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This paper presents the results of an investigation of the stability of the mass flow rates and average pellet size in the different streams of an industrial-scale balling circuit. The variables in this study include water content, drum speed and fresh feed rate. It was found that both the mass flow rates and pellet diameter are greatly disturbed when a change, positive or negative, in any of the variables takes place. For most of these disturbances to be damped don, a period of one hour or so is required. When a disturbance in the feed rate was made, the transient period extended to more than one and half hours (the duration of the experiment). It was observed that there is an interaction between the drum speed and moisture needed for producing an acceptable green ball product. The moisture content of the product pellets was found to be a function of the pellet size, with the larger pellets retaining a higher moisture content.

Jan 1, 1980

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 Fred N. Kissell, George H. Jr. Schnakenberg

INTRODUCTION A variety of instruments are available for measuring or monitoring the performance of underground mine-ventilation systems. In general terms, the instruments may be classified as those that measure air velocity and those that measure gaseous concentrations. All costs herein are in terms of 1978 US dollars. The mention of a specific manufacturer or device is not intended to be an endorsement by the US Bureau of Mines. AIR-VELOCITY INSTRUMENTS The basic instruments used for measuring the air velocity in mines are the vane anemometer and the smoke tube. Vane Anemometer Of the air-velocity instruments, the 102-mm (4.0¬in.) vane anemometer is the most common and is available as either a low- or high-speed type. The low-speed anemometer generally is the most suitable for measuring the velocities in ordinary airways. For a rough check of the velocity in an airway, it usually is satisfactory to hold the anemometer by hand, positioning it in the center of the airway for 30 sec. However, the resultant error may be as high as 25% , and such a hand-held approach is unsuitable when accurate or reliable measurements are required. To obtain more accurate measurements, the proper procedure is as follows: 1) Since holding the anemometer by hand generally causes the instrument to read about 15% high, it is mounted on a 0.6-m (2-ft) extension rod. 2) The airway is divided into equal right and left halves. A 1-min traverse is used in each half, moving the anemometer smoothly up and down in a zigzag pattern so that the entire half is covered within the allotted minute. 3) The manufacturer's correction table is applied to the readings to adjust the velocity calculation as necessary. Whenever possible, anemometer readings should be obtained in a long straight section of airway that has a constant cross-sectional area. Bends and obstructions should be avoided, since they cause turbulence and other discontinuities in the airflow and can degrade the accuracy of the velocity measurements. Although a series of velocity measurements at one location usually corresponds to within a few percent, this is not an indication that the airflows calculated from those readings are completely accurate. One reason is that the correction table provided with the instrument generally is not for that specific instrument; instead, it represents the average correction for all such instru¬ments made by the particular manufacturer. Most cor¬rection tables specify a correction factor of from 0 to 15%, depending upon the velocity. However, even after correction, the instrument error still may range from 3 to 5%. At low velocities such as those below 0.76 m/s (150 fpm), the instrument error can be two or three times greater than this, ranging from 6 to 15%. The new ball-bearing anemometers generally perform somewhat better at low velocities than did the older conventional anemometers. Another source of error is introduced when measur¬ing the cross-sectional area of the airway or entry. Under the best of circumstances, measurement errors, instrument errors, and a host of other minor errors all combine to cause a total error of at least 10% in the velocity calculation. The vane anemometer also can be used with reason¬able accuracy to measure airflows in mine-ventilation ducts. In this application, the anemometer is mounted on a rod and is held at the center of the duct end. For a duct that is discharging air, the average velocity in the duct is 85% of the centerline reading (Northover, 1957). For a duct that is taking in air, the average velocity is 70% of the centerline reading (Haney and Hlavsa, 1976). To measure the airflow discharged from a regulator or from a small hole in a stopping or bulk¬head, a correction factor for the area is necessary. A good approach in this situation is to traverse the area of the regulator or hole, holding the anemometer with an extension rod. This provides an average velocity that is multiplied by 85% of the measured area of the regulator or hole. In all cases, the manufacturer's instrument cor¬rection table must be used and applied properly. For accurate results, the anemometer should be returned to the manufacturer for periodic cleaning and checking. If it is in daily use, the anemometer should be returned about once per year, and proportionally less frequently if the usage is less frequent than on a daily basis. Smoke Tube The smoke tube may not appeal to individuals who believe that good measurement results can be obtained only with expensive, complicated, and fragile instru¬mentation. Nevertheless, smoke works about as well as anything for the routine measurement of low air velocities in mines. The following procedure yields reasonably good results: 1) Two marks are scratched 7.6 m (25 ft) apart on the floor of the airway. 2) The smoke tube is used to release a cloud of smoke in the center of the airway, about 0.9 m (3 ft) upstream of the first mark on the floor. 3) A timed interval begins when the leading edge of the smoke cloud passes over the first mark, and the interval stops when the leading edge of the cloud passes over the second mark. 4) A factor of 20% is subtracted from the cal¬culated velocity to determine the true average velocity, providing a correction for the centerline and for the spreading effect at the front of the cloud. Velocities calculated with the preceding method generally are accurate to within 10 to 15%. In some instances, the cloud from a conventional smoke tube

Jan 1, 1982

By David R. Maneval

At the turn of the century the keys to industrial success were coal and iron. At that time, Pennsylvania was the ranking mineral producer among the States and close to 90% of the Nations energy output was generated by coal. The growth of this Common-wealth into the greatest industrial State in the nation has been due to the huge resources of easily accessable coal found by the early settlers. It was the development of these deposits that provided the United States with a new source of fuel energy. For nearly a century, the Commonwealth was the chief powerhouse of our nation and the chief source of its industrial blood and bone. In 1928 it was truly said that coal was king in Pennsylvania. Pennsylvania that year mined a billion dollars worth of coal. A single train carrying this output would reach more than twice around the world at the latitude of Pennsylvania. Coal formed nearly 67% of the freight of Pennsylvania and was responsible for its four billion dollar iron and steel industry. Coal was the real reason why Pennsylvania although 32nd in size was second and not tenth or fifteenth in wealth, population and industry. Pennsylvania was slow to take-up her heritage as the greatest coal producing State in the union. But the knowledge that coal was present was not slow in accumulating. In fact the Indians were aware of its existence long before the white man arrived. They knew that many of the hills were underlain with black stones that would burn. It is reasonably certain that Arnold Viele, the first white man to cross the mountain in this State saw coal in them in 1692. But awareness was not enough. The population in these days was confined to southeastern corner of the State and the population at this time had little need or interest in the resources further inland. Pennsylvania anthracite was discovered in 1762 but was first regarded only as a curiosity. Commercial productions started in 1820. This stimulated the development of many native industries and extensive canal system and afterwards an elaborate net-work of railroads were constructed at great costs to bring the products of the hard

Jan 1, 1971

By M. L. Klohn

The Felder ore deposit is a 5,000,000-pound uranium deposit in the South Texas Coastal Plain. It occurs in the basal sand of an Oakville Formation (Miocene) alluvial system. The host sand is a carbonate-rich arkose which contains virtually no carbonaceous debris and has been reduced by the local introduction of hydrogen sulfide. The ore lies well within the reduced zone and occurs as coffinite and uraninite that fill interstices and coat and replace grains. The overall geometric configuration of the ore is that of a winged, crescentic ore roll. Weak mineralization extends the wings of the ore roll and gives greater expression to the roll character. Departures from the shape are controlled by discontinuities in bed- ding and by proximity of the ore to the surface. Associated with the uranium is a broad halo of anomalous molybdenum. Subsidiary mineralized trends suggest a pre-existing up-dip position for the main roll. Oxidation from the surface largely destroyed this previous roll by solubilization of uranium. Uranium subsequently migrated into favorably reduced sand and reprecipitated at the present roll position.

Jan 1, 1970

By W. N. Hamilton

British Columbia is geologically favourable for a rich variety of industrial mineral occurrences. Some have been produced for years, others overlooked until recently. Major production occurs for asbestos, gypsum, limestone, magnesite, silica and sulfur; lesser production for a dozen others. For several of the minerals, potential is much greater than existing production. New production opportunities exist also for diamonds, niobium and REE minerals, fluorspar, kyanite, nepheline syenite, talc, and several others. Alberta, though geologically less diverse, also has many industrial minerals. Limestone, salt and sulfur have the major production; clay and shale, silica, sodium sulfate and a few others lesser production. Opportunities exist for expanded production, and for new production in heavy (Ti -Zr) minerals, formation brines (Ca-Mg-I-Li), and gypsum.

Jan 1, 1987

By B. Berhe, R. D. Lawrence, J. Hinzer

"The fortunes of mining companies and their implied value are tied to cycles - both economic and commodity – within which they operate. Hence the determination of appropriate future metal prices is one of the most critical factors faced by mineral valuators especially for advanced projects. Price, and hence revenue, is usually the most sensitive input to the valuation model. Empirical studies of past forecasts show that the success rate for commodity price forecasting is very poor. This paper explores various approaches and comments on their strengths and weaknesses. We conclude that careful evaluation of long-term metal prices is a key element and is hardly a luxury that can be left to simple averages or rules of thumb. INTRODUCTION Discounted Cash Flow (""DCF"") analysis, is well suited for the valuation of advanced mineral development assets, for which there are identified mineral resources and some form of economic assessment (Pre-Feasibility or Feasibility Study). The DCF method is also the most commonly used method for determining value for such assets. It provides a means of relating the magnitude of expected future cash flows from profits (or losses) to the magnitude of the initial cash flow (investment) needed for capital expenditures to purchase or to develop an asset for commercial production. The method proves useful in making investment decisions as it allows for the quantification of the Net Present Value (“NPV”), as well as the Internal Rate of Return (IRR) and Payback Period on initial investment capital. Valuation by its very nature is a forward-looking exercise. The DCF method attempts to estimate the NPV of expected future cash flows, by selecting factors that will likely affect future cash flow based on informed analysis. One of the most critical and sensitive factors is the determination of appropriate future commodity prices. “The variables that have the greatest impact on a DCF evaluation are the reserves, the metal prices, and the discount rate.”1 The determination of future commodity prices for valuation purpose or during M&A negotiations is a hotly contested issue. There are a number of analysts, institutions and specialized commercial data suppliers that provide both short and long-term commodities price forecasts. Also, the use of an aggregated mean for all available forecasts, or consensus forecasted prices for valuation purposes is not uncommon. The methods used in making forecasts for each forecaster may vary and in many cases may not be publically disclosed. Nonetheless, the aggregated mean of all forecasted commodity prices of reputed institutions is usually taken to “represent the market’s view of likely future commodity prices.”"

Jan 1, 2016

By J. A. Herbst, L. B. Hales, R. Zaragoza

INTRODUCTION The Control Problem The need for automatic control in flotation plants stems from frequent and randomly occurring disturbances to the feed slurry. Typical disturbances occur in the solids feedrate, the slurry percent sol ids, and the mineralogy and flotability of the gangue and valuable minerals. In plants where semi -autogenous grinding mills (SAG) are used, feedrate changes are very large and changes in feed percent solids of 1m or more are common. Mineralogy changes between valuable minerals are common, as are changes in the host gangue rock. As these changes occur, the operator can attempt to compensate for them by making adjustments to the flotation cell levels, the dosage rates of the collectors and frothers, the pH, and in some cases the aeration rate of the cells. This problem is sometimes complicated by the fact that different sections of the flotation plant will be processing different ore types at the same time. Potential Savings with Control Flotation control brings about savings or plant improvements in two areas; the first is in improved metallurgy and the second is in reagent savings. Improved metallurgy includes increases in the average plant recovery and achieves target concentrate grade level s more consistently. Average recovery increases between 1.0 and 5.0 percentage points have been widely reported (Paakkinen, 1979; Sutherland, 1984). Reagent savings of up to 40 percent have also been reported (Hales, 1985). VARIABLES OF IMPORTANCE Control led, Manipulated, and Disturbance Variables The initial step in the analysis of the flotation control problem is identification of important variables which can affect the efficiency of separation. The efficiency of separation is represented by a group of controlled variables which can quantify the performance of a flotation circuit. This group includes variables such as grade, recovery, tonnage through- put, and circulating loads. The efficiency of separation of a flotation circuit may generally be defined by a grade recovery map, where each point on this map re- presents a set of operating conditions (see Figure 1). However, the combination of operating conditions must be set to operate close to the economic optimum. The group of operating variables which are changed in order to achieve desired circuit performance are called manipulated variables and can be classified into two

Jan 1, 1986

By Omar A. Muhtadi, David A. Milligan

7.1 INTRODUCTION The circulating fluid (generally an aqueous cyanide solution) is the major active component of heap leaching. This solution extracts precious metal from the heap and is then processed to recover the precious metal. The solution continuously recycles through the heap to recover more precious metal. With each cycle through the heap, water and chemicals are lost. Make-up water and chemicals must be added to balance losses from evaporation, seepage, entrainment, and chemical reaction in the heap, as well as any other processing losses. Adding more water than needed can result in a build-up of water as inventory in the various ponds and can eventually require that water be discharged from the operation. This is an extremely undesirable condition, and can often be very expensive in both investment in water treating facilities and operating costs. Most heap leach facilities operate under a zero discharge permit; thus, a situation requiring water discharge can result in permit violation. Chemical solution control must be managed to maintain the optimum pH, dissolved oxygen content, and cyanide content, while preventing "scaling". Calcium-based precipitation is a common scaling problem in mill reclaim waters. A buildup of these deposits can result in severe reduction of water flow, loss of pump efficiency and eventual interruption in metal production. In addition, problems of plugged sprinkler heads, blinded leach pads and fouled carbon (in carbon recovery systems) can be and often are experienced when scale control is not maintained. The remainder of this chapter discusses chemical solution control and scaling prediction/prevention in the context of a "closed" system. Sizing of ponds to accommodate precipitation and runoff, collection systems, and other similar design considerations are not discussed here. These topics are deferred to later chapters, e.g. Chapter 10 considers design of collection systems while Chapter 12 discusses pond sizing. 7.2 CHEMICAL CONTROL Chemical composition of the circulating solution must be controlled by the operator if the heap leaching process is to be successful. The primary or essential chemical control parameters in any heap leaching operation are:

Jan 1, 1988

By Donald C. Wotring, Douglas A. Gabriel, Michael G. Vitale

"The Northeast Ohio Regional Sewer District (NEORSD) recently completed the 24-foot diameter Euclid Creek Tunnel using a steel-fiber reinforced concrete segmental lining. During construction HMM, the McNally/Kiewit JV, CSI/Hanson JV, and the NEORSD performed a test program to evaluate replacement of steel fiber with synthetic fiber. A section of synthetic-fiber-reinforced cast-in-situ lining was also placed as part of the permanent works for the starter and tail tunnels. Laboratory and field testing were performed on two complete segmental rings reinforced solely with synthetic fibers. This paper summarizes the testing program, construction, and results of the synthetic-fiber-reinforced CIP lining and concrete segments. INTRODUCTION The principal feature of the Euclid Creek Tunnel (ECT) project is a CSO storage tunnel that is one component of the Northeast Ohio Regional Sewer District’s (NEORSD) Easterly CSO storage and conveyance system. The tunnel, shafts, and appurtenant facilities will be used for the collection and storage of CSOs from the northeastern part of Cleveland, Ohio and subsequent conveyance to the NEORSD’s Easterly Wastewater Treatment Plant for treatment. The ECT project includes five drop shafts that extend as deep as 220 feet below the ground surface. The ECT alignment and shaft locations are shown in Figure 1. Traditionally in northeast Ohio, tunnel excavations were supported with a two-pass system using steel ribs and timber lagging, followed by cast-in-place concrete. Control of rock overbreak proved difficult with large-diameter two-pass operation, and invert degradation led to train derailments. To mitigate these risks, the ECT was supported using one-pass, precast, bolted, gasketed concrete segments reinforced with steel fiber alone. Tunneling operations were very successful and experienced no significant delays or claims. As a result of the positive ECT experience with steel fiber-reinforced segments, Hatch Mott MacDonald, LLC (HMM), McNally/Kiewit (MK) joint venture, CSI/Hanson (CSI/H) joint venture, and the NEORSD performed a test program to evaluate replacement of steel fibers with synthetic fibers in segmental rings. This paper presents the laboratory and field testing program, construction, and results of the synthetic-fiber-reinforced cast-in-place (CIP) lining and concrete segments."

Jan 1, 2016

By G. H. Lewis, C. L. Krcil, N. S. Harding, D. E. Havekotte

INTRODUCTION Limestone injection work was completed in a 440-kW (1.5-million Btu per hour) pilot-scale combustor in preparation for a 50,000-kg/h (110,000-lb per hour) industrial steam boiler limestone injection test program. The main objectives were to: (1) identify the important variables that affect sulfur capture, (2) determine the maximum SO2 removal, and (3) determine the effect of limestone addition on furnace operability and heat transfer. A more detailed report is given by Lewis et al. (1985). The test program was divided into two phases. The objectives of Phase I were to determine the important variables that affect SO2 reduction and to investigate the effect of limestone addition on combustor operation and heat transfer surfaces. The objective of Phase II was to determine, using only the important variables identified during Phase I, the conditions that resulted in the greatest sulfur dioxide reduction. Phase I Test program Using one limestone, five variables were identified as having an effect on SO2 reduction. These are noted with their respective ranges: heat release rate [103 to 166 kW/m3 (10,000 to 16,000 Btu per hour-cu ft)], Ca/S molar ratio (0 to 3), flue gas oxygen concentration (3% to 5%), additive particle size [90% -150 µm (-100 mesh) and 90% -45 µm (-325 mesh)], and burner stoichiometry (SR = 0.75 and >1.15). A 2.3% sulfur northern West Virginia coal, pulverized to a nominal grind of 70% -75 µm (-200 mesh), was burned for the tests. The limestone used was obtained from a Maysville, KY, quarry and contained 51.75% (wt) CaO. An 11-run statistically designed program (plus three duplicates) was used to screen the five variables. The effect of limestone injection on the combustor heat transfer surfaces was determined by comparing the heat transfer rates through a slag panel and fouling probes during periods with and without limestone injection. Combustor operation was determined by noting any operating problems that occurred during the tests. Test procedure After furnace wall temperatures had stabilized at operating conditions, heat transfer surfaces (slag panel and fouling probes) and the convection section were blown clean, and a four-hour test was begun. The

Jan 1, 1987

By Zhe Liu, Liping Xiao, Jichi Bai, Xuefei Luan

"In order to study the removal efficiency of Cu2+ from acid mine drainage by prepared bentonite-steel slag composite particles, adsorption experiment was carried out. The composite particles were characterized by X-ray Diffraction analysis technique (XRD), Scanning Electron Microscopy (SEM) and Fourier Transform Infrared spectrometer (FTIR). The results show that: the composite particles can release alkali to neutralize the acid of acid mine drainage; the adsorption and chemical precipitation of Cu2+ occurred in the whole reaction process; the removal amount of composite particles on Cu2+ was 9.88 mg/g when the reaction reached equilibrium; the FTIR spectra revealed the existence of surface complexation; the SEM micrographs suggested that the composite particles would continue to adsorb and coagulate Cu2+ after the composite particles surface adsorbing Cu2+ and forming precipitate, namely, there was synergistic reaction of adsorption and coagulation; the XRD patterns further showed the existence of cation exchange and revealed that the states of Cu2+ in the surface of the composite particles was Cu-Si-O mineral phase and CuO(Cu(OH)2) polymerization precipitation. The bentonite-steel slag composite particles which can play a role of adsorption-coagulation synergism are excellent multifunctional green environmental mineral materials to treat acid mine drainage containing heavy metal ions.Introduction It is reported (Zuo and He 2013) that with the rapid development of industry, the world demand for mineral resources is increasing, while the problem of mine environmental pollution is becoming more and more serious due to the large amount of mining. The pollution of mine wastewater has become a global problem, in which acid mine drainage (AMD) is one of the most serious waste water. It is also reported (Cong and Zhao 2003) that acid mine drainage containing high concentrations of heavy metals and having low pH value, causing serious negative influence to the mining production and ecological environment. Acid mine drainage is one of the most serious environmental pollution to the mining industry. In recent years, it has been widely concerned by scholars at home and abroad."

Jan 1, 2017

By R. E. Mclaughlin, R. J. McLaughlin, A. H. Danzberger

Acid mine drainage (AMD) can occur as a natural phenome­non through oxidation of sulfide minerals. Oxidation by air and water can be exacerbated by mining. Because AMD can exhib­it low pH, it is capable of extracting heavy metals from mineral deposits. The result is highly acidic drainage with elevated heavy metal concentrations - the combination of which can cause extensive environmental damage. Because of the importance of these factors, this paper is focused on sulfate based AMD. Due to widespread mining, the occurrence of AMD is ubiqui­tous. As a result, several approaches to treatment have been developed. Chemical treatment using alkali is most commonly employed to treat AMD. There are alternative methods of treat­ment, as well as alternative ways of handling treatment sludges or concentrates. Dewatering of sludge or concentrates is beyond the scope of this paper. Within the family of chemical treatment processes, there are three principal recognized methods, namely, "Conventional," "High Density Sludge" (HDS) and "Heavy Metals Removal Process" (HMRP). All three are now employed or specified for treating AMO at mining sites which are being reclaimed under Superfund. Only limited data are available for more specialized chemical treatment processes such as "ferrate", borohydride membrane and electrochemical methods. Therefore, these data are not included here. Performance data for these three processes are presented in this paper and comparison among them is made. Comparisons include consideration of water quality, sludge characteristics, and operational ease. Capital and operational cost comparisons of these processes were addressed in a previous paper by the authors (see bibliography). The results presented here are impo1tant not only to metals mining, but are also applicable to other industries experiencing metals-laden acidic wastes including: coal mining and prepara­tion, metal plating, and electronics circuit board manufacturing. The acceptance of any process should be based on bench scale demonstration before implementation is contemplated. The three processes included herein are discussed in relation to the particular sites at which they are in use, or their use is planned. The sites selected were representative of the various types of processes and the wastewater parameters.

Jan 1, 1998

By Ernest A. Draeger, James W. Collins

Water only cyclones, sometimes referred to as hydrocyclones, are one of the newest types of coal processing equipment. These devices have achieved wide acceptance as fine coal cleaners in the past decade. Many metallurgical coal plants employ water only cyclones to process 0.5 mm x 0 size coal. Some of these plants have as many as 40 to 50 305-mm-diam units with each unit receiving as much as 9 t/h of raw feed. The growth in the utilization of hydrocyclones is due partly to the energy shortage, partly to the general increase in over-all coal preparation, and also because of a need for a simplified fine coal (0.5 mm x 0 size) gravimetric separator. Large diameter water only cyclones are also available and in use to process feed sizes up to 32 mm in size. However, the size normally employed for cleaning 0.5 mm x 0 coal are 356 mm or 305-mm-diam for single-stage circuits. In two-stage circuits, the secondary units are normally the same diameter as the primary units or smaller, such as 203-mm-diam. By using the small diameter cyclones and treating a closely sized feed, 0.5 mm x 0 only, it has been found that the probable error or imperfection of the water only cyclones is comparable to that obtained on the same size fraction treated in a large 610-mm-diam heavy media cyclone washing 38 mm x 0. The water only cyclone can approach the efficiency of a heavy media cyclone when cleaning fine solids because it creates its own dense media by accumulating high specific gravity particles in the wide angle conical section of the cyclone. A cut-away cross sectional view of a water only cyclone is shown in Fig. 1. It is also true that when treating a wide range of sizes (38 mm x 0) in heavy media cyclones, the smaller size particles (0.5 mm x 0) separate in proportion to the split of the media through the cyclone rather than by particle density.

Jan 8, 1980

By Nath S. Parate

With the extension of mining beneath residential and industrial areas, it is becoming increasingly necessary to extract coal and minerals in such a way that subsidence damage due to groundmovement is reduced to a minimum. Understanding clearly of the mechanism of subsidence development is a must in view of the environmental problems and the energy crisis today. This paper gives a simple analysis of various components resulting from subsidence and the influence of other factors on their development. An up-to-date review of the subsidence and strain profiles is presented, with particular references to the coal mining operations. The importance of this topic from mining damage to structures, buildings properties and environment in mining areas is undeniable. Field results are discussed.

Jan 1, 1982

By Robin P. Gardner

The mechanistic approach to comminution consisting of incorporating particulate breakage functions into a suitable population balance equation has been developed by a number of researchers over the past decade or so. (1) To date this approach has been developed and demonstrated primarily for batch-operated, laboratory or small-scale ball mills. However, since industrial mills are mostly quite large and operated in the continuous mode; suitable population balance models and methods for obtaining the necessary breakage and other model parameters must be developed. This paper describes and investigates a radioisotope tracer technique for this purpose, Population balance models previously developed(2,3) are adapted for this study.

Jan 1, 1974

By A. Toulet

Mining in Mexico has an immensely long tradition. In his first letter reporting to his sovereigns from Mexico in July 1519, Hernan Cortes wrote, "To our mind it is probable that this land contains as many riches as that from which Solomon is said to have obtained the gold for the temple ... " 1 But not only precious metals drew his attention; in 1524, short of tin needed for casting guns, he found it in Taxco. As an added bonus, those who went to find it "came upon a vein of iron which according to experts is of exceeding richness." 2 Since then, the mines of Mexico have yielded up enormous amounts of mineral and continue to do so. Mexico is one of the ten principal producers of minerals world-wide both in production volume and diversity. 3 Tens of thousands of tons of iron, sulphur, and coal come from its earth annually; it is the world's second largest producer of barite; third of antimony, fluorite, and graphite. 4 Copper, zinc, silver, molybdenum, arsenic and cadmium and others are wrested from its mines. The wealth is there; but maximizing its potential requires investment, technology, and markets that the government cannot itself provide. Thus mining, since the Mexican Revolution severely restricted by both law and policy, today looks to private investment to develop its potential. Beginning in 1989, Mexico's government cautiously moved to liberalize private sector involvement in mining; 5 using the experience gained from these initial regulatory changes, the current new Mining Law was adopted in 1992 6 and new Regulations were promulgated in 1993. 7

Jan 1, 1994

By Brain W. Lawrence

Sublevel stoping generally is a large-scale open stoping method. It sometimes is referred to as long-hole or blasthole stoping. This method usually is applied to strong ore bodies that require minimal support and are surrounded by strong country rock. The ore body should be fairly regular in shape and well-defined. The dip of the footwall normally should be sufficient to allow broken ore to gravitate freely, although the method has been adapted successfully to mine some of the flatter ore bodies. Sublevel stoping is not dependent on the width of the ore body. Widths of less than 6 m (20 ft), however, make the utilization of long-hole drilling techniques more difficult. The main criteria for sublevel open stoping are competent ore and stable host rock, regular ore boundaries, and a footwall dip that exceeds the angle of repose of broken ore. Basically, the method entails providing access to the ore body at various subintervals between the main haul- age levels in order to drill and blast the intervening ore. Stope drilling is carried out from drilling drifts on the sublevels, and the ore is blasted in slices towards an open face, which generally is vertical on the downholes and may be inclined towards the open face for the up holes. The blasted ore gravitates to the bottom of the stope and is collqcted through drawpoints. Fig. 1 illustrates a typical sublevel stoping mine. The method requires extensive ore body development with relatively high capital expenditures. How- ever, much of the development is in ore, and production costs are comparatively low. Productivity rates are in the 13.6 to 27.2 t (15 to 30 st)/manshift range. The drilling, blasting, and loading operations are performed independently, and equipment utilization is high. Large outputs can be obtained with few units and limited personnel. Dilution with waste rock may occur if ore boundaries are irregular or if caving occurs, but 100% of the ore within the stope usually is recovered. Pillar recovery sometimes is a problem. It is at this point that most of the waste dilution occurs. At the time data was gathered for this report (1973), the sublevel open stoping method was used by more than 21% of the metallic ore mines in the US (9 out of 42 mines producing 1088 + t/day (1200 + stpd) and provided more than 12% of the total metallic ore production from these mines. DEVELOPMENT Access to a sublevel open stope mine is gained by a slope or shaft that is normally sunk in the footwall of the ore zone, away from any possible effects from blasting or other production operations in the stoping process. Main level interval is selected and usually varies from 45.7 to 121.9 m (150 to 400 ft), depending on the vertical extent of the ore body, the number of drawpoints required to maintain output, and the eventual stope height. The main haulage drifts are located at the bottom of the stope horizon, either directly in ore or in the footwall, with access crosscuts at intervals. Raises are driven up through the ore to connect to the level above and provide access and ventilation to the sublevel drilling drifts, which are driven horizontally in the ore for the length of the stope. Depending on the drawpoint system to be used, the stope is either undercut by a raise-and-cone system, from a scram or slusher drift, or simply provided with a drilling drift from which access is made at intervals to the main haulage level. A raise driven either at the end or in the middle of the stope (if stope access is from both ends) provides a start for opening up a vertical slot across the stope to establish a free or open face for long-hole drilling and blasting. Sublevel stoping has proved to be a relatively safe and economical system.

Jan 1, 1982

By K. Mohamed, G. Rashed

A conventional mechanical anchor bolt consists of a smooth bar threaded at the anchor end, with a conical wedge, and a mechanical shell anchor. When the smooth bar is torqued, the conical wedge pushes the shell anchor against the borehole wall. Coal mines use mechanical bolts in addition to other types of bolts to control the deformation and stabilize the yielded coal ribs. Limited research has been conducted to understand how a mechanical anchor performs in a coal rib. Researchers from the National Institute for Occupational Safety and Health (NIOSH) conducted the study described in this paper to define the loading characteristics of mechanical bolts (stiffness and capacity) installed in coal ribs at five underground coal mines. Standard pull-out tests were performed in this study to define the loading characteristics of a mechanical rib bolt. Based on the strength of the coal seam and the conditions of the coal mines, the bolts were installed at different amounts of torque. A typical tri-linear load-deformation response was obtained from these tests. It was found that the anchorage capacity depends on the coal strength, and the anchorage stiffness depends on the installation torque. The outcome of this research provides essential data for rib support design. INTRODUCTION Rib falls are a serious hazard in underground coal mines. In the last 10 years, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States (MSHA, 2018). Although resin bolts has been used increasingly in the last few years to control coal ribs, mechanical bolting (mainly expansion-shell bolts) is still used in many underground coal mines. Recently, NIOSH researchers surveyed 90 sites in 14 underground coal mines in the Eastern United States and found that 79% of the mines determine their rib support requirement based on the depth, mining height, and the loading conditions, of which 35% use the conventional mechanical bolts. Conventional mechanical bolts are preferable for rib support because they do not require the miners to be too close to the rib during rib bolting. Hence, using the conventional mechanical rib bolts can reduce the risk of injuries and fatalities during the installation process. The mechanical bolt or the point-anchored bolt is the oldest type of rock bolt that is still in use today. The point-anchored bolt is a tensioned bolt anchored down the hole with a mechanical expansion shell or a grouted medium (Peng and Tang, 1984). Figure 1 shows the components of the mechanical anchor bolt. The mechanical anchor bolt consists of a plug, a shell anchor, a headed smooth bar with the top end threaded, a bearing plate, a hardened washer, and sometimes a friction-reducing washer. The shell anchor can be either standard or bail type. The shell of the standard type is held in place by a nut, while the bail anchor uses a strap to hold the shell in place (Peng, 2008). The diameter of the smooth bar is generally 5/8 in for the #5 bolt or 6/8 in for the #6 bolt. The diameter of the hole that the bolt is inserted into must be carefully controlled since an oversized hole can result in poor anchorage (Moss, 1971). The anchorage is obtained by applying torque to the bolt head, which in turn pulls the wedge down in the shell and expands the serrated leaves against the sides of the hole-wall. In the resin-grouted anchor, the resin replaces the mechanical expansion shell at the end of the hole. The anchorage is achieved by bonding between the resin, bolt, and the hole-wall.

Jan 1, 2019

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988