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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 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 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 Andrew N. Peterson

US mine operators frequently fail to investigate more cost effective and productive bulk material handling systems because surface mines seem to lend themselves to truck ore haulage. In this country, as a result, use of conveyors to move heavy loads from mine to process facilities has been minimized, if not actually neglected. In contrast, there are more than 50 conveyorized surface mines in successful operation around the world. These mine operators have learned that properly applied conveyorized systems can offer major savings in capital and operating costs, which contribute to improved profits when combined with other proven mining technologies. Growing acceptance and application of conveyorized bulk material handling in surface mines also points up how unique each mine is and how careful planning contributes to maximum mine effectiveness. Because of these differences, mining executives and technical and operating staffs need to develop an understanding of three factors in applying conveyorized bulk material handling in surface mines: • Why each mine will benefit from the type of automation permitted by conveyorized operation, •What kind of equipment is available, and • What applications most effectively demonstrate the first two factors in action - hauling either ore or waste. The conveyorized systems considered in this presentation have production rates from 0.5-2.7 kt/h (500-3,000 stph). Worldwide, these systems have been operating since the early 1960s. Advantages of Conveyors Why do you want conveyorized bulk material handling? First, it almost always provides lower operating and maintenance costs. Second, it frequently requires lower initial capital costs and almost always requires lower capital costs over the life of the surface mine. Third, it provides comparable operating availability, and finally, it frequently gives comparable operating flexibility - depending on the mine plan. Cost avoidance can be accomplished with modern production methods. These, in turn, permit increased productivity and reduced operating costs such as those for energy, maintenance, and manpower. It has been demonstrated in European surface mines and elsewhere, that conveyor systems frequently require lower initial costs than does truck haulage. Almost always such operations require lower capital costs over the mine life. Those costs include the continual addition of haulage trucks to both accommodate the increasingly difficult haulage routes and fulfill replacement requirements when trucks wear out. Conveyor systems handling ore in numerous large crushing and port facilities, which have operated since the early 1950s, have clearly demonstrated a useful conveyor life of more than 25 years. In contrast, off-highway trucks have life spans of six to eight years. The following examples illustrate comparative capital costs to purchase conveyor systems and comparable truck haulage units. Example 1 The ore haulage route from point A to point B is level and 610m (2,000 ft) long. The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be transported at a rate of 1.8 kt/h (2,000 stph). The installed capital costs to provide a properly designed conveyor that will transport the described material from point A to B is about $450,000. The capital cost to purchase three 77-t (85-st) off-highway trucks and one spare truck - which would provide equivalent capacity - would be about $1.2 million. The truck cost estimate is based on a 6 min. or 771 kt/h (850 stph) truck cycle time. Truck efficiency is estimated at 0.8. Each 77-t (85-st) truck would have an actual haulage rate of 617 kt/h (680 stph). Therefore, three trucks would be necessary to transport the designated tonnage of 1.8 kt/h (2,000 stph). A movable crushing plant would be located at point A for the conveyors and a permanent crushing plant at point B for the truck haulage system. Capital costs for these primary crushing plants were not included in the calculations for either system because the capital costs are frequently comparable. Example 2 The transport route from point A to point B is 610 m (2,000 ft) horizontally and 122 m (400 ft) vertically - on a 20% grade (Fig. 1). The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be moved at a rate of 1.8 kt/h (2,000 stph).

Jan 6, 1983

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By H. Boyd

Fluid energy grinding, or jet pulverization, refers to the process of entraining solid particles in high velocity gaseous or vaporous streams and reducing their size by impact and attrition against other particles or some sort of target. Since all desired size reduction is not achieved in one pass through the zone where energy is applied, some form of classification is necessary to return oversize particles to the grinding zone for further size reduction. Applications Fluid energy mills have normally been used to produce products where the particle size is basically finer than 325 mesh. Products having particle size distributions with a maximum size of 4 or 5 µm, and with a high percentage finer than I µm, can be produced using fluid energy mills. More recently, some fluid energy grinding units have been applied economically to produce ground products with 1 or 2% coarser than 200 mesh. It is also possible to consider usage of a jet-pulverizing zone, without the usual integrated air classi¬fier, operated either in open circuit or in conjunction with a screen. Feed sizes to be delivered to the fluid energy mill depend on the material involved, the type of jet pulverizer used, the finished product size requirement, and the upstream process. In some units feed sizes up to l in. or larger can be accepted, but practical economics usually require the starting material to be -4 mesh or smaller and some pulverizers are typically fed at 20 mesh top size. It is not unusual to find feed materials at essentially -325 mesh when extremely fine products are needed, since capacities are improved with finer feeds. Conventionally the fluid energy grinding process has been used for relatively soft and nonabrasive materials. This includes very low¬melting-point and temperature-sensitive materials or coatings, since there is no measurable temperature rise during fluid energy milling. More recently, however, jet pulverizers are being used on highly abra¬sive materials such as silica, feldspar, glass, zircon, calcined alumina, coke, boron carbide, and similar materials. Since the primary energy application is by impact of particles against themselves, there is a low rate of wear resulting in a small amount of product contamination and in low maintenance costs. Usually the fluid energy pulverizer produces a more narrow particle size distribution than other mills provide, and this is useful where such a requirement exists. It is possible, within limits, to adjust classifi¬cation and other functions of the overall fluid energy system to vary the particle size distribution curve and thereby approach the broader distribution obtained from other grinding methods. By controlling the temperature, composition, and humidity of the fluid used, it is possible to incorporate other functions with the size-reduction process. These include drying or adding moisture, coating, mixing and blend¬ing, and even oxidation or reduction and other chemical reactions or their inhibition. The finished product is discharged in the fluid stream and is conveyed pneumatically to the desired point, where a suitable dust collector must be provided. Operation and Mechanical Features One of the earliest fluid energy mills is the Micronizer now manu¬factured by the Sturtevant Mill Co. Similar units are also manufac¬tured by Jet Pulverizer Co. and others. As shown in Fig. 81, the micronizer consists of a shallow circular grinding chamber with noz¬zles arranged at the outer periphery to discharge tangentially inward. Fluid (usually compressed air or superheated steam) is provided to the nozzles from a manifold at pressures normally in the range of 100-150 psig and at temperatures up to 800°F or higher, depending on the raw material, required results, and economic factors. The com¬pressible fluid expands to high velocity through the nozzles and gener¬ates a spiral flow pattern in the grinding chamber, and discharges

Jan 1, 1985

By A. K. Mukhopadhyay, B. Sikdar, V. P. Arora

Introduction During the past several years, persistent concerns about an economic slowdown in India have tempered capital-expan¬sion plans. This has encouraged companies to find methods of increasing the capacity of existing production equipment. One of the more cost-effective methods of increasing produc¬tion capacity without major capital investment is to improve equipment maintenance to prevent needless machine down¬time. This can have a significant financial impact because of reduced maintenance costs and increased machine availabil¬ity. The maintenance program employs different techniques, but the analysis of the vibration behavior of machine compo¬nents to assess their condition was found to be the most accurate technique for monitoring various machine disorders. A dragline that handles overburden in open pit mines is one of the more capital-intensive pieces of production equip¬ment. It is imperative that the machine be run at its optimum capacity under the prevailing environmental conditions. Condition monitoring of the various components of the dragline through vibration analysis not only improves the machines availability but also ensures a long operation time free from sudden failure. This paper describes the application of vibration monitor¬ing and analysis techniques to assess the components of a W2000 dragline. The paper reveals that condition monitoring through vibration analysis can be as valuable in failure prevention as it is in failure analysis. Finally, the paper presents an approach to a knowledge-based expert system to the quick and correct diagnosis of mechanical faults in a dragline. Equipment A dragline is an intermittent-discharge-type excavator having a long boom at one end. Unlike other excavators, the bucket is not rigidly held by the frame structure. In favorable geomining conditions, the intensive operation of these exca¬vators is the least-expensive means of overburden removal. The basic working elements of this equipment are the hoist system, the crowd and retract system, the wing mechanism and the propel mechanism. The walking dragline is accepted as the most economical machine for stripping overburden at depths of up to 60 m (200 ft). They are electrically powered and the various motions are effected by individual drives. The Ward-Leonard system is used to provide the optimum flexibility in controlling different dc-drive motors. To mini¬mize the cost of mineral extraction, the largest such machine in India has a 24-m3 (850-cu ft) bucket with a boom length of 96 m (315 ft). Draglines are very costly, and considerable financial penal¬ties may arise from sudden stoppages due to component fail¬ures. Thus, operational reliability is of the greatest importance. Active measures are now being taken to minimize downtime by adopting condition-based maintenance techniques. Though a number of techniques come under this scheme, it has been observed that assessing the health of the dragline by measuring vibration is the best method of determining the dynamic behavior of its various rotating components to avoid any catastrophic failure. Based on criticality, the vibration-behav¬ior of the following components are studied: • motor-generator unit, • exciter unit, • hoist motor-gearbox unit, • hoist drum-gearbox unit, • drag motor-gearbox unit, • drag drum-gearbox unit and • swing motor unit The approach Increases in the vibration levels in different subsystems of the dragline were observed due to the presence of faults such as unbalance, misalignment, looseness and bad bearings. A vibration behavior study included the measurement of vibra¬tion level, trend analysis of time-domain vibration signals and vibration frequency analysis at bearing points of the components. A vibration meter, a vibration analyzer and a data collector are used. Vibration-level measurement. When the dragline was assembled and put into operation for the first time, vibration data were collected to generate baseline data for the equip¬ment (based on International Standard ISO 2372). After¬wards, the vibration levels were recorded regularly using a vibration meter. The measurements were taken at all bearing

Jan 1, 2000

By Robert F. Reeves, Stephens A. Avary

Introduction People, systems, methods, and organization make up maintenance productivity. Methods promote efficiency in operations, systems ensure consistency and continuity, and organization maintains direction and control. The common denominator in these techniques, though, is people. Management systems can be technically correct, but fail to work because management systems do not work by themselves - people make them work. Therefore, the foundation for mine maintenance productivity must begin with the people involved. Traditionally, the emphasis placed on mine maintenance, as opposed to production, is not proportional to the impact that maintenance has on a mining operation. The performance and productivity of mine maintenance substantially affects the operating costs, revenues, and profits of the mine. The mining industry has recently experienced a period of economic depression. While it has created financial problems in all segments of the industry, it has caused the industry to focus on the need for improved productivity. Mining companies that have made investments in productivity during this period will have a competitive edge during the recovery, and mine maintenance may be the best place to put that investment. Future lost production and maintenance cost problems can be avoided if mine management has the foresight to seize the opportunities that now exist. Background Technological improvements over the last 30 years have been dramatic in the industry. Before these changes, mining equipment was not very sophisticated. So maintenance needs were not very sophisticated. Maintenance people were not required to have any formal training in mechanical or electrical skills. At many mines there was no formal maintenance organization. The mechanic was a part of the section crew and received direction from the face boss. Mine maintenance was not a function, it was a task. Technological change in mining came about in response to an expanding economy and greater demand for mined products. The past 20 years brought about a change in the kind of individual needed to mine ore and coal, and an even greater change in the person needed to maintain equipment. Equipment went from mechanical and pneumatic to complicated electrical and hydraulic. Maintenance progressed from being a task to being a function. Types of Mechanics Mine maintenance workers can be grouped into three categories based on experience: mechanics with more than 20 years, five- to 10-years, and less then five years. These groupings have influenced mine maintenance productivity. The 20-plus mechanic has been the foundation of the maintenance force. He had no formal training and his skills were acquired on the job. As new equipment was introduced, he learned to repair it by trial and error. Not all of what he learned was correct. Personal initiative and conscientiousness are characteristics of this group. They have experienced the bad times and the good, and appreciate the opportunity to work. The five- to 10-year group worked with the older, more experienced mechanics to acquire maintenance skills. These mechanics came in during industry boom times, and many progressed quickly by being in the right place at the right time. Many in this group have difficulty with diagnostic trouble-shooting, and depend on changing out parts to identify problems. This group is generally more mobile and less apt to have a strong work ethic. Mechanics with about five years experience have vocational training, some formal education or basic skills training. Like the five- to 10-year group, these miners are mobile and tend to move readily to satisfy their changing needs. These two younger groups are less apt to feel any loyalty to the employer or obligation to be productive. Their attitudes and abilities reflect the influence of their general age groups, and has a definite impact on mining productivity. Past Requirements In the 1960s, the basic requirements for a maintenance supervisor was that he be a top mechanic, get along well with the men, and be a hustler and improviser. He learned the details of new equipment as repairs were needed. Oftentimes, this individual would be the only person on the property with the ability to read prints, use diagnostic testing equipment, and troubleshoot using an analytical approach. This approach to maintenance is sufficient when the equipment design is reasonably uncomplicated. Management Changes As technology progressed, organized labor gained strength, government regulation increased, and the job of supervising and managing people had to change. The environment has changed and

Jan 11, 1983

By L. W. Pommier

In the field of minerals processing the concentration of valuable minerals normally takes place in suspensions of ore particles in liquids. The fluid resistance offered by Newtonian liquids to the movement of free settling spheres for the limiting cases of creep flow and turbulent flow is well described by Stokes' and Newton's Laws. However, these equations are not applicable to non-Newtonian fluids, since the relationship between applied shearing stress and resulting rate of shear is a non-linear function, and more than one viscosity parameter is necessary to describe their rheological properties. For this reason, the term "apparent viscosity," which has been commonly used to explain the flow resistance of the non-Newtonian systems including suspensions, may be misleading. A rigorous analysis of the mechanisms of momentum transport of solid-liquid systems is complex indeed. There are many variables that affect the properties of such systems, and precise limitations and assumptions are necessary in order to produce quantitative relationships for predicting their behavior. Very dilute suspensions of spherical particles in Newtonian liquids of identical density as the solids have been examined in the past (12); however, the study of suspensions in which there is a considerable difference in density, between the liquid and the suspended solids has been neglected.

Jan 1, 1968

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 Raúl Castro, Marco Vera, Alfonso Ovalle

The underground mining evolution for block cave mines in Chile and worldwide has made important strides in the last decades. Nonetheless, reserves determination in the initial stages of these types of projects—mainly for long-term planning—is not studied in as much detail as it should be, especially regarding the impact on the final value of the enterprise. Today, the first estimation of the size of a block cave mine, typically expressed in tons, tends to determine very large footprints and ore bodies. The resulting large-scale mine is obtained by applying the marginal cut-off grade (COG) to determine reserves. Yet, in many cases, it is possible to obtain an equal or higher benefit, extracting fewer reserves in less time, with a smaller active area, which significantly reduces the project’s uncertainty. To explore scenarios obtained by applying a higher marginal COG in the initial phases of the project, the Hill of Value methodology is used to estimate the net present value (NPV) in relation to the production rate and the COG. However, the application of this methodology considers certain simplifications that generate results which are not completely satisfactory in specific cases. This paper suggests an improvement of this method to obtain more reliable results regarding determination of a COG profile that will optimize the NPVof a block-cave project. The methodology was developed and applied in a massive polymetallic deposit with zinc, lead, and silver mineralization, corresponding to a real block caving project. The methodology provides NPV results that differ by approximately 15% in comparison to the real mine plan, while also being considerably simpler and faster to calculate. In summary, the proposed methodology provides a reasonable and fast procedure to estimate higher than marginal COG that increases the NPV of block cave mine projects.

By Robin G. Adams

The market prices received by producers of internationally traded commodity metals have been, and continue to be remarkably volatile and unpredictable compared with most other traded goods and services. This volatility results in an erratic financial performance on the part of producing companies, disruptive and often inflationary cost changes for consumers, unreliable tax revenues and export earnings for host governments and severe social and economic pressures on local, sometimes remote, communities which are dependent on the activity in question. Generally speaking, price volatility appears to be associated with the inability of the metal producing side of the industry to regulate the volume of supply in a manner that is consistent with metal demand, which in turn reflects broad macroeconomic and industrial developments around the world. This statement does not purport to apply to all metals or to all instances of volatility of course. Price movements may occur as a result of pure speculation or even illegal manipulation of the commodity markets. Embargoes, boycotts and other politically induced supply changes may contribute. Threatened or actual strikes also cause artificial swings in sales volumes and prices. How- ever, most of these developments are of a short-term nature. Yet metals price cycles may last for several years, as demonstrated by the extended copper price trough from 1981 to 1986. This paper deals with price volatility which reflects long-term economic cycles and the adaptive processes that exist within the industry in response to such cycles. Special emphasis is placed on the role of forecasts and forecasters in these processes. It should be said at the outset that this reflects the personal experience of one forecaster working in private sector commercial consulting practices that are independent of producers, consumers and host governments. However, the experience does cover the entire period since 1973 and has involved working with a substantial majority of the major companies that supply high volume commodity metals. By way of background it should be stated that the post-1973 world economy has provided a challenging environment for forecasters. One cannot be precise about the forecasting record of either industry insiders or outside consultants, financial analysts and academic forecasters because comprehensive records have not been kept, or if they have the results are too embarrassing to be published. Institutions such as the World Bank have published forecasts of the prices of major commodities on a regular basis and these have, presumably, had some influence on financing decisions in the industry. This record along with our own experience and numerous discussions with metal producers, major consumers and host governments permits certain generalizations to be made. The most important of these is that during the early 1980s. from about 1980 through 1985/6, the forecasters system- atically and substantially over- estimated metal prices. While some fore- casters did correctly anticipate a recession in the early 1980s and some consequential decline in prices, the magnitude and duration of the decline was far greater than anyone foresaw. This systematic forecasting "failure" appears to have occurred independently of the forecasting methodology employed. Moreover, its extent cannot be explained by obvious factors such as incorrect macroeconomic assumptions. There appears to have been some kind of structural change in the world mining and metals industry over this period

Jan 1, 1989

By R. A. Holmes

Although manganese is a metallic element and is widely dispersed in nature, it never occurs except as a compound in combi¬nation with other elements. Use of such compounds in the production of glass is known to have occurred in early Egypt. The dioxide of manganese was considered a compound of iron until 1774 when C.W. Schule first recognized it as an element. In the same year, a Swedish mining engineer, J.G. Gahn, became the first to isolate manganese. In 1856 development of the Bessemer process of steelmaking gave economic importance to manganese, and later, in 1882 Robert Hadfield discovered the benefits of high manganese steels. GEOLOGY Physical Properties Elemental manganese is a silver-gray metal, resembling iron but harder and more brittle and used primarily in alloys (both ferrous and non-ferrous) and a wide variety of chemical compounds. Some of the physical properties of manganese include: melting point¬1 245°C; boiling point-2 150°C; density at 20°C-7.43 g/cm3; specific heat at 25.2°C-O.115 cal/g; latent heat of fusion-63.7 cal/g; hardness on Mohs scale-5.0; linear coefficient of thermal expansion from 0 to 100°C-22 x 10-6. Mineralogy There are over one hundred minerals that contain manganese. These minerals vary from those with compositions that are pre¬dominantly manganese to those having only minor percentages of manganese. Distribution of Deposits Manganese ore deposits are found worldwide and were formed in various geological environments, but only a rather limited num¬ber of deposits have high grade manganese ore in sufficient quan¬tities to be mined and utilized economically on an industrial scale. It is worth noting the fact that almost all of the significant deposits can be classified into two types of deposits: marine chemical sed¬imentary deposits, and residual (secondary) enrichment deposits. There are, however, a much larger number of geological types but these are not of commercial significance at this time. Sedimentary deposits are the most common and are usually stratiform or lenticular. Manganese minerals were formed by a chemical process during the deposition of marine sediments. They usually contain manganese oxides and carbonate minerals, some¬times interbedded together or with other sedimentary rocks such as limestone or shale. Examples of this type of deposit are the Russian ore bodies of Nikopol and Tchiatoura, as well as the Kalahari deposits in South Africa and deposits of Groote Elyandt in Aus¬tralia. Residual deposits were formed in a different way: by alteration of existing manganese deposits or by concentration of the manga¬nese minerals when other minerals were washed away by weath¬ering or ground water processes. The Nsuta deposit in Ghana, the Amapa deposit in Brazil, the Moanda deposit in Gabon, and nodules in the residual clays of the US Southern Appalachians are examples of this type of geological process. In the case of the Ghana and the Amapa deposits, this is only true for the outer layers of the deposit containing oxide minerals, the inner part being comprised of car¬bonate minerals including manganese carbonate, probably from marine origin. Some sedimentary and residual-type deposits have been metamorphosed, giving rise to small high grade ore bodies. These deposits are regionally metamorphosed, occurring in mar¬bles, slates, quartzites, schists, and gneisses. Some of these deposits, such as the Franklin, NJ, deposit are rich enough to be commercial without secondary enrichment; however, most of the exploitable deposits have been secondarily enriched. Due to the diversity and complexity of manganese deposits, both with respect to deposition and chemistry, a wide range of impurities are almost invariably present in the ores. Table 1 shows that the reserves (i.e., a measured resource that can be economically and legally extracted) are estimated at 814 x 106 tons of contained Mn, which equates to more than a 100 year supply at the current level of production. The reserve base is made up of the marginally economic reserves and sub-economic re¬sources and are some 4.5 times greater than the proven reserves. Also apparent from study of Table I is the fact that some 75% of manganese reserves are found in two countries, the USSR and South Africa. On the other hand, North America (i.e., the United States and Canada) have few significant deposits. The largest deposit, or at least one of the largest, in the United States is located at Chamberlain, SD. This deposit is sedimentary

Jan 1, 1994

By S. Budirsky

J.B. Gwiazda's article deals with an interesting problem that has not been studied thoroughly up to now. Gwiazda has proposed a technical solution that eliminates horizontal load imposed by the supports on the roof. That load is due to a discrepancy between the trajectories of the canopy and the overlying strata. It is true that the improved lemniscate guidance system proposed by Gwiazda could eliminate the horizontal load but, on the other hand, it would represent a higher cost of the supports, and it would complicate the already complicated mechanism of powered supports. That is why we must be convinced that the proposed improvement is actually needed. From that point of view, the following questions should be elucidated. As shown in Fig. 1, the point O travels along the curve a, which produces horizontal load on the roof in the direction of the coal face. As stated by Gwiazda, "such a high load is capable of destroying the roof above the support, causing rock debris to be scattered around the face." The results of our measurements carried out in coal mines of the Czechoslovak part of the Upper Silesian Coal Basin show that the convergence in the modern type of powered supports does not exceed several millimeters (not more than 10) per hour. Similar conclusions were drawn from field measurements performed in the Polish part of the Upper Silesian Coal Basin (see S. Romanowicz and H. Szopka in Proceedings, Scientific and Technical Symposium Simmex '85, Katowice, Poland, pp. 73 -83). The small vertical closure of the supports also induces a small horizontal displacement of the canopy. In addition to this, the transmission of forces between the roof and the canopy is influenced by the layer of debris on it, i.e., the canopy slides without imposing a greater horizontal load on the roof. This was deduced from field measurements of Voest-Alpine F 4/4600 powered supports equipped with rams instead of the common front lemniscate links (denoted as 1 in Fig. 5). In these supports, the pressurizing of the space of piston rod (la in Fig. 5) causes the rams to close, which induces horizontal load on the roof in a direction toward the coal face. After the setting of a support unit, the pressure in the piston rod space was found to drop, frequently to zero, because the small horizontal sliding between the canopy and the debris or inside the debris caused the horizontal reaction from the roof to fall to zero, although the debris were pressurized by vertical setting load. As a consequence, there was no horizontal load imposed by the canopy on the roof. A rise of the horizontal load was recorded only after the passage of the shearer and during lowering of the adjacent units. Thorough information on the results of the afore- mentioned field measurement is given in the paper, "Analysis of the performance of shield powered supports installed in a thick seam," published in Mechanizacja i automatyzacja gornictwa, 1982, No. 12, pp. 38-46. An English translation of the paper is available from the author. We have deduced from our observations that horizontal load imposed by the canopy on the roof in the direction toward the coal face is useful for roof control because it limits the displacement of the immediate roof toward the gob particularly during the rise working. We came to the conclusion that horizontal load on the roof toward the coal face should be induced on purpose. Could Mr. Gwiazda prove with the results of field measurements his contrary opinion? Has the small horizontal displacement of the canopy toward the face actually had an adverse effect on the roof? As shown in Fig. 2, during the setting of a support unit, the canopy imposes a horizontal load on the roof in the direction of the gob. In this case, I share Gwiazda's opinion that the horizontal load could have an adverse effect on the roof. Nevertheless, two important circumstances were neglected - the horizontal compliance of the supports and the interaction of adjacent units. The question of the compliance was analyzed precisely by I. Krumnacker in Gluckauf -Forschungshefle, 1984, NO. 5, pp. 219-223. He found that due to the clearance between the hinge pins and the eyes of the sheild and due to the elasticity of the steel structure,

Jan 1, 1988

By N. Edmonson

Introduction An examination of statistics for the world phosphate fertilizer industry since 1980 reveals that there is an increasing tendency for fertilizer-manufacturing activity to concentrate in the areas of phosphate-rock mining. The economic justification for this resource-oriented strategy is clear. The choice between shipping phosphate rock, a comparatively low-value product, and shipping a more valuable finished fertilizer product will always, as long as cost considerations control the decision, be resolved in favor of the latter. The historical existence of substantial phosphate fertilizer-manufacturing capacity, which is oriented toward fertilizer-consuming activities and dependent on imported phosphate rock, shows that narrow cost considerations have not always determined the industry's location. One is entitled to ask what conditions have changed to precipitate the observed relocations, and how permanent are these changes. These questions are the subject of this investigation. Recent trends in world phosphate trade Since 1980, the changing pattern of world trade in phosphate materials reflects an increased interest in manufacturing phosphate fertilizer in rock-producing areas. This is shown by the decline in the total tonnage of phosphate rock being traded worldwide compared to world consumption of phosphate fertilizers (Fig. 1). This relative decline has continued since 1984, which is a longer period of time than could be explained by temporary cyclical market factors. Other evidence corroborates this conclusion. The world's phosphate-rock production is relatively concentrated geographically. In 1992, the top ten producing countries accounted for 92% of world production. Figure 2 shows that the share of world fertilizer production in the top ten rock producers has increased. Yet, further evidence of the direction of trade changes can be had by looking at the pattern of the world's sulfur trade. The first step in the manufacture of any of the widely used phosphate fertilizer products is the acidulation of phosphate rock with sulfuric acid. In most instances, phosphate fertilizer producers have found it economical to produce their own sulfuric acid by burning elemental sulfur, and such operations account for over 60% of all sulfur consumption in the world. 1 Therefore, a concentration of phosphate fertilizer production in rock-producing countries should produce a similar concentration of sulfur imports in the same countries. Figure 3 suggests that there has been a tendency since 1989 for sulfur exports to go increasingly to the rockproducing countries. World phosphate-location strategy is not limited to the alternatives of shipping rock and shipping finished fertilizers. A third alternative consists of producing phosphoric acid, the principal phosphate fertilizer intermediate, at the rock deposit, and then exporting it to market-oriented ammonium phosphate plants. This arrangement offers the importing country the advantages of shipping a more concentrated product while retaining the economic multiplier benefits of a domestic high-analysis fertilizer industry. India, and several other countries, imports merchant-grade phosphoric acid to use as input for the domestic value-added fertilizer industries. For various reasons, the trade in phosphoric acid has never quite lived up to the promise suggested by the foregoing discussion. Figure 4 shows no growth in this trade in the last ten years, either in absolute terms or in relation to the world's consumption of phosphoric acid. Most of the decline in the trade in phosphorous in rock form has been replaced with trade in finished fertilizers. The trade in superphosphoric acid was a growth activity in the late 1970s with the growth in US exports to the USSR, and this trade survived the Carter-era embargoes of US-Soviet trade. However, no additional sizable markets for superphosphoric acid developed, and other suppliers sought out the Soviet market. The problem with superphosphoric acid is that the cost of its incremental concentration is as great as the savings from shipping the more concentrated product. The Soviet market was a special case. The political breakup of the Soviet Union disrupted the superphosphoric acid trade after 1991 and led to a shutdown of the US export capacity.

Jan 1, 1996

By D. H. Steve Zou, Cui Lin

"This paper first discusses the basic attributes of sustainable mining and the responsibilities of various steak holders including mining companies, governments, and engineers. It is followed by a brief overview of the current practice on regulations, inspections, waste management and mine closure planning in Canadian mining industry. A case study is then presented as an example of sustainable mining where mine planning, mineral recovery, environment protection, reclamation and community development were integral parts of the overall strategy. A comparison of the mine site before and after mining is made to demonstrate the ""best practice"" in modern mining.Sustainability in Mineral Resource Development Extraction of mineral resources and other natural resources on the Earth has been and will continue to be an essential activity necessary to support human’s life. We live on the planet and depend on the available resources to survive. The extracted materials form the base to support modern life. However, we must be aware of the fact that mineral resources are limited on the Earth and may not last forever. Every time we extract one ton of materials, there is one ton less. They are not able to regenerate or regrow. One can ask a simple question: How long will they last? Or, is there anything left for our great grandchildren and many generations to come? The answer is not simple. Some minerals have become scarce or about to deplete. Some minerals may last longer. The reserve of a type of minerals depends on what we have discovered and the minable amount depends on accessibility and technology we have. Even if there are minerals in the deep ground, we may be limited by our technological capability. At the same time, we also need to understand the action of mineral extraction from the ground and the consequences of that action. Mineral extraction is a very hard work. We must break the ground to extract the ore, either through surface mining or underground mining. This action means disturbance of the land where we live. The disturbance is much more visible in surface mining than underground mining. The disturbed land, if not properly reclaimed after mining, is not going to be reusable. The very purpose of extracting minerals from the ground is to support our life but the action has a negative impact on our life. How can we reduce and limit disturbance of land from the action of minerals extraction? How can we make the disturbed land reusable? These issues must be properly addressed and a balance achieved."

Jan 1, 2017

By J. M. Melluish

The metallurgical response of the Point Leamington deposit has been studied. Two series of bench tests were conducted and various ore samples and flotation products were examined using image analysis techniques. The majority of sphalerite and chalcopyrite grains in this texturally complex ore were found to be smaller than 10 micrometers. Copper recoveries were low with losses mainly as fine inclusions in pyrite. Sphalerite recoveries were somewhat better due to a generally coarser habit. Gold and silver recoveries were also low and only solid solution silver was identified in the ore. The deposit shows poor economic prospects unless additional tonnage of high grade zinc ore can be delineated.

Jan 1, 1989

By H. H. Richards, Ray L. Tobie, L. A. Thomas

GENERAL DESCRIPTION Since the beginning of operations, with the exception of a small tonnage mined by slushing, ore extraction has been by full gravity caving. Formerly, a checkerboard sequence of block undercutting was followed with the even-numbered blocks in one panel and the odd¬numbered blocks in the adjacent panel being mined. As these blocks were depleted, the intermediate or pillar blocks were mined (Fig. 1). Following this checker¬board, the mining sequence went through a number of changes, finally evolving into diagonal retreat panel cav¬ing by blocks (Fig. 2). The numbers in Fig. 2 indicate the sequence in which blocks were undercut. Gaps in the numbering sequence indicate undercutting on the level outside the illustrated area. Geology The ore body is a low-grade deposit of chalcopyrite mineralization disseminated throughout structurally weak, highly fractured, strongly altered granitic host rocks. It takes the shape of a gently dipping elliptical cylinder consisting of an ore shell of variable thickness surrounding an interior waste core. Major and minor axes of the mineralized cylinder are 1524 m (5000 ft) and 762 m (2500 ft), respectively, and length approximates 2438 m (8000 ft). Ore is sufficiently fractured to break readily into medium-coarse size. The igneous rock complex containing the ore body is covered by a wedge-shaped blanket of Tertiary con¬glomerate which was brought into place by faulting along the major regional structure of the San Manuel fault. Thickness of the conglomerate cap varies from only 9 m (30 ft) at the east end of the ore body to more than 610 m (2000 ft) at the west. Structurally, the con¬glomerate is much more competent than the igneous host rocks and, when caving, it breaks into massive chunks. Conglomerate boulders seen in drawpoints underground are very coarse. The total rock column over the initial mining area of the 1415 grizzly level was 354 m (1160 ft) of which 122 m (400 ft) was ore, 79 m (260 ft) was leached igneous capping, and 152 m (500 ft) was conglomerate above the San Manuel fault. Diamond Drilling: From 4572 to 7620 m (15,000 to 25,000 ft) are drilled annually from underground workings to delineate the ore body. MINE DEVELOPMENT Haulage Level In the south or main ore body (see Figs. 3-6), with the exception of the draw and transfer raises, all the extraction openings are concreted (Seaney and Tobie, 1965). The haulage panel drifts, which are 18 m (60 ft) below the grizzly drifts, are first driven with pre¬concrete ground support. The drift, which has an arched section, then is concreted using mobile collapsible steel tunnel forms. The haulage drifts leading from the pan¬els to the hoisting shafts are not concreted. After the panel drifts have been concreted, the raise stations from which transfer raises will be driven are constructed and the raise-station ore-drawing chute is installed. The chute is prefabricated of A-36 steel with undercut guillo¬tine gates made of abrasion-resistant 2.5-cm (1-in.) steel plate powered by 20-cm (8-in.) air cylinder installed on each side of the raise station. Transfer Raises The transfer raises are lined with 15 x 20-cm (6 x 8-in.) cribbing and are 1.22 m (4 ft) in the clear. Each cribbing is protected from wear by a high carbon steel angle which is nailed onto the cribbing. The transfer raises are driven from each side of the raise station on an angle of 1.1 rad (63°). Each raise con¬sists of a main and a backover branch. The transfer¬raise driving crew consists of two men working one shift only. Grizzly Drifts After the transfer raise reaches the grizzly level, the grizzly drift can be driven. The grizzly drifts are spaced at 10.6-m (35-ft) centers and are driven parallel to the long axis of the ore body (see Fig. 2). This drift is driven by a two-man crew working on one or more drifts at a time using feed-leg machines. The eight grizzlies in the 42.7-m (140-ft) long drift are spaced at 5.3-m

Jan 1, 1982

By G. Kraemer, J. A. Le Gac, P. R. ZETTWOOG

I. INTRODUCTION During the course of our activities in assisting mining companies, we have had access to the monitoring results for personnel from a large number of uranium mines. The results differ greatly from one mine to another. One of our objectives has been to discover the origin of these differences. It is evident that they are largely due to variations in the geological context, to mining methods, and to the organization of the work ; but we also found that the rigor with which measures are implemented to prevent the personnel from being exposed to radiations is also a cause. In order to advise mining operators effectively, we have asked ourselves the following question Given the dosimetric results of a mining site, how can it be known - if occupational exposure has been reduced to the lowest possible level (dose optimization principle recommended by the ICRP) ; - if the exposures were justified, which would not necessarily be the case following a questionable choice of the mining method, for instance, or a lack of efficiency in the maintenance and repairs of ventilation apparatus or even an excess of radiation protection. It is necessary to establish some criteria, while taking into consideration the specific conditions for each mine, in order to determine whether a mining company is adequately implementing radiation protection procedures. This need led us to attempt a modelization of the occupational exposures of uranium miners ; the preliminary phase is presented here. Although voluntarily still very basic, this model makes it possible to demonstrate the role of certain dynamic or passive variables. Moreover, the model presents the concept of specific irradiation of a mining site equal to the collective dose received per ton of uranium metal supplied to the uranium mill. The specific irradiation can therefore be used to indicate the effectiveness of radiation protection procedures at a given mining site. This model can be used for - previsional exposure studies based on the use of data gathered at each site, making it possible to compare various work methods and to determine prevention means ; - qualification of "radiation protection" procedures at a mining site ; - detection of unjustified exposures ; - research of ways to reduce inevitable exposures. 2. DETRIMENT TO BE ASSOCIATED WITH A GIVEN EXPOSURE DISTRIBUTION Respecting exposure limits ensures that delayed stochastic effects are the only health risks to workers. If the hypothesis of linearity in the dose-effect relationship is applied, the detriment to a group of workers is proportional to the sum of exposure for all the workers, or the collective exposure usually expressed as man x rem. On this point the linearity of effects hypothesis makes it possible to consider that two different standard deviation distributions having the same mean value are equivalent. It would be possible, then, to reduce the standard deviation by changing the distribution of workers at the various working places without changing the mean exposure value. In other words, without taking into consideration the practical implications, individual limits can be respected by rotating the personnel or by reducing the individual working period which would be compensated by increasing the number of workers. The collective dose would remain the same, which would justify the use of the concept of collective exposure. 3. "RADIATION PROTECTION" QUALITY INDEX AT VARIOUS MINING SITES Each mining site is responsible for its own collective dose, but it is also responsible for producing a certain amount of uranium metal in the ore sent to the uranium mill. These two quantities must be brought together in order to establish the "radiation protection" quality index of a mining site. This quality index, called-specific irradiation (Ir) is expressed in rem.ton -1 for external irradiation, in mJoule.ton-1 for inhaled [a]-energy, and in Ci.ton -1 for inhaled radon *. The specific irradiation represents the health hazard which must be associated to the extraction of one ton of uranium metal. Figures la and lb present a system of coordinates for which the axis of the abscisses is graduated in tons of U metal and the axis of the ordinates is graduated in units of collective exposures. In the EEC, the regulation is based on the measurement of the amounts of radon 222 inhaled.

Jan 1, 1981

By Qinghua Song, Jianhong Chen, Shan Yang, Zhiyong Zhou

"The development process of Metal Resources Technical and Economic Evaluation Expert System (MRTEEES) is introduced in the aspects of requirements analysis, design of the expert system, main functions of the expert system and features of the expert system. The system is based on C/B/S mixed mode and uses ASP.NET technology with .Net Framework being chosen as the development platform and metal resources database providing data support at the bottom layer. The system is an auxiliary management system for metal resources technical and economic evaluation and has the basic functions of auxiliary decision analysis, metal resources database management, data management and comprehensive query. Technical and economic evaluation model can be set up by users independently according to at which stage a project is, mainly including exploration stage, development stage and production stage, and according to the mining methods, for example underground mining, surface mining and in-situ leaching mining. Then, the technical and economic evaluation parameters can be generated. By inputting the value of each parameter in a simple and convenient way, the evaluation results can be directly calculated out and shown in the form of diagrams among others, and feasibility evaluation report can also be automatically generated, making the technical and economic evaluation process accurate and efficient. As the system can achieve the functions of scenario analysis, sensitivity analysis, shareholder’s returns analysis, horizontal comparison of different projects, it can improve the ability of project senior decision makers for rapid response to the rival and meet the demand of pricing negotiations. IntroductionWith the arrival of a new round of industrialization upsurge in both developing and developed countries, growth of demand for metal resources is entering a new stage. Population involved in the new industrialized countries is about four times that of all previous industrialized countries, implying resource consumption increases dramatically both in speed and in quantity. It is a must to seize the strategic opportunity of developing and utilizing metal resources around the world to lay abundant resource base for national industrialization. Objectively speaking, global configuration of metal resources is an inevitable tend. Presently, many issues, for example, resources exploration, mining, merge of mining companies, metal trades, mining financing, etc., generally show the feature of globalization. For the reality that competition for metal resources is becoming fiercer, industry standards of mineral resources will undoubtedly be improved. Occupying large scale mineral resources with high grade and developing them with low cost is the base for the survival of mining companies and metal resources merchants. With the growing demand of global investment in mineral resources exploration, which is at the upstream of the industry chain, geological exploration industry is promising. Globally building market relationships for mineral resources exploration companies is key to their sustainable development. Developing large and super-large deposit with low cost is also an important approach to enhancing core competence of a country or an enterprise. Mining companies around the world are rushing to purchase mineral resources globally. However, in different countries, economic, financial and tax systems for mineral resources vary greatly, and mining and processing techniques, technical parameters and equipment selection also change a lot. Thus, pro-phase research work is necessary for feasibility evaluation of mineral resources projects and it can be time wasting if taken in the traditional way. It is necessary to accelerate the process considering the complex and rapidly changing international situation."

Jan 1, 2017