dilemma of the explorationist is that until the continuity and limits of ore-grade mineralization have been established, it is difficult to make quantitative estimates of either resources or ore reserves. Further industry discussion is needed regarding standards for disclosing early exploration information in such a way as to fairly represent the potential of an exploration property.

Jan 1, 1995

By Dennis S. Kostick

Introduction Soda ash is known chemically as sodium carbonate, an important inorganic chemical. It has been produced for several centuries by processing certain vegetation and minerals. The US soda ash industry has evolved from several small sodium carbonate mining operations in the West. Now, a nucleus of six companies produce about one-fourth of the world's annual soda ash output US producers currently dominate the world market. But certain international events are occurring that will reshape the domestic soda ash industry in the next decade. Historical perspective Soda ash is used mainly in the manufacture of glass, soap, dyes and pigments, textiles, and other chemical preparations. All of these are the first basic consumer products produced by developing societies. About 3500 BC, the Egyptians became the first society to use crude soda ash. The soda ash was used to make glass containers. It was most likely obtained from dried mineral incrustations around alkaline lakes. Soda deposits were virtually nonexistent in western Europe. So people resorted to burning seaweed to obtain the ashes. The ashes were then leached with hot water and the solute was recovered after evaporating the solution to dryness. The solute, a crude "soda ash" was impure. But, it could be used to make glass and soap. These two products and industries were important to the population and economic growth of the region. About 11.5 t (13 st) of seaweed ash was required to produce about 0.9 t (1 st) of soda ash. Along the coasts of England, France, and Spain, seaweeds with varying alkali contents became important items of commerce and sources of soda ash before the 18th century. The LeBlanc process used salt, sulfuric acid, coal, and limestone. It became the major method of production from about 1823 to 1885. In the early 1860s, Ernest and Alfred Solvay, two Belgian brothers, successfully commercialized an ammonia-soda process to synthesize soda ash. It used salt, coke, limestone, and ammonia. The Solvay process produced a better quality product than the LeBlanc method. In 1879, Oswald J. Heinrich presented to the Baltimore meeting of AIME, a paper entitled "The manufacture of soda by the ammonia process." The paper compared the two processes and foretold the demise of the LeBlanc technique. World production of soda ash in 1880 was 680 kt (750,000 st). Of that, 544 kt (600,000 st) was produced by the LeBlanc process. Of the 2.8 Mt (3.1 million st) of soda ash produced worldwide in 1913, only about 50 kt (55,000 st) was by the LeBlanc method. The LeBlanc process was never used successfully in the US, except for a brief period from July 1884 to January 1885 in Laramie, WY. Previously, soda ash had been produced by burning certain plants, as exemplified by the early Jamestown colonists, or by recovering small quantities of natural sodium carbonate found in alkaline lakes, such as those found near Fallon, NV, and Independence Rock, WY. Before the 1884 startup of the first synthetic soda ash plant in the US at Syracuse, NY, most of the domestic soda ash demand in the East was met by imports, primarily from England. Large-scale commercial production of natural soda ash began in California in 1887 from surface crystalline material at Owens Lake. Production from sodium carbonate-bearing brines at Searles Lake began in 1927 (Fig. 1). In 1938, during exploration for oil and gas in southwestern Wyoming, a massive buried trona deposit, presumably the world's largest, was accidentally discovered. Recent mineral resource evaluation by the US Geological Survey and the US Bureau of Mines indicates that the Wyoming trona deposit contains 86 Gt (93 billion st) of identified trona resource in beds 1.2 m (4 ft) thick or greater. Additionally, there is about 61 Gt (67 billion st) of reserve base trona. Of this 36 Gt (40 billion st) is in halite-free trona beds and 24 Gt (27 billion st) is in mixed trona and halite beds. In 1953, the Food Machinery and Chemical Corp. (later shortened to FMC Corp.) became the first company to mine trona in Wyoming. Soda ash demand increased.

Jan 10, 1985

By J. Tuzson, H. McL. Clark

Introduction Wear of pumps, wet cyclones and pipelines by suspended solid particles will impact their operating economies by limiting their useful life, reducing their reliability and increasing the amount of servicing. Erosion is considered a materials problem. The empirical solution consists of testing different materials in field applications. Such tests are time consuming and expensive. The prediction of localized wear rates, life estimates and appropriate material selection demand a better understanding of the slurry wear process in environments corresponding to those found in working equipment. Key words: Wear, Slurry abrasion, Coriolis erosion tester The slurry wear process Solid particles suspended in liquid generally move with the flow, but such particles deviate under the centrifugal and Coriolis forces encountered in pumps. The concentration of particles will increase, and a solid layer or bed may accumulate if the acceleration is directed against the flow passage wall. Large particles in dilute suspensions may impact the wall. dissipating their kinetic energies and removing material. However, particles smaller than about 11)(1 µm encounter great drag when moving through the liquid. and such par¬ticles have little kinetic energy left when they reach the wall. They cannot penetrate the stagnant boundary layers present in pumps. Therefore, small particles tend to form beds that slide along the wall, removing material by a process that could he called sliding erosion or wet abrasion. Wear by impact erosion from large particles in dilute slurries can he predicted by appropriate analytical and empirical methods. However, laboratory tests and calculations of wear rates that correspond to sliding erosion are more likely to correlate with field experience from wear under similar conditions. Specific energy Data from laboratory slurry erosion experiments have shown that the amount of material removed is approximately in proportion to the energy expended in the erosion process. Thus. the specific energy-energy expended in removing by erosion a unit volume of material-offers a simple, practical measure for the erosion resistance of materials. The concept was found to also apply to sliding erosion (Clark et al.. 1997). Generally, the study of the material-removal process in slurry erosion continues to remain a subject of research. Conceptually, if the flow velocities and accelerations could be calculated in a pump or pipeline, the energy dissipated at the passage walls could also be estimated, and the material removal rate would become known from the specific energy determined by laboratory tests. Slurry erosion testing Laboratory devices made specifically for slurry erosion testing may be grouped into the following: jet erosion devices, slurry pots and the Coriolis test machine (Pagalthivarthi and Helmly, 1992). Jet erosion devices use a slurry pump and nozzle to produce a high-velocity slurry jet that is directed at a flat material specimen, either perpendicularly or at some acute angle. Combined impact and sliding erosion are produced. but the processes cannot be separated, and the wear is difficult to assess except by weighing. Slurry pots consist of a cylindrical vessel containing the slurry of interest. In the slurry pot, test specimens are attached to a rotating arm. The samples are rotated in the slurry and experience local flow conditions that result in combined impact and sliding erosion. Wear depth measurements from detailed laboratory experiments have been successfully correlated with results from an analytical model of the process, and specific energy values have been calculated (Wong and Clark, 1995). In industrial use, the samples are weighed to determine the total material loss. Materials are ranked with respect to some reference material. The Coriolis erosion tester (Tuzson, 1984) consists of a rotor with a central cavity, from which two opposing radial passages lead out (Fig. 1). The slurry is fed to the centerof the rotor in one pass from a stirred overhead tank. The slurry particles are subjected to the Coriolis acceleration as they flow out through the radial passages. The material specimens are mounted into the radial passage walls facing the direction of rotation and are eroded by the particles that are pressed against them as they flow out radially. A narrow groove is worn into the sample within a few minutes, using a few gallons of slurry. Material removal is estimated accurately from profilometer traverses across the wear scar at different locations. The energy dissipated by the particles can be estimated, which leads to a well-defined experimental value for the specific energy.

Jan 1, 2000

By A. D. Davis, C. J. Webb, A. Heriba

Introduction The disposal of spent ore from cyanide heap-leach processing facilities is of concern to the mining industry, the regulatory agencies and the general public. The disposal of an additional several hundred million tons of spent ore from heap-leaching operations is planned in the western United States during the next decade, and this amount could increase if gold prices rise. In the past decade, more than fifteen million tons of spent ore have been disposed of in the Black Hills of South Dakota. The criterion for offloading spent ore was set by the South Dakota Department of Environment and Natural Resources at 0.5 ppm weak acid dissociable (WAD) cyanide. South Dakota's nitrate offloading limits vary for each mine, depending on ore type and other factors. Concentrations of nitrate in the leachate from spent ore depositories could be the result of conversion from other nitrogen-containing species. One likely source of nitrate is residual explosives from the blasting of ore. Previous work on the combustion products of nitrogen explosives has shown incomplete oxidation (Johansson and Persson, 1970). Another likely nitrate source is the degradation of cyanide to ammonia, followed by oxidation to nitrite and nitrate. A third possibility may be leaching of background nitrate from soil or rock. Objective The primary objective of this research was to investigate laboratory procedures for predicting nitrate concentrations in effluent after the disposal of spent ore at heap-leach gold processing operations. Laboratory tests were performed to define the concentrations of nitrate and nitrite in leachate as a function of time. Sources and species conversions were examined by investigating the nitrogen contributions from cyanide neutralization and from blasting with ammonium nitrate/fuel-oil (ANFO) explosives. Ore processing at cyanide heap-leach facilities Gold mines in the northern Black Hills of South Dakota normally use the following steps in their heap-leach cyanidation processing: Blasting. Surface gold mines in the Black Hills use ANFO explosives (approximately 95% ammonium nitrate and 5% fuel oil). Field evidence and studies show that explosives often fail to ignite or burn completely in shot holes. One reason is that ammonium nitrate absorbs moisture readily, which reduces blast efficiency. Ferguson and Leask (1988) cited a study conducted at the Fording coal mine in southeastern British Columbia. In that study, the total nitrogen released to surface and ground water was estimated to be about 6% of the slurry explosives. However, Ferguson and Leask (1988), in a later analysis of water quality near coal mines in the Kootenay coal fields, indicated that this is likely an overestimate of the nitrogen release. Mines that used ANFO explosives in dry conditions released 0.2% of the nitrogen from blasting. Cyanidation. At surface gold mines in the northern Black Hills, cyanide species from heap leaching can be classified into free cyanides and simple ions, weak complexes, moderately strong complexes and strong complexes (Scott and Ingles, 1981). Degradation of cyanide and disposal of spent ore. Methods of destroying cyanides include hydrogen peroxide oxidation, natural degradation and evaporation, water leaching and alkaline chlorination. Hydrogen peroxide is used to oxidize cyanides in the northern Black Hills mining area. The hydrogen peroxide oxidation reaction forms cyanate, which can hydrolyze to form ammonium and carbonate ions. After formation of ammonia or ammonium ion, further oxidation to nitrite (NO2-) and nitrate (NO3-) can occur. Spent tailings from gold processing in the Black Hills normally are disposed of in tailings facilities constructed in nearby drainages. Infiltrating rainwater with a typical pH of about 5.6 is an additional component introduced into the spent ore after disposal. Methods Field sample collection. Spent ore samples were collected from conveyor belts (or other convenient points) at heap-leach mines by personnel from the South Dakota Department of Environment and Natural Resources. The samples were placed in black plastic bags that were contained within 5-gal buckets and sealed to prevent exposure to ultraviolet light. The samples had a minimum head space to limit evaporation and contact

Jan 1, 1997

By G. T. Newell, J. G. Stone, V. M. Mejia

Introduction Advances in drilling have reached a point where large diameter cores can be recovered from "tight," or weakly indurated placer gravels. In such ground, core drilling can provide more reliable data regarding tenor than can be obtained using churn drilling or similar classical techniques. It can also provide metallurgical and geological information that is not available from samples obtained through alternate methods. In 1985, Coastal Mining Co, a subsidiary of M. A. Hanna, and Western Gold Reserves began to review a Tertiary placer deposit owned by San Juan Gold at North Columbia, CA, about 14 km (9 miles) northeast of Grass Valley. The deposit is one of the largest remaining unmined portions of the formerly extensive early Tertiary ancestral Yuba river system. It has been known since the 1850s, has been the subject of much technical literature, and has been the object of at least four previous drilling programs. The eastern one-third of the 6 km (3.7 mile) stretch of the channel between North Columbia and Badger Hill was partially stripped by large scale hydraulic mining in the late 1870s and early 1880s. Mining ceased in 1884 when the Sawyer Decision prohibited further discharge of hydraulic tailings into the Sacramento and San Joaquin Rivers. By that time, about 30 to 45 m (100 to 150 ft) of relatively low grade upper gravels had been removed over some 81 hm2 (200 acres). About 90 to 105 m (300 to 350 ft) of higher grade middle and lower gravels were left at least partially stripped. In 1914, a few churn holes were drilled along a widely-spaced line. In 1938-1939, Selection Trust conducted an extensive drilling campaign to evaluate the deposit. Particular attention was directed toward the partially stripped eastern portion. In 1968, the US Geological Survey drilled three churn holes in the eastern part of the deposit. The US Bureau of Mines conducted experimental mining and drilling in the Badger Hill area. In the late 1970s, Placer Service Corp. acquired a lease on the deposit. Between 1979 and 1984, Placer Service drilled 28 large diameter BADE (a German-manufactured machine) drill holes on the eastern portion of the deposit. The surviving records from the widely-spaced 1914 drilling program are fragmentary and the reported grade not well substantiated. The 1968 holes were drilled for scientific purposes. Again, drilling details are not available. However, detailed records for both the churn drilling program and the BADE program were available and formed the basis for the initial evaluation of the property. Geology The geology of the auriferous Tertiary gravels of California have been described by Whitney (1880), Lingren (1911), and, more recently, Yeend (1974). In general, the Tertiary gravels in the North Columbia area occupy a broad channel cut into pre-Tertiary igneous and metamorphic rocks. The upper, or white gravel is overlain conformably by volcanic tuffs and volcaniclastic rocks. A middle gravel is characterized by the presence of silicified and carbonized wood. A lower blue gravel unit has relatively coarser cobbles and contains a higher proportion of igneous and metamorphic cobbles than the other units. The upper gravel consists of interbedded pebbly sand and silty, or clayey sands with prominent cross bedding. Most of the pebbles are well rounded and consist mostly of white vein quartz and quartzite. The upper unit is moderately well compacted. Exposures in the walls of the old hydraulic mine pits stand at 45° and 50° angles. The gold content of the unit is well below an economic cutoff. The middle gravel - included with the upper unit by Yeend (1974) - is coarser grained, with carbonized wood, and 75 to 100 mm (3 to 4 in.) cobbles of metased-imentary and metavolcanic rocks in a sandy matrix containing abundant lithic fragments. The upper contact appears to be conformable, but the lower portion of the unit appears in places to consist of reworked lower gravels. The unit contains less clay than the upper unit and is somewhat more friable than the underlying lower gravels. The gold content, while somewhat higher than the upper level, is too low to be of ore grade. The lower gravel averages between 30 to 45 m (100 to 150 ft)

Jan 9, 1988

Introduction The objective of the design and operation of a mine is to make as much money as possible within certain reasonable and responsible constraints. The imposition of constraints on the objective of maximizing profits can result in an optimum mine design. This paper deals with the maximization of profits. Optimization will be discussed in a future paper. In the mining industry, it is generally accepted that the maximum economic recovery is defined by identifying the reserves and the sequence of mining that results in maximum net present value (NPV) or discounted cash flow-rate of return (DCF-ROR), as indicated in Fig. 1 (see Lemieux, 2000). Current parametric analyses techniques, such as varying the mining costs; profit reservation; sales price or metal content, to define pit limits and mining sequence technically referred to single, double or triple parameterization, do not assure that the curve in Fig. 1 will be maximized. In defining the sequence, the first material to be mined should be the first grouping of recoverable mineral that has the highest after-tax unit value and that forms a feasible mining unit. The assumed mining of the first pit should be followed by the assumed mining of the second most profitable and practical pit or pit expansion. If this process of grouping and sequencing is followed until the profitability approaches zero, the general sequence required to extract the maximum economically recoverable reserve is identified. The curve in Fig. I is constructed by calculating the NPV for a series of sequences, each assuming operations are terminated on successively lower profit pits. If the operation ceases on a pit with too high of a profit, opportunity is lost, and the NPV is lower than the maximum. The NPV curve peaks and then starts to decline before the zero-profit pit is assumed mined. This occurs because the money invested in advanced stripping would have produced a greater return if it was invested at the dis¬count rate. The maximum economically recoverable reserve is defined by the pit limit corresponding to the maximum NPV or DCF-ROR. Pit and phase definition The real challenge of reserve definition is identifying the phasing and after-tax profit per ton, so that the curve in Fig. 1 can be plotted. The after-tax profit cannot be properly addressed until after the pits are designed, the production is scheduled and the cash flow is estimated. Throughout the years, many investigators have ad¬dressed this problem. A major thrust of the famous paper in which Lerchs and Grossman (1965) presented their algorithm was to identify a sequence that would produce a result similar to that identified in Fig. 1. Whittle's (1988) four-dimensional analysis is designed to address this issue. The author addressed this same issue in a paper presented to the 1968 Canadian Institute of Mining national convention (Lemieux, 1968). Without a relatively simple methodology to con¬struct the curve in Fig. 1, the number of iterations in a trial-and-error method becomes prohibitively burden-some. The pit designer seeks a simple pit-planning parameter that will serve as a proxy for the after-tax profit per ton used in defining the curve in Fig. 1. The designer's goal is to apply this pit-planning parameter using standard pit-wall location techniques to design a series of pits that will provide a guide to the sequencing. Common techniques used to position pit walls are strip-ratio limits, highwall incremental analysis, floating-cone methods or "maximizer" applications1. Costs, revenues, noncash charges, taxes and profits are usually structured on a unit basis for use in the pit-wall location analysis. Application of the proxy parameter using the pit-design tools should, ideally, result in the definition of pits and sequences that will maximize the curve in Fig. 1. 1 The pit-design tool frequently referred to in the literature as an "optimizer" defines a pit boundary that maximizes the value contained within based on the input parameters and the safe pit-wall angle. The optimization of the pit design involves selection of rate of production, cutoff grade, product quality and other considerations. These considerations constrain the maximization of value. Therefore, pit-design tools that maximize value should be called "maximizers."

Jan 1, 2000

By Henry S. Brown

At least two dozen massive sulfide deposits in the southern Appalachians have supported significant mining operations. Literally hundreds more deposits are known to exist, but are too small to be of commercial interest. Essentially all that are being, or have been, mined are those deposits that were exposed at the surface when first discovered. It is reasonable to assume that many more massive sulfide deposits, large and small, exist hidden not far beneath the surface. The ideas I am sharing with you here today are based both on my personal studies of some of the deposits and on descriptions of deposits in the literature. An attempt will be made to account for the genetic interpretation problems raised by many of these deposits posessing both "syngenetic" and "epigenetic" features. In addition, an examination will be made of the possible relationships between the massive and the not-so-massive sulfide deposits in this region. Hopefully, these observations added to those of others will eventually lead to the synthesis of a better model on which we can base exploration programs in the southern Appalachians.

Jan 1, 1976

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By C. A. Moseley, M. B. Kahle

The Gulf Coast Region has been the center of extensive lignite exploration and development activity in the past 10 years. Currently, 27.2 Mt (30 million st) of lignite are mined annually, the majority of which is uncovered by large walking draglines. Projections indicate that production could reach 125 Mt (138 million st) by 1990. As mining moves to deeper depths, removal of overburden becomes the decisive factor in mine planning and cost. The geology of a region plays a key role in the selection of an overburden removal system. While alternate stripping methods must be analyzed to determine the most economical and operationally sound method, so far bucketwheel excavator systems seem to be the best method for removing large volumes of soft, unconsolidated material that is found at projected mining depths in the Gulf Coast region. Geology of Gulf Coast Lignites Extent Distribution of the principal tertiary lignite-bearing deposits in the Gulf Coast Lignite province extend from the Texas-Mexico border to the Alabama-Georgia state line (Fig. 1). The lignite-bearing strata includes portions of eight states: Texas, Louisiana, Arkansas, Missouri, Kentucky, Tennessee, Mississippi, and Alabama. Lignites occur primarily in the Wilcox, Claiborne, and Jackson groups in the Gulf Coast. The uppermost strata of the Midway group in Mississippi and Alabama also contain significant lignite resources. Reserves Commercial lignite seams in the Gulf Coast Lignite province range from 0.6-3 m (2-10 ft) in thickness. The estimated lignite resources throughout the Gulf Coast in seams greater than 0.9 m (3 ft) in thickness and up to 61 m (200 ft) in depth currently are estimated at 32 Gt (35 billion st). Table 1 shows these resources by state. Quality Lignite quality can vary within individual mining areas and seams. However, trends in quality occur because of the depositional environment, age, and west-to-east geographical location. Better quality lignites occur in the Wilcox group in Texas and Louisiana. Poorer quality lignites occur in the northern regions of the Wilcox group and those of the geologically younger Claiborne and Jackson groups. Environmental Conditions Gulf Coast province environmental conditions are ideal for successful mine reclamation. The lignite occurs in flat to moderately rolling terrain, a topography easily reproduced from the spoil dumps or spoil rows developed during the mining operation. High to moderate rainfall facilitates rapid revegetation. Evidence from current mine reclamation programs and ongoing test plot studies indicate that mined lands can be reclaimed to reach premining productivities. Sometimes productivities are even greater than that of premining conditions. Existing Mining Methods Present Mines & Production There are seven active mines in the Gulf Coast province, all located in Texas (Fig. 2). Lignite

Jan 8, 1983

By Tore M. Larsen

The Rødsand Gruber mining company is a subsidiary of the Norwegian industrial corporation Elkem AS, and its mine is situated in the northwestern part of southern Norway. The ore from this mine is a relatively low grade, magnetite ore with a small amount of vanadium. The ore bodies are very irregular in shape, resulting in a high amount of waste rock intrusion in the stopes. The annual underground production in the mines is approximately 1 million t. Because of varying dimensions of the ore bodies, both transverse and longitudinal stopes are employed. The inclination of an ore body is usually 1.04 to 1.59 rad (60 to 90°), and the horizontal width varies up to 100 m. Sublevel stoping has been the traditional mining methods in this mine. Until a few years ago, the ore was primarily blasted using horizontal blasthole fans drilled from Alimak raises. Recently, however there has been an increasing demand for a les expensive mining method. Alternate systems were considered, and finally a method based upon large blasthole stoping was found to be the most suitable. GENERAL LAYOUT The general layout of a stope is shown in Fig. 1 As with all the figures in this chapter, only the prin¬ciples of the method are emphasized and the details may not be valid for any particular site in the mine. In transverse stopes, the horizontal width of the ore determines the length of the stopes. The average length is approximately 50 m, the width of the stopes is 20 m, and there is a 10-m wide vertical pillar between. The distance between the main levels is 60 m. Usually a horizontal (sill) pillar 10 m thick is left below the mucking level. Each stope is first undercut along the entire stope length. This undercut is made with 50.8-mm (2-in.) diam drill holes in a fan pattern. At the moment this is being done by a trackless Atlas Copco Promec jumbo equipped with two Cop 130 pneumatic rock drills. The drill drift is 4 X 4 m, and the crater is 20 m high. This leaves a 36-m stope height to mine with large blastholes. To achieve a fair portion of vertical drilling withlarge blastholes, there are two parallel drill drifts at the top of each stope. These drifts are 4 m high and 9 m wide, separated by a 4-m wide pillar in the middle of the stope. To ease the drilling of perimeter holes, the drifts extend 1 m into the vertical pillar on each side of the stope. The pillar will thus be only 8 m wide on the drill drift level, but 10 m wide down the rest of the stope.

Jan 1, 1982

By Zefang ChenLi, Laijun Ma, Lei Mao, Fang Lian

"In our work, the preparation technique of high pure manganese sulfate directly from low-grade rhodochrosite ore (MnCO3) was studied and improved intensively, including the effective leaching process and the short purifying process. Based on the same ion effect, the repeated leaching of rhodochrosite with sulfuric acid is proposed to improve the solubility of Mn2+ and inhibit the dissolution of the impurities Ca2+ and Mg2+. Moreover, the repeated leaching process could make full use of rhodochrosite and lower the dosage of sulfuric acidas raw material. With the aid of theoretical calculation, Ba(OH)2 was chose to adjust the pH value of manganese sulfate solution, and BaF2 to remove Ca2+ and Mg2+completely in the process of purifying. The results demonstrate that the extraction ratio and the recovery ratio of manganese reached 94.3% and 92.7%, respectively, which shows the prospect of industrial application. In manganese sulfate production, the heavy metal impurities have been decreased to less than 1ppm. Moreover, the content of calcium, magnesium and sodium has been decreased to less than 20ppm, which meets the standards of high pure reagent for energy and electronic materials. Our study enlightens an approach to the sustainable application of low-grade rhodochrosite.Introduction Low-grade manganese carbonate ores (=14%) becomes gradually the main resources for exploitation and applications in China according to national mineral resources survey (Tan et al 2004 and Singh et al 2017). Therefore, the electrolytic deposition technique of rhodochrosite faces great challenges including low electrolysis efficiency, high-resource input and high-pollution discharge (Fu et al 2010, Michelis et al 2009, Zhang et al 2013). Recently, more and more amount of manganese series with high or super high purity such as MnSO4, MnO2, Mn3O4 and MnO is demanded with the development of energy and electronic materials. Due to low-grade ores containing multiple elements such as iron, silicon, calcium and magnesium, conventional preparation of manganese sulphate with high purity from electrolytic product is characterized by long production-cycle and high-energy consumption (Subrat et al 2011, Mpho et al 2013). By contrast, hydrometallurgical method combined with thermal decomposition, hydrothermal synthesis and sol-gel processes has been widely studied because of its high efficiency, low consumption and low cost (Chen et al 2008). But the key problem in direct preparation of manganese series form low-grade rhodochrosite is to remove completely the multiple impurities such as iron, silicon, calcium and magnesium (Liu et al 2014). So it is urgent to develop a sustainable alternative approach with character of short process, high efficiency and environmental friendly."

Jan 1, 2017

By Nick Chlumecky

INTRODUCTION Column-type supports are used to prevent roof falls in underground excavations. The first individual column supports were made from timber that was cut to fit the irregular heights of mine seams. Because the seam height of the material removed varied throughout the mine, timber posts generally could be used only once. During the I 920s, steel friction props began to appear in British and European mines, and, by the 1950s, thousands of these props were being used. However, the steel friction props were heavy, difficult to install, and difficult to recover. Furthermore, the friction props could not be preloaded, and their yielding behavior was rather erratic. In 1946, hydraulic props were introduced in En¬gland, and, within a few years, they replaced friction props throughout Europe. The first application of hy¬draulic props in the United States occurred in the 1950s, during early experiments with longwall mining. Modern technology and high-production mining procedures have combined to make the use of hydraulic props a very acceptable method of providing temporary heavy-duty support for production areas. In areas such as development headings, where only modest roof loads are expected, lighter-duty jackscrews, ratchet jacks, and light hydraulic props are used. ADVANTAGES OF HYDRAULIC SUPPORTS The acceptance of hydraulic supports has been wide¬spread during the past 25 years, proving their reliability and effectiveness in preventing roof falls. Some of the advantages provided by hydraulic supports are: 1) Individual hydraulic props can be installed within 30 sec, resulting in a minimum exposure of per¬sonnel and equipment to an unsupported roof. 2) Hydraulic props can be preloaded to more than 4.5 t (5.0 st), increasing the frictional resistance between the rock strata. Thus, the roof is kept tight, improving the behavior of the unsupported ground. 3) If gradual roof sagging or bottom heaving im¬poses an irresistible force, the hydraulic prop yields gradually and continues to support the ground. 4) The support capacities of the hydraulic props most widely used range between 10 and 20 t (11 and 22 st). Their support capacities are maintained during yielding to an irresistible force. 5) Hydraulic props can be recovered rapidly. Where no other supports are provided for the roof, as in pillar¬ing operations, the hydraulic props can be recovered by remote means. 6) The length of hydraulic props is readily adjust¬able allowing them to be reused in seams where roof height varies. 7) Studies of mining operations indicate that the use of hydraulic props increases both personnel safety and mine productivity. DESCRIPTION Figs. I and 2 illustrate typical hydraulic props used in mines in the United States. Fig. I illustrates a prop

Jan 1, 1982

By Palmer C. Sweet

Kyanite. andalusite, and sillimanite are anhydrous aluminum silicate minerals, and they have the same chemical formula, Al2SiO3. Dumortierite and topaz are also classified as aluminum silicates although they contain boron and fluorine respectively. Mullite (a rare mineral named for its discovery site, the Scottish island of Mull) is produced when kyanite, andalusite, or sillimanite are heated. Synthetic mullite, also classified as an aluminum silicate, is formed by melting alumina and silica, bauxite, and kaolin or alumina, kaolin, and kyanite, depending on the type of synthetic mullite desired. These materials are either sintered in rotary, periodic, or tunnel kilns, or fused in an electric arc furnace. The world's largest producer of kyanite is the United States. Domestic kyanite is presently produced from one company at two sites in Virginia; Kyanite Mining Corporation mines kyanite from Willis Mountain and adjacent East Ridge in Buckingham County in the central Piedmont province of Virginia. Internationally, kymite is produced in Australia, Brazil, India, Kenya, Sweden, and Zimbabwe; the former USSR also produces kyanite (Fig. 1). Andalusite is produced by Piedmont Minerals Company, Inc. in the Piedmont province near Hillsborough, NC. Andalusite is also produced in France, India, the Republic of South Africa, and Spain (Fig. 1). Sillimanite is produced in Australia, Brazil, India, Spain, and the Republic of South Africa; the former USSR is also thought to be a major producer (Fig. 1). Some of the countries produce only for domestic use. Synthetic mullite, made by sintering and by electric arc fusion, is produced by Mulcoa (Mullite Company of America) in Andersonville, GA, and by Washington Mills Electro Minerals Corporation in Niagara Falls, NY. Annual world production of kyanite, andalusite, sillimanite, and synthetic mullite is approximately 672 727 tons (Potter, 1988). GEOLOGY Mineralogy Kyanite is a blue to light-green, triclinic mineral that occurs in long, thin-bladed crystals and crystalline aggregates in gneiss, schist, and pegmatites. The mineral has a hardness of 5 parallel to the long axis of the crystal and a hardness of 7 across the crystal. Andalusite is a yellow, brown, green, or red orthorhombic mineral that occurs in thick, almost square prisms in schist, gneiss, and hornfels. Gem quality andalusite exhibits a strong pleochroism; brownish-green in one direction and brownish-red perpendicular to it. Andalusite also occurs in placer concentrations where it has weathered out of rock. Sillimanite is a brown, pale-green, gray, or white orthorhombic mineral that occurs in long, slender, needlelike crystals in schist and gneiss. Chemical and Physical Properties Chemical and physical properties of kyanite, andalusite, and sillimanite are noted in Table 1. Kyanite is the only mineral that irreversibly expands when heated (calcined to produce mullite) and may offset the shrinkage of most ceramic or refractory ingredients when heated.

Jan 1, 1994

By N. T. Cowper

Slurry viscosity ultimately affects the majority of mineral extraction processes. However, to date its role has been disguised because mineral processing equipment has been conservatively designed to operate at minimal slurry viscosity. With most ores this still meant acceptable solids concentration was achieved. However, with the advent of treatment of weathered ores the resulting higher slurry viscosity (rheology) has meant solids concentration has had to be reduced, often to unacceptable levels. To permit operation at higher solids concentration chemical dispersants are employed. These dispersants are expensive and their addition should be controlled. Slurry Systems Pty Limited (SSPL) has developed an On-Line Viscometer to monitor and control slurry viscosity, whether by controlling water addition or by controlling dispersant addition. Using this instrument the slurry rheology can be controlled at the optimum level, maximizing solids throughput, and minimizing water, reagent, and dispersant addition. This paper discusses rheology of typical mineral slurries, the relationship between slurry rheology and percent solids in a slurry, and the need for rheological control in mineral processing. The effects of rheology on various process steps, and the use of dispersants are reviewed. The paper concludes with summarizing actual case histories and cost impact of Theological control.

Jan 1, 1992

By Jonathan M. Pachter

The Bailey mining complex is located in Greene County, PA in the valley of Talley Run, a small stream tributary to the Enlow Fork of Wheeling Creek as shown on the location map insert found on Figure 1. The Bailey complex includes two Pittsburgh No. 8 Seam deep mines (Bailey and Bailey #2), a coal preparation plant, and two refuse areas (for disposal of coarse and fine processing waste). Clean coal is shipped from the complex via a 24.15 km (15 mile) rail spur, which connects to the Monongahela Railroad in Waynesburg, PA. In early 1981, Consolidation Coal Company (Consol) began a permitting effort aimed at beginning construction of Bailey, the largest underground coal mining complex to be built to date in the United States. Construction of the Bailey Mine began in December, 1981, less than six (6) months following the first application submittal. Clean coal was produced by the Bailey Preparation Plant in September, 1984. Effective project planning, including early formulation of a project planning team, identification of "red flag" issues, innovative responses to environmental questions and development of environmentally sound project designs, all contributed to the relatively quick permitting at Bailey.

Jan 1, 1985