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By Erling Stranden, Leiv Berteig

INTRODUCTION The measurement of the total activity of radon daughters in the air of mines has become a routine procedure in order to control the radiation exposure in miners due to the inhalation of these radionuclides. Normally the measured concentration is given in terms of total potential L-energy of the short lived radon daughters. The unit most frequently used is 1 WL (working level), which is defined as the potential a-energy of 1.3.105 MeV per 1 of air. The a-dose to the respiratory tract is, however, not proportional to the inhaled potential a-energy. In addition, the degree of equilibrium between the daughter products in air and the fraction of daughter products not attached to aerosol particles (the unattached fraction) must be known. Theoretical calculations of the equilibrium between daughters and the unattached fraction have been presented by different authors (Jacobi, 1972; Porstendörfer et al, 1978). Such calculations must take into account the plateout of daughters on surfaces, the attachment of daughters to aerosols and removal of daughters by the ventilation. The situation in mines will however be very complex and no single model will cover all situations. In most cases the radon and radon daughter concentrations will be described by a combination of different models. Domanski (1979) used the experimental results of Breslin et al. (1969) from uranium mines to relate the equilibrium between the individual daughters to the equilibrium factor, F. Once the equilibrium factor is known, this seems to be a useful approach to assess the individual daughter concentrations. The concentrations of radon and daughter products may vary considerably during the day. Seasonal variations are also frequently found. It is therefore important to have knowledge of the magnitude of these variations and of the factors having the strongest influence upon the concentrations. The unattached fraction of radon daughters in uranium mines was investigated by George and Hinchliffe (1972) and Raghavayya and Jones (1974). Mercer (1975) later modified the results of Raghavayya and Jones. Both these investigations indicated that the aerosol particle concentrations was the main influencing factor upon the unattached fraction. In this paper we will summarize the main results of a study on the radiological characterics of nonuranium mines. We will also discuss the correlations between the unattached fraction of the potential a-energy and the unattached fraction of the individual daughters, and between the equilibrium factor F, and the individual daughter ratios. More detailed discussions of this study will be published elsewhere (Stranden and Berteig, 1981 and Stranden and Berteig 1981 a). MATERIAL AND METHODS During 4 two-weeks periods corresponding to the seasons in the Norwegian mine where the highest radon daughter concentrations occure, the parameters listed in table 1 were studied. TABLE 1 Parameters measured during the study [Parameter Methods Radon Scintillation flask/Continuous ionization chamber Radon daughters Tsivoglou(1953)/Kusnetz(1956) Respirable dust Cyclone Unattached fraction Wire screen (Stranden and Berteig 1981) Atmospheric pressure Continuous baragraph Temperature and Continuous thermohygrograph humidity] At a fixed location, the radon concentration was measured continuously by a ionization chamber (Stranden, 1981), during these two-weeks periods in order to study short-time and seasonal variations. RESULTS Correlation between equilibrium factor and individual daughter ratios. The mean value of the equilibrium factor in all Norwegian mines measured during the last ten years was 0.46 with a minimum of 0.08 and maximum of 0.93 (Stranden and Berteig, 1981). The correlations between the equilibrium factor F, and the individual radon daughter/radon ratios were studied and the results indicated high correlations. In table 2, these correlation formulae are listed together with the correlation coefficients. TABLE 2 Correlation between the equilibrium factor, F, and individual radon daughter ratios [Formula Correlation coefficient RaA/Rn = 1.14 F 0.551 0.76 RaB/Rn = 1.00 F 0.940 0.99 RaC/Rn = 0.77 F 1.120 0.92]

Jan 1, 1981

By K. Mulukhov

Introduction The depth of open-pit mines continues to increase. The typical designed depth for large quarries is now in the 300 to 600 m (1,000 to 2,000 ft) range, with some quarries even reaching depths of 700 to 800 m (2,300 to 2,600 ft). The cost of handling material from such deep mines can be 60% to 70% of the total cost. The capacity of using a cyclic means of transport, such as trucks or skip hoists, continues to drop as the depth of the mine increases. In addition to the problem of maintaining relatively low slope angles for haul roads, increases in environment pollution and increased fuel consumption make the use of trucks ineffective below 100 to 150 m (300 to 500 ft). In handling large quantities of bulky materials, i.e., more than 15 to 20 Mt/a (16 to 22 million stpy), under fixed terminal points, belt conveyors are preferable in almost in all cases. An exception is when the lump size of the load is too high for the conveyor to handle. Belt conveyors in mines can handle loads having a maximum lump size of no more than 250 to 300 mm (10 to 12 in.). The use of a movable crusher prior to the belt conveyor in an open-pit mine would make the process very expensive. Such an operation could not compete with truck haulage on a cost basis. Belt-car conveyor The main factor that restricts the lump size for conventional belt conveyors is collision impacts between the load-carrying belt and idlers. These impacts are completely eliminated in the belt-car conveyor, which was previously described by Mulukhov (1977). The unique feature of the conveyor is that the load-carrying belt is supported by moving cars. Each car consists of a troughed cross strap that supports the belt with the load and two rollers, on which the cars move along upper and lower rails. Cars are connected by two endless chains revolving over sprockets. The system uses conventional drive pulleys, tail pulleys and return idlers. There are no fasteners between the carrying belt and the car cross straps. The conveyor belt simply drives the cars along by the force of friction between the belt and the straps. A belt-car conveyor was first installed in 1970 at the Karatau mining operation in Kazakhstan. The system was an elevating conveyor with an inclination of 20° and was located on the final part of the truck haul road. The conveyor transferred blasting phosphate ore from the quarry to the crushing plant. A bunker was mounted between the trucks and the conveyor. The maximum lump size was typically 1,200 mm (48 in.). However, there were some lumps as large as 1,500mm (60 in.), which was greater than even the belt width (Fig. 1). Because the prototype belt-car conveyor was intended only for industrial trails, it had a relatively short length of about 50 m (160 ft). Nevertheless, the conveyor yielded a substantial profit over the cost of truck haulage. Naturally, the first question one might ask is why such a system is not now widely used. The reason was the further development of a huge transporting complex for the whole quarry at the base of the belt-car conveyor. After successful industrial trials from 1971 through 1973, which demonstrated the systems engineering feasibility, the State Planning Committee of the USSR included the development of the belt-car conveyor in its plan for the development of the national economy. However, mistakes were made. First, a single organization was not in charge of the development. Second, a plant that was not specialized in belt conveyors was put in charge of the design and manufacturing of the conveyor system. In addition, the crusher was installed before the stacker, which negated the major advantages of the belt-car conveyor. For loading the conveyor, a special blade feeder with pendulum suspension was developed, but it was found unreliable. The main conclusion was that, in-

Jan 1, 2003

By Peter V. Avotins

During last year's AIME meeting in New Orleans, Deepak Malhotra and I were discussing some of the topics that would be presented in this Symposium. It seemed appropriate that along with papers on specific reagent and flocculant types, we should address research management aspects of new product development. This development is a long, arduous task that requires a significant technical and fiscal commitment by the chemical company and potential users. A flow chart depicting the research process is shown in Figure 1. The sequence of events are plotted on an undefined time line. Our experience has shown that it takes 3-5 years to complete a typical reagent development project. A new project starts with an idea of how to do something better. The ideas range from relatively simple product modifications to opportunities created by technological change. An example of this is the use of flocculants in tailings operations to close the water recycle loop. The change was brought about by the need to conserve water and comply with environmental statutes. The result was the development and wide-spread use of synthetic polymers to help in solid-liquid separation. More recently the ups and downs of oil prices have led to development of coal water slurries as an alternative to fuel oil. The slurries contain high levels of chemicals (up to 3%) to insure proper stability and burn characteristics. There are technical changes under way right now that may lead to ideas for other new products. One exciting source of new technology is the ability to manipulate genetic material of microbes to select for specific capabilities. This has evolved by leaps and bounds over the past 5 years. Biologically assisted leaching and removal of wastewater impurities are two possible end uses. I'm sure that there are many others. The ideas can originate anywhere and are often half-baked at first, but to lead to new products they must be refined to provide project definition and eventually a research plan. A chemical ,development project cannot be started without approval to commit the company's resources. The internal selling job is often very difficult. The company must see clear financial benefits in order to launch efforts that may not pay off for many years. It is critical to project the total market, profitability, eventual market share, and timing of introduction of the new product. There is much uncertainty in this and the research manager is often skating on thin ice to retain his credibility. Once we obtain support from upper management, the research team is assembled to start the project. The teams generally consist of organic chemists, research metallurgists, and technicians. They have access to analytical scientists and technicians for surface and analytical chemistry. The projects are often augmented by help from university staff and consultants. This is particularly true where there is a need to develop fundamental understanding of the technology in addition to synthesis of chemicals. The project now proceeds along two lines. First, there is an emphasis on understanding the prior art. Extensive literature and patent searches are undertaken to provide research leads and to keep from "reinventing the wheel" and infringing on someone else's patent. The scientists and technologists are qualified to conduct computer-assisted library searches. This process usually uncovers 80-90% of the relevant information. The remainder is very much harder to get and requires the help of professional library search staff with scientific backgrounds. Along with literature searching, the team begins synthesis of new compounds. These are screened by use of techniques such as microflotation or filtration in a Buchner funnel. The scientists are looking for the direction of response as they modify structural features . Inevitably, a large number of candidates reach the next stage, a laboratory bench test. At this point it is possible to set up a research plan and use project management techniques to monitor progress. The bench tests are conducted in larger equipment, usually requiring 100-1000 grams of test substrate. Typical laboratory flotation tests are run in Denver or Wemco cells.

Jan 1, 1986

By Justus B. Deen, Randy Peterson

INTRODUCTION The General Chemical Soda Ash Operation located near Green River, Wyoming produces about 4.5 Mt of trona per year. In July 1989, Mine Ventilation Services Inc. performed a ventilation survey of this complex trona mine which had thirteen active panels spread over 36 square kilometers. The ventilation survey and subsequent study showed that the intake shaft air velocities were at their practical limits (Wallace and Rogers, 1987). The need for additional intake capacity was immediate. The ventilation network modeling delineated the best location for this new intake shaft as it applied to the 5 year ventilation plan. Unfortunately, this location, 6.5 kilometers from the soda ash processing plant, would make heating this intake air very costly. Presently, the intake air is heated by heat exchangers which use waste heat from the plant. Air heating during the winter months is necessary for miner comfort and to prevent potential freezing of water lines in the mine. It was decided to examine the feasibility of heating the air geothermally by coursing it through an old production panel. This method of air heating has been used in this region with good results at the Stauffer Chemical Operation (now Rhone-Poulenc) (Moore, 1985). In 1990, a geothermal heating study examined four different production shaft/panel configurations and a workable design was found. The operators at the General Chemical Soda Ash Operations acted on this design and began raise boring a ventilation shaft in the Spring of 1991. This paper describes the ventilation system, briefly, and the geothermal heating studies performed. Findings of the 1990 geothermal studies are then compared to field data acquired in the Winter of 1992. THE VENTILATION SYSTEM The ventilation system at General Chemical had four shafts serving the mining horizon in July 1989; three of these shafts are located at the Northwest end of the mine. The active work areas had progressed to the South over the past years and supplying ventilation has become increasingly difficult. The 6.1 meter diameter #3 Production shaft and the men and materials compartment of the 6.1 meter diameter #2 Split Shaft had reached their air velocity limits. Conversely, the 4.6 meter diameter #5 Exhaust Shaft was under utilized. Ventilation studies showed the conversion of the 3.7 meter diameter #1 Exhaust Shaft to an intake would provide adequate intake capacity for two years. The next option to improve ventilation was the addition of an intake shaft at the southeast part of the mine at the intersection of H Mains South(H-M-S)and J Mains East(J-M-E) (See Figure 1). A service shaft has been planned for the extreme Southeast end in the late 1990's. Ventilation network analyses showed this interim intake shaft would have to be at least 2.5 meters in diameter to postpone the service shaft construction beyond 1995. The greater the size of the borehole, the further the construction could be postponed. In October 1991, General Chemical commissioned the 4.57 meter diameter #4 Ventilation Shaft near the junction of H-M-S and J-M-E as shown in Figure 1. The L95 panel would be used for air heating as described in the geothermal heating section to follow. GEOTHERMAL STUDIES Geothermal heating studies were performed by MVS Inc. in May 1990. Using a modified code of CLIMSIM, an underground mine climatic simulation model, several different shaft/panel configurations were examined to see which would provide adequate air heating during the winter months. CLIMSIM simulates heat flow into a single underground airway. The program was used to evaluate the heat flow to and from the airway surface and to calculate the resulting change in dry bulb temperature of the ventilating air. CLIMSIM uses inlet air conditions, airway characteristics, and rock thermal properties input by the user to predict the variation of psychrometric and thermodynamic parameters along an underground airway.

Jan 1, 1993

By Betty L. Gibbs, Jean A. Arcamone

Introduction Microcomputers are becoming more convenient to use at all levels in the business world, including the mining industry. The state of the mining industry requires that we become more productive and innovative in our analysis of deposits and operations. With the help of computers, examination of options to increase productivity can be performed quickly and inexpensively. Effective computer use to design underground workings combines the calculating power and speed of a computer with the technical expertise and judgment of an engineer. Computer programs can perform in minutes calculations that might take several months for an engineer. But the engineer must be on guard when using computers for calculations. It is easy to assume the results produced by a computer program are always correct. However, the computer is only a sophisticated calculator. It cannot recognize logical errors or transposition of numbers. Also, some intuitive or obvious decisions made by an engineer are difficult or impossible to program effectively. In this article, selected commercial programs are discussed. These programs can provide engineers with tools to evaluate engineering designs for underground mines. The programs perform much of the tedious calculation. The engineer, though, must still make decisions and conceptualize designs. The software mentioned in this article runs on microcomputers and minicomputers. The mention of particular software, however, does not imply endorsement. Mining projects After a decision has been made to proceed with an underground mining project, more detailed evaluation of the mining method and extraction sequence begins. If a mine is already in operation, schedules and plans are needed to more clearly define future production. Existing operations can benefit from identification of bottlenecks in material movement and produce extraction options to make the mine more efficient. Detailed studies for a new mine are based on the technical engineering design. These studies provide specific details to determine equipment requirements. Detailed studies done during a project are similar to short-term planning and control carried out daily by the mining engineers. The engineering design phase includes: • Study of the infrastructures - entries, shaft, access from surface to underground; • Definition of the mine layout, hydrogeology, geotechnical considerations, transport, and ventilation; • Choice of machines and equipment; • Mine scheduling, including waste extraction and production sequencing; and • Economics, including labor and supplies. Further detailed studies include: • Blasting patterns; • Detailed geotechnical calculations; • Secondary ventilation; and • Production cycles along with coordination of the workings. Objectives The primary objective of underground mine planning is moving material out of the mine at the least cost. Development of the production schedule considers location and volume of ore units, transport distance, equipment and manpower interactions, and grade or quality of material. Two methods can be used to determine the scheduled material removal: optimization or trial and error. The microcomputer programs discussed here do not cover optimization for underground mine planning. For several reasons, optimization is often a "black box" approach. Although optimization is possible with linear programming techniques, few mining operations can be described effectively with linear equations. An interactive trial-and-error approach is desirable because engineering judgment can be inserted as a plan develops. Because of the large number of factors in an underground operation, an optimization technique seldom covers all possibilities. Few of the programs available consider all of the possible parameters. The programs use an interactive sequencing or a simulation approach and are either event-oriented or deposit-oriented. The most desirable programs offer an integrated approach. They consider a broad range of operating conditions including equipment, labor, costs, material movement, timing, and mine layout, among other factors. The two types of models generally used in designing underground workings include: • The interactive model, which requires direct, incremental input from the user. Outlines of the section to be mined may show on the screen. The engineer causes the model to develop piece-by-piece. The program may offer feedback on parameters such as tons and grade, or other values, as the section is "mined." • The simulation model, which is usually an automatic technique. Parameters related to the process are entered. The program then calculates the mining sequence over a designated time period or until the economic material is exhausted. A simulation model imitates step-by-step performance of equipment or other processes.

Jan 3, 1988

By J. Tuzson, H. McL. Clark

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

Jan 1, 2000

By Robert W. Miller, Rhoda S. Kriesel

INTRODUCTION With each passing year, the general public becomes increasingly aware of the potential hazards associated with many products and services previously considered unharmful. This new consciousness is especially evident in the workplace, both as the result of advances in scientific understanding and the influx of concerned, responsible professionals into the labor movement. For example, data produced by epidemiologists doing retrospective mortality studies which related mortality to toxic substance exposures has often been the prime motivating factor behind reductions in exposure standards. This certainly has been the historical case with many radiation exposure standards, including the limitations set for "Working Level" exposure. The concept of "Working Level" was introduced in PHS Publication No. 494 as a result of both the difficulty in relating radiation units to biological effects and the complexity of interaction of radon daughters with the physical environment. WORKING LEVEL One working level is defined as any combination of radon daughters in one liter of air that will alternately release 1.7 x 105 MEV of alpha energy during decay of 210pb (RaD). Its usefulness is that working level can be readily measured in both the field and the laboratory. Because of this, existing occupational exposure standards for radon daughters in both the U. S. and other countries use working level as their basis. In the U. S., the last twenty years has witnessed a significant reduction in radiation exposure standards from 120 working level months to 4 working level months per year. New international standards are being proposed that would be only slightly higher than these current U. S. standards. POTENTIAL EXPOSURES TO RADON DAUGHTERS The universal use of nuclear power, along with the transfer of nuclear technology to developing nations, has significantly increased the demand for fuels. As a result, uranium mining operations have expanded and increased numbers of miners are regularly exposed to radon daughters. In addition to potential exposures in uranium mines, other deep mines such as tin, gold, platinum and tungsten, also have potential exposures due to the distribution of uranium and its accompanying radium throughout much of the earth's crust. The energy crisis has also prompted interest in general population exposure to radon daughters in buildings. Many homeowners, using additional insulation and caulking to seal cracks and conserve energy, have reduced indoor ventilation to the point where working level exposure could become significant, especially in homes with unventilated crawl spaces. In fact, several serious exposure situations have occurred where mine tailings were used for building materials. MEASUREMENT TECHNIQUES Historically, rapid and convenient measurement techniques have been actively sought to improve the ability of ventilation engineers to limit mining exposures through ventilation control. A number of measurement techniques have been developed, including the Kusnetz method (further elaborated by Shalaynev) and techniques developed by Rolle and Tsivoglou. All, however, suffer from similar handicaps that prevent their usefulness in the mining environment. Their main limitation, in terms of minimizing employee exposure, is the elapsed time from the start of sampling until the results can be calculated. It is usually 40 to 90 minutes. This time delay presents both serious economic limitations as well as the potential for exposing workers to high radon daughter concentrations until calculations are completed and corrections made. Further reducing the utility of the Kusnetz and Tsivoglou techniques is the need for tedious calculations that increase the possibility of human error along with the need for cumbersome sampling equipment that is not ideally suited to the mining environment. INSTANT WORKING LEVEL METER To alleviate these problems, the idea of an Instant Working Level Meter was conceived by several groups. Several such instruments were proposed, built, and tested with disappointing results. Although the concepts behind them were sound, the instruments simply failed to measure working level without excessive distortion. In addition, their cost was relatively high, gamma background presented a problem, and sophisticated computations were necessary to determine working level. In short, these systems lacked the design engineering that would make them suitable for the demands of a mining environment.

Jan 1, 1981

By Glenn A. Shirey, Robert D. Stoessner, S. Gopalakrishnan

Under a contract awarded by the U.S. DOE-PETC, the Pennsylvania Electric Company and New York. State Electric and Gas Corporation are conducting a test program to evaluate and optimize the performance of a two-inch air-sparged. hydrocyclone (ASH) for cleaning fine coal. The test program is being conducted at the Homer City EPRI Coal Quality Development Center, utilizing Upper Freeport coal classified to minus 0.15 mm as the process feed material. A matrix of experiments has been conducted to establish the influence of significant design and operating variables (including overflow and underflow orifices, inlet slurry pressure, airflow rate, and reagent dosage) on the production of deep cleaned coal. Performance of the ASH is being compared to conventional froth flotation, utilizing advanced techniques and selected flotation reagents. This paper focuses on work completed to-date to optimize the performance of the air-sparged hydrocyclone for cleaning fine coal. INTRODUCTION The Pennsylvania Electric Company (Penelec) and New York State Electric and Gas Corporation (NYSEG) have been evaluating advanced fine coal cleaning processes for the production of compliance grade coal (1.2 pounds S02/million Btu). Laboratory bench-scale froth flotation results were compared with results of advanced flotation processes as performed by various organizations, utilizing the same coal feed. Among the processes evaluated, the air-sparged hydrocyclone (ASH) - developed by Dr. Jan Miller at the University of Utah - was considered to offer the best potential for retrofitting into an existing coal preparation facility because of the increased capacity per unit volume. The ASH has been shown to process over 100 times the throughput of conventional froth flotation cells. In May of 1988, Penelec/NYSEG was awarded a contract from the Department of Energy's Pittsburgh Energy and Technology Center (DOE-PETC) to optimize the performance of a five cm (two inch) ASH for producing compliance grade coal from raw classified minus 0.150 mm (1.00 mesh) Upper Freeport coal, and to compare its performance with conventional froth flotation utilizing advanced techniques and selected flotation reagents. The test work is being conducted at EPRI's Coal Quality Development Center (CQDC) in Homer City, Pennsylvania, utilizing a circuit processing 0.10 to 0.25 tons of coal per hour. Other members of the Penelec/NYSEG contract team include the University of Utah, Advanced Processing Technologies, Inc., Davy Dravo, Kaiser Engineers, Ebasco Services, and Management and Technical Systems. ASH DESCRIPTION AND OPERATION The ASH-2 System is a five cm ID (actually 4.76 cm) air-sparged hydrocyclone manufactured by Advanced Processing Technologies, Inc., under license from the University of Utah, for the purpose of research, evaluation and development. A schematic drawing of the air-sparged hydrocyclone is presented in Figure 1. Basically the air-sparged hydrocyclone consists of two concentric right-vertical tubes, a conventional cyclone header at the top, and a froth pedestal at the bottom. The inner tube is a porous tube through which air is sparged radially. The outer nonporous tube simply serves as an air jacket to provide for even distribution of air through the porous inner tube. The slurry is fed tangentially through the conventional cyclone header to develop a swirl flow in the radial direction. Air is sparged through the jacketed, inner porous tube wall and is sheared into small bubbles by the swirl flow. Hydrophobic particles in the slurry collide with these bubbles and, after attachment, are transported radially into a froth phase which forms on the cyclone axis. The froth phase is stabilized and constrained by the froth pedestal, thus moving towards the vortex finder of the cyclone header and discharged as an overflow product. Hydrophilic particles generally remain centrifuged in the slurry phase and are discharged as an underflow product through the annulus created by the froth pedestal.

Jan 1, 1989

By G. Sciocchetti, G. F. Clemente, F. Scacco

INTRODUCTION Radon, thoron and their airborne daughter products may result in a significant biological hazard when they are concentrated in enclosed places such as mines, houses and caves. High concentrations of radon and its particulate daughter have long, been known to be a causative agent for lung cancer to uranium miner: (Archer 1973). More recently increasing concern has been expressed on the possible health hazards associated with long term exposures over large population groups to low levels of radon daughters (Cohen 1980, Stranden 1980). Due to its peculiar geological features, Italy has various natural environments where the content of uranium and thorium in the soil may increase underground radioactivity. The presence in Italy of many underground waters rich in radon may be considered a further important source of radon in underground mines. A preliminary survey of the radon and radon daughter air concentrations in italian underground mines was started in 1971 (Bottino 1973). Furthermore many studies (Sciocchetti 1978, Sciocchetti 1980, Clemente 1980) have been performed during the last ten year to characterize the natural radiation environment in Italy. It has been also demonstrated that 21OPb/21OPo skeletal burden in spa workers and non-uranium miners can be heavily increased by the exposure (short or/and prolonged) to high levels of radon daughter air concentration (Clemente 1980). All these studies have been of great help to put in the right prospective the radiation hazards linked to the underground working activity in non-uranium italian mines. Two main actions have been prosecuted from 1971 until today in order to evaluate in greater detail such radiation hazards and then, when needed, to obtain their reduction to levels which may be considered acceptable. The two actions may schematically described as the following: - Measurement of the radon and daughter (and thoron and daughter when needed) concentrations in most of the italian non-uranium mines. The assessment of the radiation hazards linked to both the annual average lung exposure and the 210 Pb - 210 Po annual average dose to the skeleton is then performed for each considered mine by taking into account the main parameters characteristic of the mine. - Set up of a national regulatory recommendation on the permissible limits for the radiation exposure of workers due to underground mining activity. Such recommendation have been established in 1978 and have been designed in order to implement a national radiation protection practice during the underground mining activity. In this paper a summary description of the results obtained during the execution of these two actions will be given. MATERIAL AND METHODS Measurement of Radon, Thoron and their Daughter Products The excution of the planned survey requires the availability of specific techniques and instruments for measuring the different parameters characterizing the radon, thoron and their daughter exposure in mines. A reference laboratory was set up for the development and the standardization of the measurement procedures. Measurements of radon, thoron and daughter concentrations are performed by various methods: instanteneous or grab sampling methods and continuous read out monitors. The radon concentration is usually measured by a modified version of the two filter method. Mine air is drawn into a polyethylene bag and the sample is there after injected into a two filter tube; after sampling the exit filter is counted by the alpha spectrometric technique. The measurements is completed in a total time of less than 25 minutes at a sensitivity level as low as 3.7 Bq/m3 for a tube volume of 15 1. The measurement of radon and thoron daughter concentration and of the relevant potential alpha energy concentration is performed by means of a portable spectrometric apparatus which was developed at our laboratory (Sciocchetti 1981). The device allows multiple operating modes: instantaneous or brief sampling period and automatic cycles; computer data processing is then provided. The instruments that measure continuously radon concentration over long periods of time are of two types: a continuous scintillation monitor and an automated two filter

Jan 1, 1981

By C. Allen Born

There is a saying, the greater the gratitude, the shorter the speech. That puts me in a quandary, because while I am grateful for the honor you do me today, I also want to talk about a subject very important to all of us. A few months ago, when I was asked to submit a title for my talk, I chose "Quality where it counts." I did so because perhaps the most important thing we at Amax did in our restructuring was to commit to quality in all our businesses - quality people, assets, and operations. Together, they create quality performance, which produces quality growth and increased shareholder value - our prime corporate mission. We are dedicated to quality. At Amax, quality means everything. Today, I want to discuss quality in another context - the need for it in American education and what we have to do to achieve it. If we ignore this need, we will all suffer the consequences. And they will be upon us far sooner than the next century, barely a decade away. I think almost everyone in this room works for an organization that had to restructure in the 1980s. Mining companies that did not, just didn't make it. Well, we did. As an industry, we are seeing improved profits, positive cash flow, and a more competitive position in the world marketplace. But all the restructuring we have done, all the hard work, and all the pain that was part of it, will be in vain if, down the road, we do not have the educated people we need as workers. We need them to design, manufacture, and market the products that use the basic materials produced by the mining industry. Simply put, we are at a crisis point in education today. That is no new revelation. Nearly six years ago, the National Committee on Excellence in Education published its report, "A Nation At Risk." Since then, things have not gotten better. The Wall Street Journal, Business Week, US News & World Report, and Fortune have all recently published stories about the sorry state of education in the United States. The subject has become the topic of TV talk shows as well as national commissions, and for good reason. The pool of young people entering the US labor force is dwindling. In the next decade, 10 million fewer will enter the job market compared to the 1970s. But fewer of them will be qualified to do the jobs available to them. Here's why: • 20% of today's high school students drop out before they are 18 and another 20% graduate functionally illiterate; 40% of today's fourth graders think the world is flat; • 20% of today's sixth graders cannot locate the United States on a world map; • one of every three ninth graders cannot figure the change for a two-item meal - and I'm talking about a Big Mac and fries. These are not just problems of minority kids in big city ghettos. Fully 40% of all young working Americans cannot add up their lunch bill. Three out of every five college freshmen need some kind of remedial work. These statistics are frightening. Particularly so because, while the work force is becoming less qualified and our public education system is getting worse, our products and technologies are becoming more complex. We are approaching a time when we will have machines that recognize handwriting but people who cannot write. We will have technology to take voice commands, but people who do not know what they are talking about. The accelerating pace of technology is changing the nature of the job market. The fastest growing jobs now, and in the years ahead, are in professional, technical, and sales positions requiring the highest education and skill levels. Most of these jobs will require the use of a computer. Yet, 95% of today's graduating high school students have never laid a finger on a computer keyboard. While our education system is failing the students it is supposed to serve, other countries are doing far better. In a

Jan 1, 1990

By James N. Grassby

Metal production has traditionally been a cyclical industry. 1977 in the base metals industry has highlighted that fact with a steady decline in the market place for base metals. Corporate aggregate planning is based on good forecasting and when changing conditions make that forecasting inaccurate, alternate analysis of schedules must be accomplished quickly. The computer is the only medium powerful enough to answer management's questions within an acceptable time frame. Worldwide communications networks now make it possible for the many multinational mining companies to arrive at coordinated corporate production plans based on similar basic data from several producing divisions. The Ontario Div. of Inco Metals Co., a unit of Inco Ltd., encompassed during 1977 14 producing mines, 2 nonproducing mines, 1 nonproducing open pit, 5 flotation mills, I smelter, 2 nickel refineries, 1 copper refinery, and an iron ore recovery plant. Several new mines are under development. Mine production approximates 63 500 tpd (70,000 stpd) from some 1000 workplaces. The need for longrange mine planning and scheduling is obvious and until recently has been satisfied by manual methods. However, while manual methods are often quite practical for a small number of mines, or for mines with relatively short lives, or for mines with only one mining method, the situation in the Ontario Div. is too complex and long-term to allow effective, timely planning by such methods. In 1967, a year after the first computer was installed in the division, work was begun on the computerization of the more than 30,000 borehole logs kept in the mines exploration files with a view to speeding up the assessment of mineral resources. Also in 1967 one of the first substantial systems developed on the main business systems computer was the monthly scheduling of development and production for all mines. This system encompassed tons productions per workplace, feet development per workplace, auxiliary and service activities and the requisite labor for each of these. These two dates are mentioned to show that concurrently with the acquisition of computers in the division there was an early understanding of the importance of automating mine scheduling and planning and of the potential for effective and useful computerization of such activities. In 1970, a pilot program was developed in one of the mines containing a large number of workplaces (350) to schedule the production on a network basis. Varying degrees of interest were shown by the individual mine engineers based on the complexity of the production planning problem at each mine. In an evolutionary process, mines were added to the system as its value became obvious in terms of time saving, flexibility, and accuracy. In 1975, the central mine engineering department decided that the system should be extended to all mines so that a complete division-wide long-range scheduling system would be available upon which to base development and production planning. This system would use basic ore data, allow the addition of scheduling commands, and calculate a schedule for the life of the mine. Many reports could be produced for each mine and all-mine summaries would also be available. The significance of the system can be stated in one phrase-speed, accuracy, and flexibility. Whereas in the past a long-range all mines schedule might have been done once a year, taking from two to three months, now it can be done in a day. The chief mines engineer can incorporate changes in input data and external constraints and call for a "new look" at the long-range schedule to assist him and mines production management in their decision-making processes. Using previous manual methods this could not be done quickly enough to be useful.

Jan 1, 1979

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

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

Jan 9, 1988

By E. Haug, Quamrul Ahsan, R. Norcross

Cyanide leaching is the most common process employed in the majority of the gold extraction plants. These gold extraction plants naturally generate tailings and/or effluents that contain highly toxic cyanides, In geographic locations with excess precipitation and therefore net positive water balance, a portion of these tailings must be discharged. Before these highly toxic cyanide-containing effluents can be discharged to the waterways, cyanide and any other toxic substance must be destroyed. Geographic locations with high evaporation and low precipitation, as in the western USA, result in net negative water balance and therefore these plants can operate with a zero discharge option. However, with the current stringent environ­mental regulations and concerns for bird and wildlife mortality, even these operations may be required to partially treat their tailings to preclude potential harm to wildlife. Furthermore, heap leach opera­tions are now generally required to detoxify the heaps before abandonment. It is, there­fore, apparent that all gold extraction plants may be required to install some form of effluent treatment in order to achieve the above objectives,.

Jan 1, 1990

By Donald R. East

For many years there has been increasing pressure on mining companies, not without some justification, to design tailings storage facilities which guarantee some degree of protection against process solutions entering local groundwater and surface water flows. A fairly straightforward approach to achieving this has been to design conventional water-retaining type structures, and this has often been reinforced by the need to store large volumes of process water and seasonal run-off within the tailings facility. The conventional design uses naturally occurring geological barriers and a conventional embankment with possibly a grout curtain, or artificial barriers around the entire facility, using either imported low permeability soils, admixtures, or synthetic liners. While these approaches may achieve the primary objective of minimizing seepage, various other factors should be considered. Firstly, by depositing the tailings slurry in a sub-aqueous manner, the tailings, especially if relatively fine grained, will settle to very low densities and little or no on-going consolidation may occur over the operating life of the mine, There have been many cases where measured pore pressures within the tailings mass equal the total overburden stress, i.e., no consolidation and no strength, many years after decommissioning.

Jan 1, 1998

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By Sidney A. Hiscock

INTRODUCTION Lead and zinc are today rightly regarded as sister metals. Historically, however, they differ markedly with lead being known and in general use as the metal since 3000 BC in several countries which at that time might have been described as mature economies. Zinc was not isolated and recognised as an industrial metal with distinct and valuable properties until about 1700-1800 AD (although it had, of course, been utilised as a constituent of brass for some 2000 years previously). c were well established as industrial metals by the beginning of the 20th century. The present paper briefly reviews the overall growth in world consumption* and changes in the three main areas which today represent the mature economies or industrialised areas, i. e. ,Europe, the United States and Japan. Trends in overall consumption are determined by the demands for the individual uses for lead and zinc. The changes in patterns of use in the industrialised areas - individually and collectively - and reasons for the changes are also considered. LONG TERM TRENDS - 1900-1984 The World Picture Since the beginning of the 20th century the world consumption of lead (3.9 million tonnes in 1984) has grown almost fourfold and that of zinc (4.7 million tonnes in 1984) is almost eleven times greater than it was in 1900. The long-term trend has been one of continuing expansion for both. metals although the pattern has been severely disturbed at certain periods, for example in the years immediately following the two world wars and at times of marked recessions in industrialised areas. Apart from such major setbacks, new levels of consumption have been established regularly every few years. Recently, however, consumption has not really grown and with only a slow and partial recovery in world industrial activity at present, there is no indication that the peak consumptions of 4.2 million tonnes of lead, set in 1979, and 4.8 million tonnes of zinc (1973) will be exceeded in the near future. The growth in world lead and zinc consumption since 1900 is shown in Figure 1 (which includes changes in copper and aluminium for comparison). Overall growth over the period 1900-1984 has been about two percent a year for lead and about three percent for zinc (annual growth for copper has also been about three percent). Increases in tonnage consumption and growth rates by decade for lead and zinc are summarised in Table I. However, the growth rates by decade sometimes conceal very large individual annual increases and decreases. For example, in 1975 lead consumption fell by 13 percent and zinc consumption by 22 percent compared to 1974. In 1976 consumption of lead and zinc rose by 12 percent and 18 percent respectively. The years when consumption 'landmarks' (ie one, two, three and four million tonnes) were first reached are shown on Table 2. Clearly, the pattern of shorter periods being required to attain each extra million tonnes of consumption was broken in the 1970's. The consumption of zinc overtook that of lead for the first time in 1940, and since 1946 has always been at a higher level. Trends in industrialised areas The two key industrialised areas at the beginning of the 20th century were Europe and the United States which between them accounted for some 95 percent of world lead consumption and 98 percent of zinc consumption. Figures 2 and 3 show the growth in lead and zinc consumption, and Figures 4 and 5 the percentage shares for each area since 1900. Europe has usually been the major user of lead except for some years in the 1920fs, most of the 40's and early 50's when consumption in the United States was larger. In 1900, Europe accounted for about 64 percent of world lead consumption and the United States just over 30 percent. Currently, Europe takes about 40 percent of world consumption and the

Jan 1, 1986

By Geoff W. Seward, Charles J. Maes

INTRODUCTION The work of the metallurgical engineer engaged in developing a leach, solvent extraction electro- winning process usually begins in the laboratory and generally proceeds via pilot plant testing before finally resulting in a commercial operation. The path leading to the development of the eventual process will involve a test program on leaching in order to derive information on the amenability of the copper-bearing material to acid dissolution, under various conditions. Commonly in North America the leach process of choice is either dump or heap leaching but it may also be in-situ leaching or the treatment of mine waste- waters and bleed streams. In the past, agitation and vat leaching have also been included in the list of options. Currently the economics of copper production dictate that only material of mineralogy and grade, capable of being leached without prior milling, will lead to a viable process. This is not necessarily true in other parts of the world where grades are higher and milling costs lower and therefore this paper con- fronts all the techniques available to the metallurgical engineer in terms of the evaluation of the solvent extraction process. Understanding the objectives of a laboratory and pilot plant evaluation of a solvent extraction route is important and these are described in some detail in this paper. Of at least equal importance is understanding limitations of such a test program and this aspect of process development is also discussed with reference to the kind of equipment and experimental practices that have been shown to produce results consistent with the objectives. WHERE TO BEGIN? At the inception of a project much information is developed on the variables of the leach process and indeed in some instances the type of leach that best suits the ore to be treated is the subject of intense investigation (Wadsworth 1983). The decision to treat the leach solution by solvent extraction focuses attention on the selection of the extractant most suited to the nature of the chemical composition of the solution. A laboratory screening program will probably need to be arranged in order to assist in narrowing the options and the following is suggested for this initial screening. Selection of a Representative Leach Solution This is preferably obtained from leaching a representative sample of the ore under conditions likely to pertain in the commercial leach circuit. As the conceptualized circuit flow sheet will undoubtedly involve the recycle of aqueous solutions via the leach, then it is desirable to know the nature of and the likely level of the impurity ions in the solution as well as the concentration of copper. The other most important parameter that is required is the free acid (Anon, Acorga Analytical Method) content of the solution. As a rule of thumb, it is always true that recovery across SX is maximized as the pH of the leach solution rises. Whilst a high pH may be desirable for SX recovery, it is not always desirable for the pH (or more correctly, the free acid) of the leach to rise too high. The working range, therefore, tends to be between 1.5-2.5 pH. At an early stage in process development it is likely that a number of ill-defined conditions will exist, making it difficult to decide on typical leach conditions in which case the screening tests should be carried out using a synthetic leach solution prepared from laboratory chemicals. In doing this the copper and free content should be carefully controlled along with the appropriate addition of major impurity ions as their sulphate salts. Preparation of Suitable Organic Solutions Manufacturers information sheets on the reagents to be tested are required at this stage The information necessary to make an informed judgment on reagent performance is: Copper Uptake Value. This will indicate the amount of copper that can be extracted by a reagent as a function of its operating concentration in a

Jan 1, 1986

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By Kumara Rachamalla

North American mining companies are lagging behind their global competitors in participating in the outstanding opportunities in India. The Indian government has liberalized foreign equity participation in the mining sector by up to 50% and, in some cases, even higher. Delegates from Europe, North America and South Africa learned this at an information seminar held in London, England, Attendees were welcomed by L.M. Singhvi, the UK's high commissioner for India. He introduced a government of India delegation headed by B.P. Baishya, minister of steel and mines. Singhvi is an eminent jurist and leading constitutional expert. He reiterated the soundness of India's legal system. He also outlined the recent Investment Protection Treaty between India and the United Kingdom. Baishya emphasized thee geological diversity and strengths of India's domestic market with its population of more than 920 million people the second largest in the world after China and its reservoir of skilled labor. He also outlined the potential of India's untapped natural resources. The private sector is the backbone of the Indian economy. It accounts for 75% of gross domestic product (GDP). The current minimum program of the new United Front government envisions 12% growth in the industrial sector, 7% in GDP and direct foreign investment of US$10 billion a year. "Mining is an area that can attract a sizable part of this investment," Baishya said. "Projected growth of the Indian economy will require increasingly large quantities of basic raw materials, such as coal and base- and precious-metals to meet the needs of domestic and export markets." Administration of India's mining sector is divided into the Ministry or Mines for regulating and developing the country's mineral resources, five public sector Mining Enterprises, the Geological-Survey of India (GSI), the Indian Bureau of Mines (IBM)and 25 states and seven Union Territories. The GSI is the second oldest (founded in 1851) and the third largest organization of its kind in the world, Baishya said. It has geologically mapped more than 90% of India's 3.2 million kmz (1.2 million sq miles) at a scale of 1:50,000. Several promising mineral projects have emerged from regional exploration programs conducted by GSI and the Mines and Geology State Governments. IBM recently completed a national mineral inventory. It covers 13,000 deposits/prospects of 61 nonferrous minerals. GSI also compiled a similar inventory on 61 coal fields. India is attractive to exploration companies for several reasons. These include favorable geology, accessible locations and a large mineral database. India also has many experienced geoscientists with well-equipped and efficient laboratories, Baishya said. Secretary to the Ministry of Mines A.C. Sen emphasized the largely untapped-geological and mining potential of India. He also discussed the new vistas that have opened up opportunities for exploration and mining. India has large quantities of mineral reserves, Sen said. Its vast Precambrian Shield - like those in Canada and Australia - is endowed with gold, platinum group and base metals, as well as coal and industrial minerals. Annual mineral production is valued at more than US$7 billion. Sen pointed out that India is the largest single consumer of gold. And domestic gold prices command at least a 20% premium above international prices. Recent diamond, gold and base-metal discoveries and prospects uncovered by GSI have generated investment interest from abroad, he added Delegates heard that the Indian Constitution gives the central government the job of framing legislation and the regulation and development of minerals. This ensures that mineral laws are uniform throughout the country. However, the right to grant mineral concessions, such as prospecting licenses and mining leases, rests with the minerals' owner. In India's case, that is the state government. The Indian government has formulated several guidelines that regulate the granting of prospecting licenses for large areas. ? The central government will consider the requests of state governments for the granting of prospecting licenses for areas exceeding 25 kmz (9.6 sq miles). But the license must include a provision to conduct aerial prospecting of the area. ? Any prospecting licensing area should not exceed 5,000 kmz (1,930 sq miles). for a single license. And the total area held by one company should not exceed 10,000 km2 (3,861 sq miles) for the whole country. ? The grant of larger areas will be linked to a mini- mum expenditure commitment on physical targets. State governments will monitor these expenditures. ? The granting of large areas for prospecting will be linked to a schedule of relinquishment.

Jan 1, 1997

By Manfred Hager

The ability of conveyor belts to transport large mass throughputs economically over previously unprepared ground has resulted in this system achieving great and extensive use. A significant component in this development is the conveyor belt itself. The development of high strength steel cord conveyor belts involves the optimising of splice design, the use of excellent rubber material especially in the splice, good craftsmanship during splice production and modern field vulcanisation equipment. The durability of a splice for belts of class St 3000 up to St 790 is expressed by the fatigue strength under dynamic stress. The results obtained with a test method and equipment developed by the University of Hannover indicate the present state of the art in this field. Belts of high nominal strength used on inclined long distance conveyors reach splice fatigue strengths of about 3000 N/mm. The design should only exploit approximately 50% of the fatigue strength determined in the tests as maximum operating stress. IMPORTANCE OF CONVEYOR BELT SYSTEMS The use of conveyor belt technology has expanded considerably over the last decades in the bulk goods transportation sector. Because of the favourable transport costs and the technology's adaptability to specific topographies, belt conveyors frequently represent the most economic solution. When flows of goods are large, as in German lignite mining operations, where masses of up to 40,000 t/h are given, it is the only technically feasible alternative. The use of belt conveyor systems allows the sensible distribution of the waste and the production of the coal (Hager, 1981). But this transportation alternative is also suitable when the masses involved are smaller. Whereas conveyor belts were only suitable for loose bulk goods up until a few years ago, today they are frequently also used for hard rock open cast mines, e.g. in the production of copper ore. Because of the favourable transport costs it is frequently still economic to use mobile crushers with a throughput of up to 10,000 t h 6 which breakdown the large blocks produced y explosives into transportable grain sizes. In many operations, trucks are only used to provide flexibility between the excavator and the crusher, i.e. over short routes, and the long, frequently steep, transport paths to the processing plant or to the spreader are undertaken by belt conveyors (Einenkel et al., 1992). The advantages of belt conveyors, i.e. to cope with large mass flows over inclines and down slopes of up to 1:4 have resulted in this system gaining very extensive use throughout the world. The conveyors can be installed over for the most part unprepared ground with suitable vertical radii, adapted to the locally available belt material. In the process individual conveyor belts with lengths of up to 15 km, and in the underground sector with lift heights of up to 1 km, have been built and operated. This great variety of application possibilities is complemented by a limited ability to pass through horizontal curves. This can be achieved with the help of design measures on the conveyor whilst bearing in mind the characteristics of the belt. STRESSING OF CONVEYOR BELTS IN DIFFERENT TYPES OF PLANT The advantage of conveyor belts compared with other systems is also to be found in the large available service-time window, i.e. the excellent reliability and the low costs of energy consumption and maintenance. It is in particular the reliability of a belt transport system which depends to a great degree on its main component, that is the belt - in all its different types and variations. For this reason, the belt is also given the greatest attention in the development of the individual components, because it is the belt which must be designed optimally with regard to various factors. The specific properties of the conveyor belt then assert a considerable influence on the design and sizing of most other components in a

Jan 1, 1993