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By Marc Lonoff

Byproducts are more important to the copper mining companies than to the copper market. Copper ores frequently contain gold, silver, molbydenum and cobalt. With the increase in the prices of these metals copper mining has become more profitable. If byproduct prices remain high, additional firms will be attracted to the copper industry; the increased supply of copper will reduce the long-run copper price. However, this is unlikely to happen because the markets for the copper byproducts, especially molybdenum and cobalt, are small. Increased recovery from copper and other sources will depress the prices of these byproducts before the copper price is significantly affected. Should byproduct prices remain high the copper market will both settle at a lower long-run price and be more unstable in the short run. Since changes in the copper price do not affect the entire earnings of a multi-metal mine, the copper price must fall further before such mines reduce output in the short run. Higher byproduct revenues will make copper supply less responsive to copper price; shifts in copper demand will result in greater copper price fluctuations in the short run than they did before.

Jan 1, 1980

By Miguel Carrera, Jose Arnal, Julian Ortiz, Maarten Haest, Fouad Faraj

ShovelSense is a complete solution for bulk ore sorting at the mine face that combines robust shovel bucket mounted X-ray fluorescence sensors, with real-time material classification, fleet management integration and supporting cloud-based data infrastructure, to ensure each truck always goes to the right destination. Correct truck destinations are a challenge in many current mining settings because the material classification of a truck, which is the key input into the truck destination, is often based on estimates at block resolution (~20x20x15m), with limited adjustment for blast movement and mixing. In this contribution, we present a geostatistical, data driven approach to estimate the potential value of bulk ore sorting at the mine-face for seven open pit mines. The approach is premised on the assumption that truck loads within blocks follow a simple gamma distribution. Standard ore control data from the seven mines have been used to estimate the potential bulk sorting value at these operations and ShovelSense-integrated dispatch datasets have been used to validate these estimates for specific benches on which ShovelSense operated. ShovelSense results validate the ore control data-based value predictions and show how at mine-face bulk ore sorting value varies between distinct deposit types. In general, deposit types that have higher bulk ore grade contrasts, display more potential ShovelSense value. The manuscript concludes with the limitations and opportunities of applying the proposed value estimation approach to evaluate at mine-face bulk ore sorting opportunities for open pit operations.

Jun 25, 2023

By Robert R. Dorfler

AMPS[. has completed a rationed grinding ball test in a 2.93 m x 2.93 m ball mill at Climax Molybdenum Company. The test consisted of purging the ball distribution formed by charging 100%, 76 mm carbon steel balls with a distribution formed by charging 50% 76 mm and 20% 51 mm alloy steel balls. The resulting change in ball surface area yielded improvements in both tons ground and fineness of grind with a significant reduction in total ball consumption. This paper summarizes preliminary studies leading to the plant test, grinding efficiency results, and future plans to optimize the ball charge.

Jan 1, 1981

By Terry B. Linde, William M. Dowall

The recently completed In-Pit Crushing and Conveying Project at Phelps Dodge's Morenci Mine is rated among the world's leading materials handling projects. The operation involves the crushing and transport of more than 109,000 tonnes per day of copper ore, utilizing regenerative conveyors capable of providing over 5.5 megawatts (7,000 hp) of equivalent electrical energy. This paper describes the key aspects of the project and includes: Original project objectives; System design criteria; System description; Optimization studies; Major design features; Equipment description; and System commissioning.

Jan 1, 1993

By A. Lashgari, V. Kecojevic

This paper presents research on the environmental impact of digging and loading equipment used in surface coal mining. Environmental impact was assessed through the measurement of equipment exhaust and dust emissions, and sound level. In this study, exhaust emission refers to carbon dioxide (CO2), carbon monoxide (CO), nitrogen oxides (NOx), sulfur oxides (SOx) and volatile organic compounds (VOCs). Dust generation refers to fine particles (PM2.5), inhalable particles (PM10) and total suspended particulate matter (TSP). Sound level refers to the sound pressure level at different distances from the equipment. Modeling of the environmental impact of digging and loading equipment was conducted using the Microsoft Visual Studio.NET software package. This research was a part of a broader project involving the development of a software system for the selection of productive, cost-efficient and ecofriendly mining systems, sponsored by the Appalachian Research Initiatives for Environmental Sciences (ARIES). This research may assist mining professionals in quantifying the environmental impact of digging and loading equipment, and developing strategies for reducing the overall impact of such equipment in surface mining.

Jan 1, 2013

By John D. Abbatt, H. B. Newcombe

The Eldorado Epidemiology Project formally began in late 1978. It consists of a retrospective cohort study, to be followed by and to leave in place the mechanism for a prospective cohort monitoring program. These present and future activities are intended to merge into one another. The current study includes all Eldorado employees past and present for whom records are available. The objectives of the undertaking are - 1) to obtain cause-and-effect, dose-response data with which to evaluate the risks to workers in radon and radon daughter containing atmospheres, and to provide additional quantitative information on which to base possible improvements in working conditions; 2) to identify any dead employees of E.N.L. whose cause of death suggests that a potential compensation claim right exists for their survivors, and to similarly identify living ex-employees of E.N.L. whose work histories and states of health suggest potential compensation claim rights. The nature of the study has been determined by the history of Eldorado, which will be described briefly. Eldorado began (as Eldorado Gold Mines) in the late 1920's, operating relatively unsuccessfully in Manitoba. In 1930-31 the emphasis shifted from gold to uranium, and what became the Port Radium mine was first staked. From the very early 1930's until 1942 the company's purpose was almost entirely devoted to the mining of pitchblende - uranium ore - from the Port Radium site, and the extraction of radium from this ore at the Port Hope refinery in Ontario, a process from which the uranium was a largely stockpiled byproduct. In 1942 the company was requested by the Canadian government to accelerate the production of ore, and of uranium. At the same time, all its stockpile of uranium (then being stored in silos as a "waste" product which might later "come in") was acquired for the Manhattan Project. In 1944 the company shares were bought by the Canadian government and the company itself became Eldorado Mining and Refining (1944) Limited; it has operated as a Crown corporation ever since. The name was changed to Eldorado Nuclear Limited in 1968. The Port Radium mine was closed in 1960, and since 1953 the Beaverlodge uranium mine complex has operated on the northern shores of Lake Athabasca. Radium production was discontinued in 1954 when the radium circuit was taken out of the refinery, and since that time the refinery's products have been various types of refined uranium (uranium dioxide, uranium trioxide and uranium hexafluoride) plus metallic uranium. The refinery, of course, processes the company's own yellowcake but as it is one of only five refineries in the western world, most of its work consists of custom refining for a large variety of customers. In addition to the mining and the refining operations, there have been three other divisions involved, and these are, respectively, Aviation, R. & D., and Exploration, in addition to the financial, corporate, administrative and support activities. MATERIAL AND METHODS The human study population (Abbatt, J.D. et al., 1980) consists of all Eldorado employees who have ever worked for E.N.L. and for whom records are available. The total nominal roll to December 31, 1980 is approximately 21,000 and consists of employees of the Mining, Refining, R. & D., Aviation and Exploration Divisions, as well as a relatively small number of other employees (Head Office, etc.). The basic epidemiologic design, as has been mentioned, is for a retrospective cohort study merging into a prospective cohort study, and depending very heavily upon automated record linkage and file matching. The original design called for an initial assembly of the bulk of the nominal roll, prior to the first of two matches with the National Mortality File. The first of these two matches is now complete and the second is to be separated from it in time by 18 months to two years. The second match will be the definitive match prior to the analysis, the interpretation of the data, and the publication of the results. The period between the two matches was designed to permit completion of the nominal roll, including: - addition of previously missing individuals and groups, elimination of duplicates, and amplification of identifying information, - assembly of work histories and exposure information for each individual, - digestion and application of lessons and modifications resulting from the first match with the National Mortality File at Statistics Canada, - preparation of files for the definitive second match. The analysis of data will be carried out by the National Cancer Institute of Canada Epidemiology Unit based at the University of Toronto. This group, with whom Eldorado Nuclear has a memorandum of understanding, have for a variety of reasons, including their familiarity with record linkage and their expertise, been members of the Project Team since

Jan 1, 1981

By Thomas O’Neil

Part 2 of this article continues with a discussion of features that distinguish contemporary investment analysis in mining from earlier methods. Part 1 traced the evolution of evaluation methodology from the traditional and discounted cash flow methods to the present. Quantitative investment models of the past 20 years played an important role in mine evaluation, but a new integrated approach to mine investment decision-making is now developing because of growing economic, technical, and political uncertainties. One of those uncertainties -inflation-was discussed in Part 1. Increased Risk At a time when mineral production technology and analytical evaluation tools have never been more sophisticated, it may seem anomalous to claim that mining ventures are riskier. Whereas technical risk has, indeed, been notably reduced, business risk the chance of a major loss due to market factors-has increased. This has been caused by two factors: the rapidly rising capital requirements of new projects and increasingly unpredictable future economic conditions. These difficulties are compounded by the long lead times of large, new mining projects. Rising Capital Requirements A $100 million investment is no longer considered to be a large mining project in most cases. For a large metal mining project, the total cost of the concentrator alone can exceed $100 million. Amoco Minerals' Thompson Creek molybdenum mine being developed in Idaho will cost over $350 million, for a relatively modest sized (23-kt/d or 25,000-stpd) ore operation with no extraordinary features. For a major grass roots copper operation producing blis¬ter or refined copper, the capital requirements are billions of dollars, not millions. Project costs are higher for two reasons. One, the projects themselves are larger; and two, the capital cost per unit of production is rising rapidly. Projects are larger because lower grade ores, yielding smaller margins, must be mined in greater volume-particularly in view of the high fixed costs of modern technology. When analysts contemplate any capital project, an important consideration is the minimum economic size of the facility. This is the smallest size that can be built and still be competitive on a total unit production cost basis. In the conventional copper smelting business, for example, a production rate of at least 91 kt/a (100,000 stpy) of blister copper is about the absolute minimum for reducing unit fixed costs to competitive levels. Clearly, many variables affect the minimum economic size of a mining facility, but with few exceptions that size continues to grow. Unit capital costs of new mining projects also are continuing to rise. The accompanying table shows that the capital costs in primary copper production have risen faster than the cost of living. Of particular note is the capital charge of over $1.70/kg (77411b) required for a new, fully integrated facility in 1981. Thus, the mid-1982 copper price would not even cover the cost of capital for a new copper producing complex without considering the direct operating costs for such a facility. The breath-taking costs of new mining projects are caused by factors such as increasingly complex technology, the rising social costs of creating attractive living and working conditions for employ¬ees, and the expenditures re¬quired for environmental protec¬tion. For whatever reasons, how¬ever, unit capital costs in mining continue to exhibit staggering growth rates. Future Economic Conditions Risk is not only a function of the size of the potential loss-an im¬portant measure of which is the capital intensity of mining dis¬cussed above-but also of the like¬lihood of incurring the loss. It is here that the ever increasing lead times of mining projects and the glaring limitations of economic forecasting are so important. In comparison to nearly all other in¬vestments, mining has more capi¬tal at stake for a longer period of time before commercial uncer¬tainty is resolved by distant eco¬nomic conditions that are subject to high forecasting errors. A sobering example of the pre¬ceding points occurred with the Lakeshore copper mine in south¬ern Arizona that Hecla Mining Co. undertook to finance and manage in 1969. The project was a joint venture of Hecla and the El Paso Natural Gas Co. and included two mines-one producing oxide ore and a second in the sulfide zone¬and a complex processing facility producing both electrowon cop¬per cathodes and copper precipi¬tates. The project originally was

Jan 12, 1982

By William T. Griswold

Few metals or industrial minerals can boast a better growth record over the past fifteen years than the borates. This period of growth occurred before the influence of any large-scale use in plastics or high-energy fuels. These new developments not only provide expanding markets, but the programs related to their development are generating research information that will lead to many other new uses. The most dynamic growth areas for refined products of boron or borates are in the acid, the oxide and the reduced forms, such as the metal and the hydrides. Expanding requirements for these products will modify the emphasis on the use of naturally occurring sodium borates. Any borate from which boron can be economically extracted will become a potentially important ore. This is particularly true of the calcium borate, colemanite. Only one large deposit in the Western World has ever been developed where the principal economic minerals are sodium borates, This factor alone would discourage any exploration program directed solely toward the discovery of these ores; on the other hand, ample exploration opportunities for calcium borates exist in the desert regions of California and Nevada. It is therefore reasonable to restrict a program to this region, in spite of geologically favorable areas in foreign countries such as Turkey, Argentina and Chile.

Jan 1, 1959

By Neil D. Policow

Dewatering of the fine clean coal fraction, - 28 mesh (595 microns) represents one of the most difficult and costly separations in typical plant circuits. Since 1969, the Screen Bowl Centrifuge has offered an economical alternative to filtration/thermal dryer combinations. The performance of Screen Bowl Centrifuges has been continuously modified and improved to dewater not only coarser coals, but also today's variety of ultrafine material. Actual test comparisons with alternative methods show the Screen Bowl Centrifuge to be most cost effective and efficient when all factors are considered.

Jan 1, 1982

By David A. Heitman

The Buick Concentrator is a highly automated operation which utilizes a process control system for continuous monitoring and control. The process control system development, installation, and application to the concentrator grinding and flotation circuits are presented in this paper. The design and implementation of analog field sensing instrumentation, a continuous on-stream X-ray analyzer, grinding and classfication circuit controls, and a recent digital process control system utilizing a centralized computer are discussed. Emphasis is placed on the project planning and installation of the digital control system including a presentation of future plans for continuing application of computer control.

Jan 1, 1982

By Norbert Patzelt, Johann Knecht

INTRODUCTION Grinding mills have been used in the minerals processing industry for over 100 years. Their dimensions have grown continuously during this time. Besides increasing throughput rates of grinding plants due to the depletion of high grade ores, the lower specific in- vestment costs, as well as reduced operating and maintenance requirements are major reasons for this trend. When selecting new plant equipment one must consider that design principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger size of equipment. Modern calculation methods as for instance the Finite Element method already contribute considerably to the safe design of the huge equipment being built today and are a standard tool of the design engineers. More recently, modern computer programs are also being used in order to size the equipment to meet the process requirements. Today, two design principles are on the market - one which supports the weight of such a unit on trunnion bearings through cast conical endwalls and one which is supported through slipper pad bearings arranged at the circumference of the mill shell (Fig.1). The reason for the development of this alternative grinding mill design can be found in the past. During the sixties and seventies the growing sizes of ball mills with high LID ratios caused many mill failures due to cracked endwalls. The accuracy of the calculation methods as well as the quality standards for castings were not developed to a degree required for such kind of heavy equipment. One way to overcome these problems was the increase of the manufacturing quality standards as well as the introduction of the finite element method based on the analysis of the experience available. The biggest grinding mills being built today are large size SAG mills with cast conical endwalls and trunnion bearings (Fig.2). This is due to the fact that mill manufacturers who had come from the conventional ball mill design adopted these principles as well to their SAG mills. These grinding mills perform well without special concern to the operators. Other manufacturers overcame the problems as mentioned above by eliminating completely the heavy castings and trunnion bearings and the problems associated to it (Fig.1). This design was originally applied to ball mills for the mining and other industries. Due to the success of these shell supported ball mills, this design principle was also applied to SAG mills(Fig.3). Despite of the fact that the majority of today's grinding mills are built to the conventional design it is also interesting to have a look at this alternative. Principles which have proven their reliability on sizes of today's equipment do not automatically warrant a successful operation on the ever larger equipment if bigger mill sizes are realized only based on the pantograph principle. With growing grinding mill sizes, the mass and volume flows through the equipment increases rapidly. Thus it is very important not only to concentrate on the safe design of the structural components of the equipment but as well on the process requirements. The influence of the design on important process parameters of dry and wet grinding plants are discussed thereafter. It shall be shown how modern computer programs can assist in the optimization of the design of components in order to fulfil the operational requirements of such large size equipment. PROCESS REQUIREMENTS OF LARGE SIZE GRINDING MILLS Dry Grinding Mills The world's biggest ball mill is a dry grinding ball mill having a diameter of 6.2m and an overall length of 25,5m with a drive power of 11,200 KW or 15,000HP. This grinding mill dries and grinds gold ore at a rate of 500 tons per hour at a moisture content of up to 9,5%. As shown in Fig.4 this mill was built as a shell supported unit. In fact only this design principle allowed to meet the process requirement. This mill could hardly be built with cast conical endwalls due to the constraints of the trunnion bearings limiting the mill inlet. The following case shows how modern computer programs can help to meet the design criteria of the air system of large size dry grinding plants. For dry grinding plants, the gas flow through the SAG mill has to match the drying, as well as the material transportation require-

Jan 1, 1998

By V. R. Degner

The minerals industry trend toward very flotation machines, exceeding 10 m3, began in the late 1960's and responds to the continuing desire to achieve better plant economics through improved plant floor-space utilization, power requirements, and operational /maintenance manpower heeds. In general, as process design tonnage increases, the benefits of the larger flotation machine becomes more evident. The development of the larger flotation machine can be distingushed from earlier, smaller, machines in demanding a rational "scale-up" procedure to guide the basic design, thereby producing a machine prototype capable of satisfactory-operational evaluation at minimum cost. Proven flotation machine "hydrodynamic" scale-up consideration, successfully applied in the development of both the WEMCO No. 164 (28.3 m3) and No. 190 (42.5 m3) machine will be reviewed, and the resulting economic benefits of large flotation machines in high tonnage applications will be summarized.

Jan 1, 1985

By K. A. Natarajan

Platinum electrodes are not inert as originally thought to be. The reactivity of platinum electrodes can explain their erratic behavior in many electrochemical measurements of metallurgical interest, e.g., in flotation systems, streaming potential measurements, contact-angle measurements, and in leaching systems. The anomalous behavior of platinum electrodes in redox potential measurements in aqueous systems was studied through Eh and pH measurements in water-oxygen, iron-water-oxygen and manganese-water-oxygen systems. Stability relations between Fe++ and Fe(OH)3 and between Fe(OH)2 and Fe(OH)3 were studied to judge the correspondence between experimental and theoretical equilibrium lines. The practicality of redox potential measurements in estimating ferric-ferrous ratios in aqueous systems was investigated along with their suitability as indicators in leaching operations, e.g., the removal of iron by aeration from manganese leach solutions.

Jan 1, 1970

By J. L. Thompson

Lime made in a coal-fired, rotary kiln has two distinct parts: the interior of the pebble which constitutes about 90% to 95% of the mass, and the surface dust layer which is the remaining 5% to 10%. Uncalcined carbonate and impurities in the product may be present and will not contribute to the hydration reaction. Sulfur, in the form of calcium sulfate, inhibits the hydration reaction out of all proportion to its actual weigh-percent concentration (perhaps more than 100 times). If the stone contains dolomite, this will also inhibit the reaction as its rate of hydration may be nearly 3 orders of magnitude slower than that of calcium oxide. Early in the hydration reaction dolomite appears practically inert, but given sufficient time, it will hydrate. The heat of reaction is nearly the same as for calcium oxide, but the time necessary for complete hydration may be hours; whereas for calcium oxide it is only fractions of a second. The dust layer on the surface also inhibits hydration as it contains a collection of compounds which have low heats of hydration. This layer can tie up additional calcium and delay or prevent its reaction. In addition to these chemical processes, an important physical characteristic of the lime is its specific surface area. This is effected by the thermodynamics of calcination and the original stone grain chemistry. All of these factors are agglomerated in the gross slaking rate of lime. The total hydration of the chemical complex that is called commercial lime can be more easily understood by considering the individual reactions and combining them to form a more comprehensive picture of the situation. Calcium oxide and magnesium oxide hydrate similarly, differing only in their reaction kinetics. Dolomite, however, reacts differently than a simple mixture of calcium oxide and magnesium oxide as its hydration is strongly influenced by calcination thermo-dynamics and stone grain structure. Finally, the dust hydration and composition should be examined Understanding these reactions enables one to examine hydration reaction graphs with more insight into the factors which determine their shape.

Jan 1, 1978

By V. P. Petrov

There is a great number of asbestos deposits in the USSR. Some of them are quite large ones. The main deposits of chrysotile asbestos can be seen in the sketch map (fig. 1). Bazhenovo deposit situated in Sverdlovsk region is the largest of them. It is on mining operations since year 1889. Systematical studies of asbestos deposits were begun from the first years of Soviet Government and connected with the names of K.E. Tarasov, P. M. Tatarinov, B. J. Merenkov, N.D. Sobolev, F. V. Syromjatnikov, V. N. Lodochnikov, V.P. Petrov, V.P. Eremeev, L.A. Sokolova, V. R. Artemov, V.F. Dybkov and many others. This investigations served the USSR to be one of the most important producers of asbestos in the world. In the limits of the report a characteristic of the main asbestos deposits of the USSR may be done in general outline, from the point of view of their genesis, and in order to show their common features and differences that must be taken into account for forecasting, exploration, and mining of asbestos. All chrysotile asbestos deposits of the USSR were formed of magnesium rock as a result of comparatively lower-grade temperature hydrothermal alteration and free crystallization of fibrous chrysotile in open cavities that served as catalyst As a matter of fact an important mineral for industry is not chrysotile in general, a widely distributed rock-forming mineral, but a long fibrous and mainly vein variety of chrysotile. It must serve for geologists as an important prospecting criteria. A geologist should search for the mineral not only places of possible chrysotile generations, but places where asbestos may be crystallized in its long fibers that could be fiberized well. The Soviet geologists distinguish two groups of chrysotile-asbestos deposits: (1) related to ultrabasic rocks and (2) related to sedimentary dolomite terranes, taking into account what the maternal rock was affected by hydrothermal solutions. The largest deposits related to ultrabasite. The largest deposits in dolomite are small ones but their asbestos contains almost no iron. So much asbestos is a better filler into insulating plastics. There is a few types of asbestos deposits in the limits of the ultrabasic group. The largest of them is a Bazhenovo type which the deposits of Urals are attributed, and Bazhenovo deposits is the first instance. The Bazhenovo type is characterized by the most complete zonality in disposition of various kinds of asbestization. This type of deposits is practically the only one to be in working now in the USSR.

Jan 1, 1978

By David A. Zegeer

Since 1980, more than 40 miners have lost their lives by being trapped and suffocating in stock-piles, bins, silos, and similar facilities. Mine Safety and Health Administration (MSHA) is concerned about these accidents, most of which could be prevented if management and miners appreciated the hazard. In February 1986, the coal industry had a suffocation accident that drew national attention. A group of engineers were standing on a coal stockpile. It gave way under their feet. Five of the engineers were buried in the coal and suffocated. As with all fatal accidents, MSHA investigated the cause of this tragedy. I also asked my staff to look into the history of similar accidents. I found out it is rare to have multiple fatalities due to suffocation in material. But, over the years, there has been a long series of individual fatalities from this type of accident. Here are some examples. In Oct. 1985, a coal mine employee in Virginia entered a hopper to collect a coal sample. The feeders were started while he was standing in the hopper. He sank into the coal and suffocated. In Sept. 1985, an employee at a Texas sand company climbed onto a surge pile, to check a problem with sand flow. When other employees found him, the sand had swallowed him. Only his face and one hand were showing. Though the victim's face was clear of the sand, he suffocated before rescuers could free him. Many people do not realize you can suffocate in this kind of situation, though your head is free. Every time you exhale, the material has room to shift and press harder against your chest. Then you have less room to inhale. If this goes on long enough, it gets to the point where you cannot inhale at all. That is what happened to the victim in this accident. In May 1985, a plant operator at a Wisconsin sand and gravel operation purposely slid into the drawhole of a surge pile. He tried to break up large rocks that were clogging the drawpoint. Material piled on the side of the surge pile slid down and buried him. The operator was wearing a safety belt and line. But he was covered by 3 to 3.7 m (10 to 12 ft) of material and could not be pulled free. This incident illustrates another point. A safety line can help. But it is not enough. Workers should never put themselves in a position where loose material can fall on them from above. In Jan. 1985, a young Utah coal miner was waiting for a ride home from a co-worker whose job was at the tipple. A jam developed in the coal bin. The men could not clear the jam from the outside. So the young miner jumped into the bin. He started digging with a shovel at the toe of the coal pile. It caved in on him and he suffocated. The young miner had only three months of mining experience. He had not been trained for work around the coal bin. Since 1980, 16 coal miners and 26 nonmetal miners have lost their lives in loose-material suffocation accidents. Suffocation accidents occur in many areas around mines - bins, hoppers, stockpiles, surge piles, and silos, among others. One recent accident occurred at a surface coal mine in Illinois. The victim was walking over an area that had been drilled and blasted the week before. He fell into a void that had been crusted over with earth. He too suffocated. And many materials have been involved in suffocation accidentscoal, crushed stone, sand, salt, different kinds of ore, ammonium nitrate pellets used in blasting, and a powdery byproduct of iron ore milling called calcine. In addition to fatal accidents, there are near misses. We do not know how many. Most of them are not reported. A person who gets trapped in material, but is rescued, is usually unhurt. MSHA sometimes hears about these accidents by chance, or because they are reported in the local media. We do know that near misses are happening far more often

Jan 1, 1987

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

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 Janet A. Johnson, T. B. Borak, K. J. Schiager

INTRODUCTION In the USA regulatory agencies have adopted standards which limit the allowable concentration of ore dust in underground uranium mines. The American Conference of Governmental Industrial Hygienists has recommended a threshold limit value for silica which has been incorporated by reference in federal regulations (30 CFR 57, 1980). The Nuclear Regulatory Commission (NRC) imposes a limit for unprocessed uranium ore dust for reducing the exposure to radioactive materials from inhalation (10 CFR 20, 1980). We have been unable to trace the origin of the NRC standard. Recently the ICRP has published recommendations concerning limits for intakes of radionuclides by workers (ICRP, 1979). We have used the ICRP methodology to compute the dose equivalent commitment to lung and bone from inhalation of insoluble ore dust particles. These values were used to derive air concentrations which would provide an acceptable risk for induction of cancer from occupational exposure. Results indicate that the new ICRP recommendations are more restrictive than the present NRC standard for radioactivity. The maximum allowable dust load in a uranium mine will depend on both the concentration of silica (% Si02) and the amount of radioactivity (% U308). We have combined these standards assuming complete independence of biological hazard to establish boundary conditions for which either silica or radioactivity is the dominating factor. Based on these results we present suggestions for analysis of ore dust to insure compliance with regulatory agencies. SILICA STANDARDS Silicosis, fibrotic lung disease produced by crystalline silica, was one of the earliest forms of occupational disease to be recognized. The pulmonary lesions caused by silica, silicotic nodules, consist of concentrically arranged bundles of collagen fibers. Fusion of these nodules results in progressive massive fibrosis causing alveolar changes which decrease ventillation and blood flow in the lungs. The current Threshold Limit Value (TLV) for mineral dusts is based on the concentration of free silica. The standard is expressed in three forms: (1) Particle count standard: TLV (mppcf) = 300/(% Si02 + 10) (mppcf = million particles per cubic foot) (2) Respirable dust mass standard: TLV (mg/m3 ) = 10/(% respirable Si02 + 2) (3) Total dust standard: TLV (mg/m3 ) = 30/(% Si02 + 3) The most commonly used form is the respirable mass standard. These curent values for silica are based primarily on epidemilogical studies of Vermont granite workers and other occupationally exposed workers. In the Vermont study no cases of silicosis were seen in individuals employed after dust control measures were initiated and whose subsequent exposure averaged less than 5 mppcf. The free silica concentration in the airborne dust to which these workers were exposed averaged 25%. A concentration of 10 mppcf of granite dust is considered equivalent to 0.1 mg/m3 quartz. The current respirable mass standard allows a maximum silica concentration of 0.1 mg/m3. NIOSH (1974) has recommended a reduction in the current TLV for respirable silica to 0.05 mg/m3 Si02. As with radiation standards, the TLV for mineral dusts is an upper limit for time weighted average exposure. It is recommended that dust concentrations be maintained as far below the TLV as current practices permit. RADIATION STANDARDS The present concentration limit for airborne natural uranium in ore dust prior to chemical separation is 75 µg of uranium per m3 of air ( 10 CFR 20, 1980) . This corresponds to 1.85 Bq/m3 (50 pCi/m3 ) of natural uranium, U-238, U-235 and U-234 and is equivalent to 3.7 Bq/m3 (100 pCi/m3 ) of gross alpha activity with radioactive equilibrium between the long-lived alpha progeny through Ra-226. The origin or basis for this standard is not known, but it presumably includes both chemical toxicity and risk from somatic radiation injury. Recently the International Commission on Radiological Protection (ICRP, 1977) has published recommendations concerning the objectives and criteria for limiting radiation exposures. This was followed b y a revision of the limits for intake of radionuclides by workers (ICRP, 1979; ICRP, 1980). The previous compilation of limits for internal emitters was published by ICRP Committee II in 1959 (ICRP, 1959). The latest version includes new methodlogy outlined in ICRP Publication 26 (ICRP, 1977) as well as extensive metabolic information

Jan 1, 1981

By Rafael Moure

On April 21, 1980 the Oil, Chemical and Atomic Workers Union (OCAW) submitted a petition urging the revision of the Mine Safety and Health Administration (MSHA) permissible exposure limit to radon daughters in underground mines. This petition asks for: 1. Annual worker exposure to be lowered from the current 4 WLM/year to 0.7 WLM/year; 2. Full economic protection for workers whose exposure exceeds the proposed 0.7 WLM/ year; and 3. The primary means of lowering workers' radiation exposures shall be environmental controls. I will present the scientific rationale for the lower standard requested in our petition to the Department of Labor. [1. Radiation Exposures to Uranium Miners] Two types of radiation exposures are present in the environment of uranium mines: A. [External Exposure] to gamma sources which until 1978 was considered as negligible compared with internal alpha radiation. R. E. Utting (1), in an investigation of three Canadian uranium mines, found levels of exposures adding to an annual dose equivalent of 3 rems (30 mSv)/year. He concluded that gamma dosages may be a major fraction of the maximum permissible dose for many individual miners. The U.S. experience has shown that there is significant gamma exposure in U.S. uranium mines and, therefore, it could contribute to the total radiation exposure to the lungs in combination with internal alpha radiation. Before ventilation improvements in uranium mines, the contribution of external radiation to total lung dosage was relatively small. After improvements in ventilation, the reduction in internal radiation exposure appears to be in 1980 as low as 1/10 of the 1950 dosages. However, ventilation improvements do not affect external gamma radiation. Gamma radiation remains at the same 1950's level, and it represents today a higher proportion of the total radiation burden to the miner's lungs. MSHA regulations (57.5-47) state that if gamma exposure is below 2.0 milliroentgens per hour, then workers are not required to wear gamma dosimeters and cummulative gamma radiation records need not be kept. Assuming an exposure of 1.9 milliroentgens per hour per 173 hours/month for 12 months,a miner would accumulate a dose from external radiation of 3.9 rems/ year. This dosage would, of course, go unrecorded. MSHA, therefore, should consider a downward revision of the external radiation permissible exposure limit of 2.0 milliroentgens/hour. MSHA should make clear that this exposure level, when added to the internal exposure, does not make for sound health physics -even though the measured current levels do not approach 2.0 milliroentgens/hour. B. [Internal Exposure] to alpha radiation is still the larger contributor to lung irradiation in uranium miners. The target organ is still the lung. The parameters to be considered in determining the biological radiation load appear in Table I for radon daughters present in the upper respiratory tract (URT). TABLE I RADIATION DOSE PARAMETERS IN THE URT OF URANIUM MINERS Parameter Uranium Underground Mining [1. Quantity of the Absorbed Radon daughters (Po-218, Nuclide Pb-214, Bi-214) deposit in bronchial and lung tissue. (Measured by air concentration, Working Levels) 2. Type of Radiation Emitted Alpha Particles 3. Energy of Radiation High-Linear Energy Transfer 4. Half-life of the Nuclide Ra 222, 3.8 days in the Organ Po-218 (RaA), T1/2 = 3.1 minutes Pb-214 (RaB), T1/2 = 26.8 minutes Bi-264 (RaC), Vi = 19.7 minutes 5. Selection Factor of the Direct inhalation with mine Organ dust to be deposited in the tipper Respiratory Tract 6. Mass and Form of the Organ Especially cancer sensitive, tracheobronchial region and bronchi* 7. Distribution of the Activity Dependent on deposited dust in the Organ distribution in the Upper Respiratory Tract 8. Chemical Nature of All compounds (RaA, RaB, RaC) Incorporated Substance biologically insoluble. (Only localized, irradiation) *See Reference (2)] The parameters of Table I depend, in turn, on the demography of the exposed population, especially: 1. Age 2. Sex 3. Health Status 4. Genetic Constitution 5. Lifestyle (Smoking, Diet) Some of these demographic characteristics of uran-

Jan 1, 1981