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By David W. Cannon, Russell V. Mann

In a dispersion of particles in liquid a net charge will develop at the particle-liquid interface. This surface charge is usually due to the adsorption of charged material from solution. The existence of this surface charge gives rise to the formation of the electric double layer of counter-ions which surrond the charged particle. The particle surface charge and the electrostatic repulsion which exists between similarly charged particles is the primary stabilization mechanism for lyophobic colloids (1). The separation of charge which occurs at the particle-liquid interface gives rise to several dynamic phenomena associated with colloidal systems or with solid-liquid interfaces in general. These phenomena are known as electrokinetic phenomena and the four classic electrokinetic phenomena are; electrophoresis, electroosmosis, streaming potential, and sedimentation or Dorn potential. The actual driving force for electrokinetic phenomena is not the surface charge per se, but the charge at the interface between the liquid which is hydrodynamically bound to the particle surface and the bulk fluid. This interface is known as the slipping plane or the plane of shear and the potential at this interface is the zeta potential (2). The factors linking the electrokinetic phenomena is that they involve a relative motion between the liquid and the charged particle or solid surface and the driving force is the zeta potential of the solid. In addition to the four classic electrokinetic phenomena there are two additional electrokinetic effects in disperse systems; the electro-acoustic effects. When an alternating electric field is applied to a cooloidal dispersion the particles will move in the field due to their net zeta potential. If there is a density difference between the particles and the fluid this motion will result in the development of an acoustic wave. The effect was discovered at Matec and has been termed the Electrokinetic Sonic Amplitude or ESA (3). ESA is the pressure amplitude generated by the colloid per unit applied electric field strength and has SI units of pascals per volt per meter. When an alternating pressure field (acoustic wave) is applied to a colloidal dispersion the inverse of the ESA effects occurs. A density difference between the disperse phase and the continuous phase leads to a relative motion between the particles and the surrounding liquid. This means that there will be a periodic displacement between the charged particle and the oppositely charged counter-ions in the electric double year. This displacement results in the development of an alternating dipole moment at the frequency of the applied field. This effect is termed the Ultrasonic Vibration Potential or UVP and was first predicted for electrolyte solutions by Debye in 1933 (4). UVP is measured in units of volts per unit velocity amplitude of the applied acoustic field or volts per meter per second. In 1938, Rutgers (5) and Hermans (6) pointed out that the effect would also be present in colloidal dispersions. A detailed theory for UVP effects in colloids, also called CVP, was first presented by Enderby in 1951 (7). Extensive studies of the UVP in electrolytes have been carried out by Yeager et al (8). Recently, O'Brien (9) has developed a general theoretical treatment of electro- acoustic effects in colloids and has derived a reciprocal relation linking ESA and CVP effects. The most commonly studied electrokinetic phenomena is electrophoresis. Electrophoresis is the movement of charges particles in an applied electric field. The velocity of the particle divided by the applied electric field strength is the electrophoretic mobility of the particle. The zeta potential can be calculated from this mobility (2). The magnitude of both the ESA and CVP effects are directly proportional to the electrophoretic mobility of the particles. The mobility determined by the two electro-acoustic effects is the dynamic or AC mobility of the particles.

Jan 1, 1990

By V. Anthony Cammarota

MINERALS INFORMATION PROGRAM Mineral data collection by the Federal Government actually began in 1882 with an appropriation by Congress to the U.S. Geological Survey (USGS). Although the Bureau of Mines was instituted in 1910, it was not until 1924 that the collection, interpretation, and dissemination of worldwide information on production, consumption, and availability of minerals became the responsibility of the Bureau. The objective of the Bureau's minerals information program is to maintain and provide reliable, timely, and comprehensive data for analysis of national minerals problems. This program is implemented by a staff of specialists who cover about 100 mineral commodities and the mineral industries of 168 countries. Data are also provided on the minerals industry on the State and county level. The specialists analyze information to identify trends and problems, forecast future U.S. and world supply and demand for the nonfuel minerals, and relate trends and developments to their impact on the domestic minerals industry. Data Sources and Publications The primary sources of data concerning the domestic minerals industry are the companies that produce, use, and trade in these minerals. Approximately 60,000 survey forms are sent to industry, on a monthly, quarterly, semiannual, or annual basis, to obtain the data. Domestic trade data are provided by the Bureau of the Census, U.S. Department of Commerce. Information on mineral developments abroad is developed by Bureau specialists from mineral reports supplied by U.S. embassies worldwide, from primary source documents, from a review of the technical mining literature, and from onsite mine and plant visits in foreign countries. The information is disseminated through a number of Bureau publications: "Mineral Industry Surveys" provide aggregated data on a monthly. quarterly, or semiannual basis to give timely information on the most important mineral commodities. "Mineral Commodity Summaries," published each January, provides preliminary annual data on supply and demand and discusses significant events of the previous year. The final data are published later in the "Minerals Yearbook," which is the definitive historical record for mineral statistics. "Mineral Facts and Problems," an encyclopedic work issued every 5 years, covers industry structure, technology, supply-demand analysis. economic factors, reserves, and forecasts of U.S. and world demand for individual commodities. "Mineral Perspectives" and "Mineral Issues" cover important international mineral developments and significant issues of current U.S. policy as they relate to minerals. Discussions and analyses of mineral trends and developments also appear in other Bureau publications, such as "Minerals and Materials" and "Information Circulars." These publications provide the basis for analysis and policy determinations by the Bureau and by virtually every Federal agency concerned with minerals, defense issues, the national economy, and foreign relations. The data are also used by industry in planning capital investments in mines and processing plants, by financial analysts, by U.S. delegations in their work with international commodity groups, and by a host of other minerals related organizations. Data Bases The data bases used by the specialists for preparing publications or in response to a continuous stream of inquiries include both those that contain individual commodity data from respondents, and those that are summary data bases on production, trade, capacity, employment and a number of other minerals-related items, not all of which are derived from data reported to the Bureau of Mines. The data bases maintained on the mainframes are not accessible by non- Bureau users. A 10-year data base of supply- demand data, published each year as an update of tables that appear for each commodity in Mineral Facts and Problems, is available on a floppy

Jan 1, 1987

By Jack O?Hearn

A liquid and solid waste management plan for an operating uranium ore processing mill was developed to efficiently and effectively dispose of existing solid wastes and related, affected surface and ground water resources. The goal of this management plan was to achieve satisfactory clean up of these existing wastes ad their by-products as well as to prevent similar situations from occurring in the future. An existing, modern system of synthetically lined tailings ponds, constructed after many years of mill operations, now provides an approved site for the disposal and containment of waste streams currently generated by the mill. This system was thought to be readily adaptable to serve as a key element of the waste management plan. The potential presence of extra capacity in these ponds could eliminate or reduce expenditures for new facilities required solely for proper management of liquid and solid wastes generated by operations prior to construction of the modern tailings ponds as well as for a range of anticipated future operations.

Jan 1, 1984

By N. B. Prenn

Case Studies in Open Pit Design Using Whittle 4D will be discussing the following four examples: 1) Optimization of process plant design using Whittle 4D as a sizing tool. In this case, the home office and project team had different views of the size of the final project. The home office wanted to reduce the size of the project to reduce capital expenditure, while the project team suggested a larger project. Whittle 4D was used to determine Net Present Value and Internal Rate of Return using the Whittle scheduler (FDAN) to schedule many alternatives, and calculate the NPV and return for each alternative. 2) Comparison of pit optimization methods: Lerchs-Grossman vs floating cone. Many pits are developed based on floating cone economics. A comparison of methods in one large gold mine resulted in the stripping ratio of the pit being reduced by 10% in the Whittle 4D optimization as compared to the floating cone optimization. 3) Pit optimization to determine if a new mill is justified. A large heap leach gold and silver mine needed to determine if a new mill was justified. The planning and scheduling would take several man months if completed without pit optimization and scheduling software. Whittle 4D software proved to be very effective in providing optimized pits and a mine schedule in a short time frame. 4) Pit optimization to determine how to mine and size. Open pit optimization determined where the underground operation should start, and whether the material should be processed by milling or heap leaching. One of our clients asked for help in determining where the pit should end and a new underground operation start. We used Whittle 4D to help determine where the pit should end and where to go underground. In addition the operation needed to determine how to process the material.

Jan 1, 1996

By William R. Henkle

High-calcium limestone has been intermittently produced from the Monarch quarry since the late 1800s. Initial production was used in sugar processing and for a smelter located near Salida, Colorado. During the 1920s, Colorado Fuel and Iron Company (CF&I) began to exploit the deposit as a source of flux stone for its blast furnace steel operations at Pueblo. The deposit was operated for many years as a captive quarry by CF&I. In 1984, CF&I closed the quarrying operations at Monarch; flux stone was no longer required since the blast furnaces were taken off line in 198 1. In 1994, Colorado Lime Company, a subsidiary of United States Lime and Minerals, Inc. acquired the property. Limestone was mined from the Castle Butte Member of the Leadville Limestone (Lower Missippian). Some high- magnesium dolomite was also mined during WWII from the Redcliff Dolomite Member which directly underlies the Castle Butte. Past production was from both an upper and lower quarry at elevations ranging from 10,000 to 11,000 feet (3,048 m-3,353 m) mean sea level (m.s.1.). The reserves at Monarch quarry are in a dip slope deposit, occupying the northwest slope of Monarch Hill. The structural geology of the deposit is complex on both a region- al and mine scale. The reserves are located on the eastern limb of the Syncline Hill syncline. The beds on this limb of the structure strike to the northeast and dip to the northwest from 30 to 45 degrees. A major reverse fault separates the reserves into an upper and lower deposit. Numerous smaller normal and reverse faults cut the deposit, requiring a flexible mine plan. At the Monarch quarry, the Castle Butte Member con- sists of two stratigraphic units. The Upper unit (Mlu) is a thick-bedded, fine-grained crystalline dolomitic limestone from 0 to 300 feet (0-90 m) thick. The unit generally has an Si02 content of less than 4 percent. A bedded chert horizon and a karst horizon at the base of the unit locally raise the Si02 content as high as 20 percent. The underlying Quarry Stone unit (Mlq) is a light-gray, massive, bedded, calcilutite/micrite limestone from 90 to 130 feet (27-40 m) thick. The Quarry Stone unit is remarkably uniform chemically throughout it's entire thickness. Average analysis is CaC03 - 95 percent, MgC03 - 3.5 percent, and Si02 - 0.5 percent. The Footwall dolomite unit (Mld) of the Redcliff Member of the Leadville Limestone is a dark gray to black, blocky, medium-to-thick-bedded, calcilutite/micrite dolomite from 150 to 180 feet thick (46-55 m). The upper 100 feet (30 m) of the unit is a low-silica facies with an average Si02 content of less than 0.5 percent. MgC03 content is con- sistent at 39.8 percent throughout the upper 50 feet (15 m) of the unit but varies from 15 percent to 39.8 percent in the lower portion (50 to 100 feet [15-30 m]) of the low-silica facies. The lowest portion of the unit, 100 feet (30 m) to base of unit is high in silica and is not considered chemical grade. Current in-place, surface-minable reserves at the quarry are estimated to be 14 million short tons (st) (13 million metric tons [mt]) of limestone and 50 million st (45 million mt) of dolomite. A 1.4 million st (1.3 million mt) stockpile of finely crushed limestone is also present on the property. Colorado Lime is presently processing approximately 2,000 st (1,800 mt) per month from the crushed limestone stockpile. This material is pulverized and sold to underground coal mines for use as rock dust, some of this product is also used for environmental purposes. Limestone mining is scheduled to resume in spring of 2000.

Jan 1, 2001

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

GENERAL HYDROLOGIC REGIME The total hydrologic cycle must be considered when evaluating potential ground water problems in a mining environment. Precipitation is the ul¬timate source of all surface water and essentially all ground water. Stream flow during low flow periods often orig¬inates as seeps or stream gains (base flow) where the water table intersects the land surface. A stream that is higher in elevation than the ground water system may lose flow, thereby recharging the ground water system. The magnitude of stream gain or loss is dependent upon the hydraulic conductiv¬ity and related properties of the underlying hydrostratigraphic units. Ground water moves constantly from recharge areas to discharge areas (streams and springs) usually at very slow rates. Large quantities of ground water may be in¬tercepted in many underground mining environ¬ments. Construction of an underground mine in saturated formations may encounter ground water inflows ranging from very small to very large and from continuous to intermittent depending upon the hydraulic properties of the hydrostratigraphic units penetrated. SURFACE WATER VARIABLES Precipitation occurs in different forms, varying amounts, and with a wide assortment of temporal and areal distributions in the various mining dis¬tricts within the United States. Mine inflow can be affected in a matter of hours by precipitation events where there is a direct hydraulic interconnection between the mine and the surface through fractures or a mine opening. The timing, form, amount, and distribution of precipitation should be evaluated in a mine inflow study. A portion of the precipitation that falls in a topographic basin moves as sheet flow to runoff channels (Fig. 1). Some of the precipitation infil¬trates the unsaturated soil zone where it can either move laterally as interflow to the surface system to augment stream flow or move downward as re¬-charge to the ground water system. The unsatu¬rated zone (called the zone of aeration or vadose zone) can range in thickness from essentially zero to a few hundred feet. The saturated zone or ground water system occurs below the water table (zone where hydrostatic pressure is greater than atmos¬pheric pressure) where all the pores are filled with water (Freeze and Cherry, 1979, p. 44). Moisture can be lost from any of the systems (surface water system, unsaturated zone, and ground water sys¬tem) by the processes of evaporation and transpir¬ation which return moisture to the atmosphere. Stream flow consists of runoff and interflow components plus an additional base flow compo¬nent (Fig. 2). Base flow constitutes ground water discharge to a stream. Base flow plays an increas¬ingly significant role as total stream flow declines. Base flow is important because it constitutes much if not all stream flow between periods of precipi¬tation when runoff and interflow have largely dis¬sipated. INTERACTION OF SURFACE WATER AND GROUND WATER SYSTEMS Recharge from a surface water system to ground water systems can occur via unsaturated or saturated flow. The amount of recharge is depen¬dent primarily upon the vertical hydraulic conduc¬tivity of the formations underlying the stream and the availability of water. Streams that flow over low hydraulic conductivity formations lose little water even if the ground water levels are significantly lower than the stream. A ground water system may contribute flow to a surface water system but nor¬mally only by saturated flow. This contribution usually takes the form of seeps, springs, and the base flow of streams. GROUND WATER VARIABLES Introduction The variables that affect ground water flow are discussed in increasingly complex theoretical situ-

Jan 1, 1986

By L. C. E. FEIO, R. RAMSDEN

In many parts of the world deep level mining relies heavily on the use of refrigeration for providing safe thermal underground conditions. The trend in South Africa in recent years has been to locate the refrigeration plants on surface and to use these plants to chill water that is then distributed throughout the mine. The surface cooling installation normally consists of a pre-cooling tower in which water is first chilled before it is sent to one or more refrigeration plants for further cooling. Presently, the majority of these refrigeration plants use CFC refrigerants; however, to conform with the Montreal Protocol, these plants will have to be replaced or converted for use with alternative refrigerants. The optimum design criteria for installations using the new generation of refrigerants will be affected by their characteristics and are likely to be different from those currently in use. In light of these developments the industry is examining a number of alternative systems for future installations including water vapour refrigeration. Factors that affect the decision on the type of refrigeration system to be used in mine installations are reviewed. INTRODUCTION Gold mining has been carried out in South Africa for over 100 years and most of the reserves close to surface have been mined out; this has led to mining taking place at greater depths, approaching 4000 m below surface. One of the consequences of operating at these greater depths is that it is necessary to install refrigeration plants to cool the ventilation air so as to provide safe and healthy working conditions. With the current low precious metal prices there is pressure on mines to cut costs and in recent years there has been considerable effort devoted to the design of more cost effective and efficient refrigeration plants and cooling systems. In the last 10 years there have been many studies into the use of ice for specific applications (Eschenburg et al., 1986; Hemp, 1988 and Ramsden and Lloyd, 1992) and at present there are three ice making plants in operation on South African mines, each of which produces ice in different processes. In addition, detailed design studies have been carried out on the "split ammonia system" (Napier and Patterson, 1992). In this refrigeration system the evaporator is located underground close to where the cooling is needed, whilst the condenser and heat rejection system is on surface. As far as refrigeration plants are concerned the trend has been to move away from standard package units to custom built refrigeration installations. This has given the design engineer greater flexibility in the choice of key components for refrigeration installations. For example both screw and centrifugal compressors are used, the heat exchangers for the evaporators and condensers may be either plate or shell and tube, and in a number of installations evaporative condensers are used. In addition, the following refrigerants are in use, R12, R22 and ammonia. The acceptance of the Montreal Protocol and its amendments by the South African government will restrict the use of CFC refrigerants and it will be necessary to re-evaluate various refrigeration systems, including those that do not use conventional refrigerants, such as air cycle and water vapour systems, for new surface installations. Although new systems and refrigerants are being developed, only the following refrigeration installations will be compared in this paper: - a refrigeration system using R134a which has properties similar to R12 - refrigeration system using R22 or a new alternative refrigerant with very similar properties - an ammonia refrigeration system - a water vapour refrigeration system The benefits of new cooling systems are not discussed in detail, since they are largely mine specific and outside the scope of the present paper. Since the optimum design for an air cycle system (del Castillo, 1988) is dependent on the mine layout, this system is also not included.

Jan 1, 1993

By Andrzej Frycz

The development of the theory of reliability creates more and more possibilities of its application in different domains of technics. The methods evolved to date, both theoretical analysis of the engineering systems and investigations techniques in practice can be put in use a d translated from one science to another one [1,3,7J . It needs to consider the specific of the different sciences. But it is necessary to take account of the specific characteristic of each line of science. The reliability investigations needs, justified in technical and economical viewpoint appear too in mine ventilation technics and systems. The mine ventilation system is a complex of workings and ventilation plants, which serve to supply the networks with a sufficient quantity of fresh air from the atmosphere and then carry off the vitiated air. The condition necessary for a ventilation system to perform definite task is the correctness of all its elements, that is a capability to preserve in a expected range of tolerance, the values of basic parameters. The element of a system which ceases to work correctly is called failed. The basic parameter of all elements of a ventilation system is the characteristic curve, pressure loss - air quantity relation ?p = f(Q) (1) This curve concerning the workings and ventilation stoppings is given by the formula: ?p = R Q2 (2) thus the resistance R (of the workings or stoppings) may be taken as a characteristic parameter. The course of this characteristic f(Q) will be given in a complex form in concern with fans, ejectors, air curtains. The failure of a ventilation plant or working is therefore an event, relying on a change of its characteristic, thus for a definite time the losses of the plant's or working's capability set in for the fulfilment of its function in the ventilation system. The failures are divided in viewpoint of their nature on sudden and progressive. We are in concern with the sudden failures which are characterised by a violent change of the ventilation element parameter for example a fall of roof, or silt up of the working, opening or failure of the ventilation stopping, the stand-off of the fan or ejector etc. We distinguish two kinds of failures on the viewpoint of their action on the air flow: the "opening" failure and "closure" failure. The prevalence of factors, which causes the failures is in connection with the intersection in the mine of roads of haulage, roads of underground transport of materials and lining elements, roads of people motion with airways, as also depends on the rock pressure reaction on the workings lining and plants placed there. Other factors will yet be decisive upon supply decay of fans, ejectors and curtains. The appearance of failure of the ventilation element is therefore of random nature and determines the change of the element, when it is passing from operating state to unoperate. Qualifying the ventilation system element to one of these states is easy, when the characteristic change is clear and occurs suddenly, that was assumed. The investigations conducted by Patrushev and Karnaukh [91 in concern with the ventilation stoppings resistance change in the case of their opening, had shown the possibility of resistance value change in the range of several hundreds higher.

Jan 1, 1980

By C. D. Anderson

"IntroductionTungsten carbide is a valuable material used in a variety of different applications, including the manufacturing of cutting tools, bearings and as a cost-efficient alternative to industrial diamonds. Both the high hardness value (Mohs 9) and scratch resistance of this material make it a valuable commodity for use in the mining industry.Usually, the production of tungsten carbide is a multistep process involving one or more of the following unit operations: hydrometallurgical digestion, solution purification, crystallization (as ammonium paratungstate), calcination, hydrogen reduction and carbon synthesis. The proposed three-step process eliminates three of these steps (crystallization, calcination and hydrogen reduction) by implementing a combination of hydro/pyrometallurgical techniques. These include alkaline pressure leaching, carbon adsorption and carburization roasting. By eliminating some of the intermediate processing steps, there is potential to reduce energy consumption and decrease overall operating costs.Methods and resultsInitially, alkaline pressure leaching with sodium carbonate (Na2CO3) was employed to selectively leach tungsten, as tungstate (WO4 2-), from an industrial Scheelite (CaWO4) flotation concentrate. Experimental variables included lixiviant (Na2CO3) concentration, temperature and residence time. Experimental constants, based on previous work, were percent solids at 13 wt% and agitation at 400 rpm (Queneau and Strathmore, 1969). Leaching results showed it was possible to extract 97% of the tungsten present. The most favorable conditions for leaching were determined: 2M Na2CO3, temperature of 180° C and a residence time of 30 minutes."

Jan 1, 2015

By R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981

By Fred Leighton

Miners have long known that rock noise, or the popping and cracking of the rock commonly heard during mining, can indicate instability of the mine structure. For many years, miners have "listened" to the rock talk and many times have interpreted changes in rock noise activity to be a warning of failure and have retreated from the potential failure area. Microseismics, or the study of rock noises, was begun in the early 1940s partly because of this historical fact. Microseismics is based on using geophysical equipment to detect and analyze rock noises on audible and subaudible levels. Thus, these systems are much more sensitive than the human ear and "hear" even more of the rock "talk" than do miners. Research has shown that micro-seismics can be used to precisely locate portions of a working area that are generating rock noise, and that rock noise release rate information from each area can be used to analyze its stability. On the basis of research results, microseismic systems are now commercially available and have been purchased and installed in many mines around the world. Introduction When a rock mass is subjected to changing stress conditions, such as those caused by mining, small-scale adjustments occur within the rock that release seismic energy. When this energy is in the audible range it is called rock noise. Those areas where stress changes occur are also the areas of the structure most likely to fail. Individual rock noises can be detected and analyzed to determine their precise location relative to the mine structure. Over a period of time plots of this data provide a pictorial representation of where stress changes are occurring. This is because rock noise activity tends to concentrate in those structure areas most actively adjusting to the changing stresses. Since these areas are the most likely to fail, they can be pinpointed and mapped relative to the structure. Rock noise rates, or the number of rock noises occurring per unit of time, also tend to vary dramatically before failure. Thus, the ability to locate the source of individual rock noises provides a means of determining where failure may occur, and rate counting within each area offers a means of assessing when failure may occur. This information, properly treated, can be used to avoid, control, or warn of impending failure. Successful applications of this technique have provided recent impetus to the effort of developing microseismics into a practical, reliable, and economically feasible tool. The phenomenon of naturally occurring rock noise was discovered in 1938 by Obert and Duvall (1939) who were measuring seismic velocities in mine pillars. Seismic energy other than what they were generating continually registered on their recording equipment. Further study by these researchers showed that the extraneous seismic energy being recorded was from rock noises that were being generated within the rock in highly stressed areas. Pursuing this interesting phenomena both in the laboratory and in the field, Obert and Duvall documented the dramatic change in rock noise rate before failure (Obert, 1941). They also established the fact that, in many instances, rock burst failures could be predicted by monitoring and listening to the rock noise activity in rock-burst-prone areas (Obert and Duvall, 1945). These early efforts clearly showed that microseismics had great potential as a tool for measuring or estimating mine structure stability. Extended testing, however, showed that sometimes rock bursts occurred with no apparent microseismic warning, and that at other times sharp increases in microseismic data were not accompanied by failure. Also, because precise location of individual rock noises was not possible at that time, one never knew where the failure was going to occur, only that failure near the observation point was likely. Thus, while the technique clearly offered promise, it was not considered practical then. In the mid-1960s, the Bureau of Mines undertook a new research effort to improve the microseismic technique. Major improvements were judged possible mainly because new and vastly improved electronic and system components had become available as an offshoot of space program instrument development. Thus, in 1967, development of a multichannel, broad band, microseismic system began (Blake and Leighton, 1970). Application of this system in rock-burst-prone mines showed it worked well and provided the incentive to develop methods whereby the source location of individual rock noises could be easily and directly calculated (Leighton and Blake, 1970; Leighton and Duvall, 1972; Redfern and Munson, 1982). The improved system and the ability to locate rock noises quickly showed that the microseismic technique offered new and increased poten-

Jan 8, 1983

By Ryne Goodnough

INTRODUCTION The Wyodak Mine is located in the northern Powder River Basin of Wyoming and is thought to be the oldest continuously operating coal mine in the United States, having recorded yearly production since 1923. The mine has expanded over its 80-plus-year history from an underground room and pillar operation to a modem surface snip mine. Currently, the mine feeds four on-site power plants using conveyors, has a filly automated truck load-out, and ships coal by rail to nearby Wyoming power plants. Annual production of 4.8 MTFT is accomplished on a single-shift basis with 53 fill-time production, maintenance and administrative personnel. In 1990 a decision was made to modernize the coal feed system to the power plant complex with modular conveying equipment. The project involved constructing coal storage silos, overland conveyors, feeder breakers and modular conveyors to follow the mining benches. Total project cost in excess of $25 million was used to construe or commission: A 2,150-TPH 48-inch overland conveyor with horizontal curve A 2,150-TPH 48-inch plant feed conveyor 13 shuttle conveyors, each 150 feet in length 2 front-end loaders equipped with large coal buckets 2 crawler-mounted feeder breakers with compatible discharge conveyor 1 high-lift shuttle conveyor with adjustable discharge height 1 high-angle conveyor 2 4,500-ton coal silos with automated truck load- out capabilities. Conceptually, the plan was to develop 33 MTONS of reserve adjacent to the plant complex and use front-end loaders to tram ROM coal to in-pit feeder breakers. The primary crushing would reduce coal to 6-inch minus and feed through a series of shuttle conveyors to an overland supply conveyor. The overland conveyor would feed the secondary crusher, and the crushed product would be moved to any of four storage silos using conveyors. The in-pit crushing and conveying system is a unique fit for the power plants and provided many interesting challenges for the coal mine to overcome. System integration, maintenance windows, equipment availability and the like had a major role in engineering the system to a power plant complex having no emergency coal stock- pile and limited plant storage. In crediting the designers and operators, the in-pit conveying system at the Wyodak Mine has delivered coal to the power generators since 1994 with little or no downtime. ENGINEERING CONSIDERATIONS Yearly production for the mine is based on 3.5 MTPY which requires about 1,550 TPH on an 8 hour, one-shift- per-day, 5-days-per-week basis. Considering efficiencies for starting and stopping, crew transport, loader trams, and operational delays, a nominal conveyor feed rate of 2,150 TPH or 72% system utilization was used in engineering design application. This ton-per-hour basis is used to size and commission the remaining components of the in-pit system. Coal quality requirements also played a role in equipment selection. The mine reserve grades vertically and coal from Upper and Lower benches needed to be mixed in any given proportion. This required equipment to be filly variable and capable of producing all or part of the 2,150-ton-per-hour requirement. Bisecting the 80-foot- high coal face for quality required construction of a high lift module capable of reaching a single overland conveyor 40 to 50 feet above the pit floor. Designers built a high-lift, adjustable and portable module specific for this application Having two bench systems capable of each producing the required feed for the power plants required computerized control processes to control on-line production. RSLinx, RSLogix, and RSView were chosen as the communication and control software, and a Data Highway Plus serves as the communication backbone. Each of the two in-pit benches are controlled with a data center equipped with Allen Bradley 5-40 processors. This

Jan 1, 2004

By Richard R. Klimpel

Part 2 of this article continues the discussion of a 10-year, multimillion dollar research and plant testing program on slurry rheology and grinding circuits. The first part of the article (ME, Dec. 1982) presented the basic concepts identified by the research and some laboratory test results. This section Illustrates typical Industrial scale test results and Identifies some industrial operating implications of controlling rheology by different methods. At least four controllable factors decide the rheological character of a slurry-slurry density or percent solids, particle size distribution, chemical environment, and slurry temperature. The second factor has two interrelated facets: the shape of the particle size distribution which controls packing behavior of the solids, and the fineness of the distribution. Finer particles increase interparticle forces and viscosity. As indicated in Part 1, during a given grinding test it is possible for all four factors to change. However, regardless of the particular settings of the four factors in a given test, if the resulting rheological character is either dilatant, pseudoplastic, or pseudoplastic with yield, the associated breakage rate is correlated with the current rheological character. It is obvious, for example, in batch grinding tests run at constant percent solids, that the second and third factors, where appropriate, are changing during grinding because size distributions are changing and the production of fresh surface area takes up unadsorbed chemical. Thus the corresponding rheological character change in batch tests with increasing grind time would be dilatant to pseudoplastic to pseudoplastic with yield. The degree to which this transformation occurs depends on the changing setting of the four factors over grinding time. In continuous grinding tests, one might logically expect that dramatic changes in any one of the four factors will be less likely to occur. It will be shown that continuous mill operations offer some unique opportunities to take advantage of possible rheological transformation by more direct operational control of the settings for the four factors. One extra observation noted in the rheological studies was the variability in the location and extent of region B (pseudoplastic behavior) for the various coals and ores tested. The location of region B was usually in the region of 45-55% solids by volume and was of the extent of 0-8%, or 2-11% with chemical addition. The corresponding increase in net production ranged from 0-10% in region B and 0-21% in region B&apos;. When region B is small or zero (no pseudoplastic character is exhibited) no increase in production will be observed and the use of chemicals is often marginal. There are several reasons for some materials exhibiting this quick transformation from dilatancy to high yield values often at surprisingly low percent solids by volume such as 30%. One condition identified was for materials containing high levels of viscosity-producing elements such as- carbonates or clays. A second condition documented was for materials that exhibited unusually fine primary fragment distributions (fine B1,; curves). This corresponds to materials having small y values of < 0.5. Such a slurry developed a yield value quickly during grinding because of the rapid buildup of fines. A related problem can be presented by materials exhibiting excessively coarse primary fragment distributions, y values > 0.9. Breakage of this type of material produces size distributions that give poor packing efficiencies even at long grind times, thus hindering the normal rheological transformation presented earlier. A final condition that can cause the occurrance of region B to be small or zero is when the media void volume filling of slurry is < 100%. Figures 5a and 5b use previously published data of this study to demonstrate the influence of solids loading and weight percent solids on the net production of taconite ore in a laboratory batch ball mill. In particular, Fig. 5a illustrates several trends not generally recognized until this study. They include these two: • Increased slurry density allows for increased solids loading before passing through the maximum in the net production curve where the fall-off is due to non first order breakage. • The use of rheology control chemicals such as GA-4272 allows this trend to be extended to higher slurry loadings with an increase in net production over any previous condition by keeping grinding first order. Figure 5b shows the same data as Fig. 5a plotted for constant weight loadings. These various figures of net production versus percent solids show the location and extent of the regions A, B, and C presented in Part 1 as a function of solids loading. It is obvious that the rheological transformation pattern described earlier does not hold true for slurry loadings corresponding to less than the void volume of the media, which is also a region of

Jan 1, 1983

By H. Polat, Q. Hu, M. Polat, S. Chander

The electrostatic charge on spray droplets of ionic surfactant solutions and coal particles was measured and the results were correlated with the dust collection efficiency. When various surfactant were added, the magnitude of the droplet charge increased significantly and it was observed to be a function of surfactant type and concentration. The concentration of maximum droplet charge coincided with surfactant concentration where maximum collection efficiency was observed for these surfactants. Particles of coal also carried substantial amount of charge magnitude of which seemed to be a function of coal rank. Based on the results presented in this paper, it was concluded that ionic surfactant primarily act as a strong electrostatic charge inducer for droplets. Due to interactions between these highly charged droplets and naturally charged particles, the efficiency of droplet-particle collisions play a primary role when compared to the wetting and engulfment phenomenon which could only follow a successful collision. INTRODUCTION Water spray are widely used to suppress airborne dust in mine atmospheres (Walton and Woolcock, 1960; Kobrick, 1970; Hamilton, 1974; Jayaraman et al., 1986). Several investigators have considered the use of surfactants to enhance the effectiveness of water sprays especially for difficult to wet particles such as those of coal (Glanville and Wightman, 1979). The capture of dust particles by water droplets involves droplet-particle collisions, adhesion of particles to droplets, and engulfment of particles into droplets. Surfactants affect these sub processes through their influence on droplet charge, surface tension, and wetting. The last two mechanism have been thoroughly studied in recent years (Walker et al., 1952; Cohen and Rosen, 1981; Glenville and Haley, 1982; Chander et al., 1988; 1991). However, little attention has been paid to the role of electrical charge on particles and droplets on the collision and adhesion of spray droplets and dust particles. Airborne particles of dust have long been known to carry a significant amount of electrostatic charge (Hopper and Laby, 1941; Kunkel, 1948; Kunkel, 1950; Dodd, 1952; Liu et al., 1987; Kutsuwada and Nakamura, 1989). It is reasonable to assume that presence of charge on particles will effect their agglomeration and particle-droplet interactions. Polat et al., (1991) showed that virtually all freshly generated dust particles were agglomerated in air. They suggested that electrostatic charge and humidity were important factors responsible for agglomeration. Previous theoretical studies on the interactions between charged particles and collectors by Nielsen and Hill (1976) show that, the collision efficiency is a strong function of the particle charge. In addition to charge on particles, spray droplets might also carry substantial amount of charge (Chapman, 1937;1938; Blanchard, 1958; Iribarne and Mason, 1967; Jonas and Mason, 1968; Byrne, 1977; Bailey, 1988). In theoretical studies of interactions between a spherical collector and airborne particle, it was found that the collision efficiency was significantly altered depending on whether the collector and the particles were charged. If neither the collector nor the particles carried a charge, the collision occurred by inertial and gravitational forces. The collisions took place on the front part of the collector (the front capture). If either of the collector or the particles were charged, the collision was enhanced due to the induced image forces. If both the collector and the particles were charged, the collision efficiency was significantly affected by the sign of the charge as well as its magnitude. For oppositely charged collector-particle pairs a collision could take place on the rear of the collector even if the particle flied past the collector upon approach (the rear capture) (Nielsen and Hill, 1976; Wang et al., 1986; Chang et al., 1987). On the other hand, the columbic force became negligible as the particle size increased and the inertial force became dominant. The electrostatic attraction was predominant for particles of less than about 2.5 µm in diameter. For particles larger than about 8 µm the inertia of particles was sufficient to overcome the columbic force and inertial impaction became the dominant collision mechanism (Grover and Beard, 1975; Chang, 1987). Previous studies of dust suppression using charged spray droplets generated by applying high voltage to the spray nozzle showed significant improvements in collection efficiencies (USBM open file report, 1983; McCoy et al., 1985). However, it was considered that highly charged spray droplets obtained by direct charging might have

Jan 1, 1993

By Jozef S. Stachulak

Inco Limited operates 11 underground mines in the Sudbury District. The mines are located on the rim of the Sudbury Basin, an oval with the axis in the range of 27 and 60 km. The ore dips to at least 3000 m below surface. The ores are mined primarily for nickel and copper. Total ore production from underground is in excess of 55,000 tons per day. Over 40 surface fans have been installed since the late 1960&apos;s. All of the fans are adjustable pitch, axial flow units. A major factor influencing ventilation design in the last 30 years has been the introduction of diesel equipment underground. Volumes per fan have ranged from 60 to 330 (cubic metres per second), with motors from 100 to 2500 hp. Fans of the axial flow type have been in common use for main fan installations at Canadian mines for many years. The standard arrangement has been to mount these fans horizontally, i.e. with the fan shaft and the long axis of the housing horizontal. This is a natural arrangement for an underground fan, but for a surface installation, a vertically mounted fan has definite advantages. The surface area taken up by a typical vertical fan installation is generally about one quarter of that with a horizontal fan of the same capacity. (1) This is not a problem with isolated fans and flat surface outcrop sites, but where the installation is to be near existing buildings, or where there are poor surface soil conditions, space and cost considerations greatly favour vertical fans. MAIN FANS INSTALLED There are 43 main surface fans in operation at 11 mines. Twenty-seven of these fans are supply units, and 75% of them are vertical installation. The remaining 16 units are main surface exhaust fans, with predominantly horizontal installation. Within the last five years, some 20 main booster fans have been installed underground at several mines. Axial flow fans, with adjustable pitch blades, are used for both surface and underground installations. Exhaust fans are equipped with stainless steel or cast aluminum blades. Main underground fans are arranged horizontally and the majority of them have a floating shaft between the fan shaft and the motor. (2) The size of the fans in service varies from 1.8 m to 5 m in diameter, with the majority ranging from 1.8 to 2.5 m. The pressure produced by these fans varies from 0.25 kPa to 2.0 kPa. At Inco Limited, two main fans in parallel are preferred, rather than a single fan, so that if one fan fails, the remaining fan can still supply up to 70% of normal air quantities, while the damaged fan is repaired. This requires closure doors on each of the fans so the fan can be isolated in case of failure. It is more expensive than a single fan, but results in less production interruptions. The fan installations are well away from sharp inlet and outlet bends. FAN DESIGN INTEGRITY Both the mine operator and the fan manufacturer must understand that the main fan is critical to the mine operation, and that everything technically possible must be done in design and manufacture to ensure the highest degree of reliability. Some of the design parameters and criteria, based on Inco experience, are outlined and discussed below. RESONANT FREQUENCIES AND HOUSING MODEL ANALYSIS Any fan assembly will have many different resonant frequencies. It is a challenge to the designer to arrive at a design in which forcing frequencies do not coincide with any of these resonant frequencies to produce unacceptable vibration levels in operation. Finite element analysis is a useful tool that can be used to identify the most critical of these, so that housing and blade stiffness can be adjusted to change any resonances that might be close to forcing frequencies. Shaft critical speeds should be at least 25% above the fan operating speed, and there should be sufficient separation between other resonant and forcing frequencies to avoid excitation that might result in high vibration levels. QUALITY ASSURANCE Radiographic Blade Examination Mine fan blades are normally cast in small lots. To ensure that the castings are sound, a full radiographic examination is recommended in the highly stressed lower 1/3 of the blade.

Jan 1, 1995

By B. P. Boisen

INTRODUCTION The rapidity of onset, rate of increase, and magni¬tude of loads in an underground support system can be measured using load cells or pressure cells, whichever is appropriate to the type of support. These instruments should be installed at the various instrumented station locations immediately after excavation. If the loading of a member is of interest, the load should be measured, whereas if the deflection or strain of a member is of interest, the strain should be mea¬sured. It makes as little sense to use a load cell to allow computation of strain as it does to measure the strain and then back calculate the load. Tunnel steel set de¬signers today invariably rely on the early work of Karl Terzaghi, the basis for which is load, so load measure¬ment should be the main concern. The current trend of using strain gages on steel sup¬port systems stems from the inability by some to eval¬uate unusual loading conditions caused by uneven block¬ing. In fact, uneven blocking will obscure almost any attempt (whether with load cells or strain gages) to properly evaluate support performance. The load curve in Fig. I shows the development of a characteristic early peak load sometimes called the "rear abutment load" seen in many underground open¬ings. It is thought that the peak reflects the coupling of the rock and the support system, and that the magnitude will increase until some yielding, or minor failure, of the support system occurs. At that time, the slight deforma¬tion of the support system promotes the formation of minute shears in the opening walls, and these shears tend to distribute stress between the support system and the adjacent rock in proportion to the relative rigidity of the support elements and the rock. In normal rock, as a result of this stress redistribu¬tion, subsequent load magnitudes generally do not reach the magnitude of the early peak load. In physically unstable (squeezing) or chemically un¬stable (swelling) rock, however, the loads experienced after passage of the early peak load may in fact show a slow, continuous increase. One of the very important purposes of load instrumentation is to provide the means for recognizing such long-term adverse trends, thus enabling the proper remedial steps to be taken. Another purpose of load instrumentation is to pro¬vide a comparison between the magnitudes of the early peak load and the subsequent stable load. The ratio of these two loads is analogous to a safety factor, and may be used to evaluate the efficiency and economy of the support system design. Aside from considerations of economy, it may be well to design support systems which do not have excessively high ratios of early peak load to subsequent stable load. Should these ratios be exces¬sively high, the support system may be so rigid that the yield or failure associated with stress redistribution may occur with explosive violence. Load cells for use in mines, tunnels, and on con¬struction projects come in many forms. Almost all, however, employ the same procedures for installation, readout, etc. Therefore the following comments are almost universal in application. LOAD CELL INSTALLATION Load cells should be installed with bases parallel to the surfaces against which they bear. Care must be taken to orient the cells so that their signal cables are protected from accidental damage as a result of con¬struction, maintenance, or cleanup activities. Most electronic load cells are compensated for tem¬perature variations likely to be encountered during nor¬mal operations. However, if a large difference is an¬ticipated between the calibration temperature [21°C (70°F)] and the average operating temperature, the cells should be conditioned to the operating temperature for at least 8 hr prior to installation. This is to insure that the initial reading, made under no-load conditions prior to installation, provides a stable value to which subse¬quent measurements can be referred. Hydraulic load cells tend to be temperature sensitive and should be used with that in mind. Also, hydraulic load cells tend to be soft compared to electronic types and will sometimes allow movement to take place in a system intended to be semirigid. Furthermore, hydrau¬lic load cells tend to be difficult to read remotely. Care must be taken to use bearing plates on both sides of load cells which are sufficiently rigid and of high enough bearing capacity to prevent bending and crush¬ing under load. This is very important with tieback load cells (basically a ring of steel) which can easily dig into bearing plates. MAINTENANCE AND TROUBLESHOOTING Except for a direct hit by a miner&apos;s axe or flyrock from a blast, most hydraulic load cells are nearly in¬destructible and require little, if any, maintenance. Hy¬draulic oil on the bearing plates is a good indication of a leak and the need for corrective action. Field maintenance of electronic load cells involves protecting the instrument and signal cable from mechani¬cal damage and from unnecessary exposure to dirt and moisture, and recognizing and correcting damage and the effects of normal wear and tear.

Jan 1, 1982

By S. K. Sharma

The Bureau of Mines has investigated the fundamentals of the flocculation process. The objective of this study was to determine the effects of various physico-chemical parameters such as shear rate, polymer dosage, and polymer concentration on floc growth and to determine the applicability of a model in predicting polymer flocculation behavior. A rapid in-line particle size analyzer was used to follow the progress of both the floc growth and the breakage as a function of time. A genetic algorithm (GA) was used to determine the unknown parameters in the model and to fit a curve to the experimental data. Results show that the model, using GA as a tool to determine the unknown parameters, reproduces the experimental data with only reasonable accuracy and was relatively insensitive to the parameters. Results of the flocculation study showed that increases in polymer dosage had greater effect on the time required to reach the maximum floc size than on the floc size itself. It was also determined that, given all the aforementioned parameters, the degree of slurry mixing during the flocculation process had the greatest bearing on the floc growth.

Jan 1, 1993

By Warfield Garson

Few, if any, of the non-ionizing radiation health hazards to be found in either surface or underground coal mining are uniquely different because of their being found in the work environment. Hence, they can be considered generally for their bio-effects on the worker when found in the mining work environment. The same may not be said, however, for the lack of non-ionizing radiation and its bio-effects, particularly as it relates to underground coal mining. The term "non-ionizing radiation" refers to various forms of electromagnetic radiation of wavelengths longer than those of ionizing radiation. As the wavelength gets longer the energy of electromagnetic radiation decreases. Therefore, all non-ionizing forms of radiation have less energy than cosmic, gamma, and X-radiation. In order of increasing wavelength, non-ionizing radiation includes ultraviolet, visible light, infrared, microwave, and radiofrequency radiations. The energy frequency and wavelength range of both the ionizing and non-ionizing electromagnetic forces are shown in Table I. To convert the wavelengths of various radiations to Ångström units, one multiplies millimicrons by ten. In a vacuum, all electromagnetic radiation has the same velocity, namely 3 x 1010 centimeters per second. The natural source of radiant energy here on earth is our sun which emits radiation continuously over a wide spectrum. This radiation on average reaching earth ranges from 290 nm in the ultraviolet range to over 2,000 nm in the infrared range with a maximum intensity of about 480 nm in the visual range. You will note this falls into the visible blue wavelength and accounts for our blue sky and blue ocean and deep water effects. We are all familiar with the fact that the passage of solar radiation through the atmosphere to the earth changes the spectrum considerably because the atmosphere absorbs and scatters many of the sun&apos;s rays. The ozone in the upper atmosphere absorbs the shorter ultraviolet wavelengths and water vapor absorbs some of the infrared wavelengths. Smoke, dust particles, gas molecules and water droplets scatter the rays, especially those of shorter wavelengths. In addition to the sun, every gas, liquid or solid object at a temperature above absolute 0° radiates energy. Solid objects emit almost continuous spectra. At low temperatures only radiation of the longer wavelengths in the infrared range is emitted, but as the temperature of the object is increased, more and more of the shorter wavelengths are added. This fact is most readily demonstrated by heating a piece of steel. When a piece of steel reaches a temperature of about 1,700° Fahrenheit, it gives off radiation at the red end of the visible spectrum and appears dull red. As the temperature is further increased, the shorter rays are also emitted, until at about 2,100°F, the metal appears white, due to the emission of wavelengths throughout the entire visible range. Gasses, on the other hand, when heated emit radiant energy only at certain wavelengths, which are characteristic of their chemical structure. This latter fact is of importance in underground coal mining as high intensity gas and vapor lamps are becoming more and more utilized for illumination in underground coal mining. The biologic effect of non-ionizing radiation exposure depends upon the type and duration of exposure and on the amount of absorption by the miner. The effects of this radiant energy on the miner fall into four distinct types: (1) the heating effect of infrared radiation, (2) the effect on the eye of visible radiation, (3) the effects of ultraviolet radiation, and (4) the growing potential effects of the misuse of microwave radiation. Each non-ionizing type of radiation will be considered individually. ULTRAVIOLET RADIATION The sun is the major source of ultraviolet radiation, which is of concern in open pit and surface mining at certain seasons and in certain climes necessitating protection for the surface miners under those conditions; nonetheless, there are some man-made sources such as electric arc lights, welding arcs, plasma jets, and special ultraviolet bulbs for illumination underground that demand surveillance in the underground environment to be aware of whether the miners are at risk above the threshold limit values allowable. Since ultraviolet radiation has little penetrating power, the organs that are affected are the skin and the eyes. Ultraviolet radiation is strongly absorbed by nucleic acids and proteins, and the effects in man are largely chemical rather than thermal. Short-term effects on miners include acute changes in the skin. These are of four types: (a) darkening of pigment, (b) erythema (sunburn), (c) increase in pigmentation (tanning) and (d) changes in cell growth. Ultraviolet radiation also causes acute effects on the tissues of the eye. Overexposure can lead to keratitis, inflammation of the cornea, and conjunctivitis. Long-term effects of ultraviolet exposure include an increase in the rate of ageing of the skin with degeneration of skin tissue and a decrease in elasticity. Late effects of ultraviolet on the eye include the development of cataracts. The most serious chronic effect of ultraviolet exposure is skin cancer. Ultraviolet radiation effects are increased by some industrial materials and drugs. After exposure to such compounds as cresols, the skin is exceptionally sensitive to ultraviolet radiation. Photosensitivity reactions occur after exposure to a variety of other chemicals and drugs including dyes, phenothiazines, sulfonamides, and sulfanylureas. On the other hand, we must remember that ultraviolet radiation has an important role in the prevention of rickets. Vitamin D is produced by the action of

Jan 1, 1981

By Richard E. Ray

INTRODUCTION Main mine fan installations at underground coal mines must be designed to comply with the requirements outlined in 30 CFR (Code of Federal Regulations) Part 75. However, very little emphasis is typically placed on the aerodynamic optimization of shaft collar designs, ductwork configurations, and the selection of a fan isolation, or "closing door," system. Field testing has shown that poorly designed systems may produce large pressure losses between the shaft and fan inlet, as well as air turbulence problems which adversely impact fan efficiency and sometimes fan structural integrity. This paper analyzes each component of a surface exhaust fan system for an underground coal mine: • The shaft collar arrangement, • The 45 or 90-degree turning duct that isolates the fan motor from the return airstream, • Alternative ductwork arrangements for parallel fan installations, • The self-closing doors or dampers that are required for multiple fans, and • Horizontal and vertical evas6 arrangements on the discharge side of the fan. While discussion is limited to axial fan installations at shaft mines, most of the recommendations would also apply to centrifugal fan installations and to main mine fans for drift mines. FEDERAL REGULATIONS The mandatory safety standards for underground coal mines that apply to the installation of main mine fans are listed in 30 CFR Part 75.310. Stipulations which impact the design of the fan sys- tem arrangement are contained in paragraphs (a)(l), (a)(5), (a)(6),(d), and (1) of 75.31 0: • Each main mine fan shall be installed on the surface in an incombustible housing; • Each main mine fan shall be protected by weak walls or explosion doors in direct line with possible explosive forces; • Explosion doors or weak walls shall have a cross-sectional area at least equal to the area of the return air shaft or drift opening through which the pressure from an underground explosion would be relieved; • Each main mine fan shall be offset by at least 4.57 m (15 feet) from the nearest side of the mine or shaft opening; • Automatic closing doors shall be provided to prevent possible air reversals through the fans in mines ventilated by multiple main mine fans. The installation of main mine fans underground is prohibited by 30 CFR Part 75.310, (a)(l). Providing explosion doors or weak walls in line with a possible underground explosion and offsetting the fan 4.57 m (1 5 feet) from the edge of the shaft preclude the installation of the fan directly over the shaft. In addition, 30 CFR Part 75.507-1 specifies that all electrical equipment used in return air courses be permissible, which requires locating the fan motor outside of the airstream exhausted through the fan. To comply with the Part 75.310 regulations, explosion doors are typically located directly over the shaft as part of the shaft cover or &apos;bonnet." A sufficiently long section of ductwork placed between the shaft enclosure and fan, equipped with a 45 or 90- degree turn, enables the fan to be offset at least 4.57 m (1 5 feet) from the shaft and the motor to be isolated from the airstream. The need for the turn in the ductwork can be eliminated by driving the fan from the discharge side, instead of from the inlet side, and providing for a 90-degree bend in the evas6. For multiple main mine fan systems, self-closing doors can be located on the inlet or discharge side of the fan. FAN SYSTEM COMPONENTS A poorly designed exhaust fan arrangement will produce turbulent conditions at the fan inlet and significant pressure losses between the shaft and fan. Field tests have shown that pressure losses through inlet ductwork at some installations may constitute as much as 30 percent of the total mine head. Unfavorable flow conditions may cause a change in the operating point on the fan performance curve, resulting in reduced fan efficiency and pres- sure generation capability. In some instances, poor airflow distribution through the fan may create mechanical stresses that lead to fan blade or fan housing fatigue failures. The objective in designing an air duct arrangement for a fan system is to provide a low cost duct that will move the air from the shaft to the fan in a uniform air distribution and with minimum energy requirements. The optimum solution is a compromise between technical and economic factors. Aerodynamic design considerations for the shaft collar and cover, fan inlet ductwork for both single and parallel fans, self-closing doors, and evas6 are discussed in the sections below. Shaft Collar/Cover To prevent excessive turbulence and non-uniform air distribution through the fan, a smooth transition for the change in direction of airflow at the top of the shaft must be provided. However, the concrete lining of the shaft is abruptly terminated at the surface at many coal mine installations (see Figure I), creating a sharp 90-degree bend. The airflow exiting the shaft becomes heavily concentrated in the upper portion of the horizontal fan in-

Jan 1, 1997

By K. Tokutake

Kobe Steel, Ltd. improved the quality of iron ore pellets for blast furnace burdens when it enlarged its blast furnaces. However. there still remained room to increase reducibility at high temperatures for better blast furnace operation. In particular, it was still necessary to lessen the reduction retardation above 1200?C. In the last few years, by employing iron ore and additives of optimum size structure, large pore pellets with a high porosity have been obtained. By using this method. Kobe Steel. Ltd. has accomplished the technology to produce pellets with remarkably high reducibility.

Jan 1, 1989