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By M. Machina

Anew approach to the use of projectiles for rock fragmentation uses recently-developed technology to optimize projectile shape and composition to achieve the desired type of breakage. Projectiles can also be made of various materials, as determined by the requirements of downstream processes. This paper describes laboratory tests with an 21-mm (8-guage) gun, and field tests with a 60-mm cannon.

Jan 1, 2003

By Raja V. Ramani

Worldwide, promulgation and enforcement of health and safety laws and regulations have been the primary means of ensuring minimum and uniform standards of health and safety in the mining industry. In U.S. coal mines, a new era was ushered-in in 1969 for the control of health hazards from respirable coal mine dust. Since then significant progress has been made on all aspects of the respirable dust problem-its definition, monitoring, reduction, and control. In the intervening years, our ability to monitor, analyze, and characterize the mine atmospheric environment and the health of miners has increased many fold. Also, advancements in production technology have far outdistanced enhancements in environmental control technology. Therefore, questions are continuing to be raised on the adequacy of the characterization and control measures with regard to both the airborne respirable dust in mines and the health effects of respirable dust. Late in 1995, the Secretary of Labor appointed an Advisory Committee on the Elimination of Pneumoconiosis Amona Coal " Mine Workers (Dust Advisory Committee) to review the progress towards elimination of coal worker's pneumoconiosis (CWP) and silicosis, and in the program to control respirable coal mine dust in underground and surface mines, and make recommendations including improved standards and other appropriate measures for the control of exposure to airborne coal mine dust. The Committee's report was released in October, 1996. In this paper, a summary of the recommendations of the Committee is presented. A brief discussion of the background, and the issues and findings leading to the recommendations is also presented. The Committee concluded that the progress made towards making mines safer from the health hazards of respirable coal mine dust is substantial but that it is not sufficient to achieve the intent of the 1969 Coal Act. The Committee's recommendations encompass a wide range of issues addressing permissible exposure limits (PELS), sampling, dust control plans, hazard and medical surveillances, role of miners, training, and research and development.

Jan 1, 1997

By Tuncel M. Yegulalp

This article summarizes some of the results of the continuing research activity at the Henry Krumb School of Mines, Columbia University, directed towards a better understanding of the stochastic behavior of slurry flows in underground coal-slurry pipeline systems. The slurry system under consideration here replaces all conventional haulage systems between the coal face and the coal preparation plant on the surface. At the beginning of the pipeline system, flow rate and frequency of coal slurry is very closely related to the activities of production equipment at the coal face. Due to several scheduled and unscheduled interruptions in continuous miner's coal mining activity, resulting slurry flows possess stochastic characteristics. Proper understanding of these characteristics is believed to lead to a more realissic and economic design of slurry handling equipment. Fundamental ideas supported by field observations are presented and a Semi-Markov model of the slurry flows is developed. As a result of this formulation, behavior of combined slurry flows from N independent coal faces can be expressed in analytical form. The contents of a slurry surge tank receiving slurry flows with certain stochastic characteristics is investigated as a function of discharge pump characteristics and time.

Jan 1, 1981

By R. T. Torrie, J. R. Mernagh, D. B. Chambers

l, INTRODUCTION Uranium mine and mill workers are exposed to external gamma and beta radiation fields from radioactive ore. In the past the average uranium content of ores mined in the United States and Canada ranged from about 0.1% to 1.0% U308 with pockets of much higher grade ore. Holiday (1973) reported that radiation surveys in the U.S. uranium mines found mean gamma "radiation rates ranged from 0.20 to 0.70 mrem/h. Such radiation rates cause relatively insignificant exposures." Others also concluded that the external gamma radiation fields associated with uranium mining did not result in significant worker exposures (Federal Radiation Council, 1967; Simpson, S.D. [et al], 1959) Gamma exposure levels in most modern-day Canadian uranium mines are reported to be low with average annual exposures estimated to be less than 1 rem/a (Frost, S.E. [et al], 1981). However, two developments are taking place which affect the potential significance of external radiation fields in the uranium mining environment. The first development stems from the most recent system of dose limitation developed by the ICRP which is intended to limit the workers' overall risk from exposure to ionizing radiation through the adoption of a sum rule (ICRP 26, 1977) which combines external and internal radiation exposures. In the case of exposure to radon daughters the sum rule will have the effect of reducing the annual exposure limit below the recommended limit by an amount that depends on external radiation and other sources of internal exposure such as the inhalation of ore dust (ICRP 1980). The Atomic Energy Control Board of Canada (AECB) is reviewing this subject and is expected to produce its recommendations shortly. Irrespective of the form of the sum rule eventually adopted by the AECB, it is clear that the net effect of the sum rule will be a collective reduction of the individual dose limits for individual exposure pathways. The second factor is the increasing development of high grade uranium deposits in Northern Saskatchewan. Some of these ore bodies have an average ore grade of 1% to 5% U308 or greater. Since the potential external radiation fields increase in proportion to the ore grade, it is apparent that increased effort in radiation protection planning is required in order to develop safe yet workable methods for mining and milling such ores. This paper is intended to provide information which can be of assistance in the formulation and development of a mining and milling plan. The principal focus of the paper is source identification and the design of radiation protection measures to limit external gamma radiation exposure. The exposure of workers to external beta radiation fields is also discussed. The paper is organized as follows: - Section 2 deals with source characterization. - Section 3 discusses the effects of finite source size and distance (i.e. geometry effects). - Section 4 presents selected data that are useful in evaluating shielding requirements. - Section 5 discusses the potential beta radiation fields. - Section 6 discusses practical data requirements for worker exposure scenarios. - Section 7 presents a variety of work exposure calculations. - Section 8 is a summary of this paper. 2. SOURCE CHARACTERIZATION This section develops the basic formulae for estimating the fluxes and doses from external gamma radiation. The calculation of the radiation flux due to a distributed source (i.e. a linear, area or volume source) as a function of distance assumes that any distributed source can be treated as a summation of point sources. [ ] Uranium ore contains radionuclides from both the decay chains of U-238 and U-235. In this paper the radioactive daughters are assumed to be in secular equilibrium with the uranium parent. (If natural thorium were present in the ore in significant quantities, the gamma rays originating from Th-232 would have to be added to the gamma rays from the uranium series). In all, there are over 50 separate gamma rays (as well as alpha and beta particles) emitted from the U-238 and U-235 radioactive decay series (USHEW 1970). The total

Jan 1, 1981

By James R. Hindman

Vermiculite is the mineralogical name generally applied to a group of hydrated, trioctahedral ferromagnesian aluminum platty or micaceous silicate minerals. In microscopic particles the mineral as mineral is differentiated from montmorillonite by a higher Si:AI ratio of approximately 3:l and a higher layer charge and higher cation exchange capacities. Commercial vermiculite, the vermiculite which is mined, beneficiated, and discussed in this chapter, is an important industrial mineral which is formed by the alteration of macroscopic particles of biotite and iron-bearing phlogopite. In natural form it has the size and shape of mica, while possessing the ion exchangeable properties of zeolites and some clays. Within the internal structure of vermiculite lie water molecules, which when rapidly heated to high temperature, transform into steam causing the vermiculite particles to increase in volume. This process of thermal exfoliation produces a lightweight product which finds use in various construction products, agriculture, horticulture, and other industrial applications. In exfoliated form it has the low density and biological inertness of expanded perlite, while maintaining a chemically active surface. Most of the uses of vermiculite are for the thermally exfoliated material. In addition, the cation exchange potential of vermiculite and its layer-silicate structure are properties which can be used as the basis for numerous products including intumescent coatings and gaskets, inorganic films, and the treatment of nuclear and other toxic waste. When chemically and physically modified, it possesses some properties common to organoclays and synthetic compounds. In high purity concentrates and manufactured products, these properties can be altered and optimized in applications to produce superior performance. GEOLOGY Mineralogy Vermiculite was first described as a mineral by Thomas H. Webb in 1824. Webb believed vermiculite to be a variety of talc and described its diagnostic property of thermal exfoliation thusly: "If subjected to the flame of a blowpipe, or that of a common lamp, it expands and shoots out into a variety of fanciful forms, resembling most generally small worms having the vermicular motion exact." Observing that with one exception, all commercial vermiculite has formed by the alteration of iron-bearing phlogopite, a simple definition of vermiculite might be the alteration product of trioctahedral micas, macroscopic particles of which display marked exfoliation when heated rapidly to temperatures above 374°C (the critical temperature of water). The first vermiculite described was found near Worcester, MA, but it was later pointed out by Brush (1866) that vermiculite had been known for some time in Japan as a children's novelty. Because the techniques for mineral analysis were somewhat primitive com- pared to those routinely available today, and because the structure and crystal chemistry of the micas and related minerals were not widely understood, the early literature generated at least 20 "new" minerals which were nothing more than variations in the color and chemistry displayed by vermiculite. Trade names such as Dugarlite, Mandolitem, Strong-LiteTM, and Zonolitem are sometimes used to denote commercial vermiculite, thermally exfoliated vermiculite and their products. Vermiculite particles resemble mica and vary in color from colorless through tan and green to black, depending on the chemical composition. Most physical and optical properties are close to those of biotite, with perfect basal cleavage, hardness between 1.5 and 2.5 Mohs, and specific gravity between 2.2 to 2.8. Fine-sized particles will feel talcose, especially when wet. However, for commercial vermiculite the diagnostic test is that of thermal exfoliation. Vermiculite can display a wide range of compositions, depending on the composition of the original mica and the change of chemistry during weathering and ion exchange. Chemical data for some commercial vermiculites are given in [Table 1]. There are numerous analyses for vermiculite given in the literature, and the analyses of vermiculites from commercial deposits will fall within the range indicated by the average analysis in the table. Many analyses separate the adsorbed water (H2OP) from crystalline water and water as hydroxyl (H20+). The basis of separation of adsorbed water is generally considered to be 100°C. However, hydrate water within the vermiculite interlayer begins to be lost at 75°C and lower. Therefore, measured water content and loss on ignition are combined in the table as H20[total].

Jan 1, 1994

By A. W. Easton

INTRODUCTION This presentation will address changing trade patterns in lead and zinc from the mid 1980's to the early 1990's. Lead and zinc deposits are developed where they are found in many parts of the world. As many are not in major consuming areas, there is significant international trade in both lead and zinc concentrates and in refined metal. Forecasts of mine production have been analyzed, primary and secondary refining capacity and future consumption levels have been estimated, to arrive at the changes in trade patterns expected during the next half decade. This paper will first address concentrate and metal trade patterns for lead, then for zinc, and finally some remarks will be made on captive mine/smelter agreements, on regional trade agreements, and on barriers to trade. I would like to note that. there is only a small international trade in semi-fabricated products such as zinc die-cast alloy, zinc oxide, lead oxide or rolled lead and zinc. Most of the world's zinc and lead trade is in concentrates or refined metal. LEAD TRADE PATTERNS Lead Concentrates Six countries are exporters of lead concentrates. Three countries are in the Americas - Canada, Mexico and Peru, with two in Africa - Morocco and South Africa (including Namibia), and Australia. Due to the limited growth in lead consumption and consequent minimum investment in new production facilities expected, mine and metal production levels and future consumption levels are not forecast to change much between now and 1990. Accordingly, the tonnage of lead concentrates exported will remain essentially unchanged from these six countries. In total they have exports of 480,000 metric tonnes (metal content)* and by 1990 this figure is forecast to rise only slightly with the only increase taking place in Mexico. The concentrates are exported mainly to Western Europe & Japan. Apart from these six main lead concentrate exporting nations, a significant change is expected in the United States. In the mid * All concentrate trade figures are based on metal content

Jan 1, 1986

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 Jacques B. Wertz

It is well known that the major fractures and lineaments presently recorded in Western North America are preferentially oriented along a general northwest direction (Figure 1). There are other features, however, more subtle, more elusive as a rule and which can only be inferred, occurring at right angle to this broad NW direction. It is intended a) to examine some of these NE or transverse trends, b) to compare these regional NE breaks to similar ones that transversally cut through the eastern part of the European Tethys, more specifically in Iran and Turkey, and c) to study their potential relationship to mineralization in both areas.

Jan 1, 1973

By P. A. Plunkert, D. E. Morse, L. McCartan, S. F. Sibley

Following the recession of 2001, restrained consumer spending, declines in domestic manufacturing and industrial output, and increased unemployment led to a lack-luster year in the domestic economy. In 2002, the U.S. real gross domestic product (GPD) grew only 2.4%. Significant production declines in the U.S. metals industry were registered by copper, lead and zinc producers. Steel producers, despite a slight improvement in shipments, continued to face strong foreign competition, higher energy costs and lower prices for their products. Homebuilding and other domestic construction sectors— major consumers of nonmetallic mineral products such as cement, brick, glass and stone — remained strong enough to help raise the output of industrial mineral materials slightly above 2001 levels. A strong U.S. dollar relative to other national currencies continued to weaken the competitive stance of U.S. metal and non-metal mineral materials companies alike against foreign producers in markets at home and abroad.

Jan 1, 2003

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 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

By D. W. Fuerstenau

One of the more energy-efficient methods for size reduction, namely particle-bed comminution under high compressive loads, was employed to grind eight minerals and ores, using up to 250 MPa pressure. A new approach based on the self-preserving character of the ground product size distributions and a linear energy-reduction ratio relationship is proposed for analyzing this mode of size reduction. The grind ability of the material when comminuted in a particle bed is about one-half that of the single-particle crushing mode, which is in agreement with previous estimates of the relative efficiencies of energy utilization based on direct measurements of surface area produced. The loss of efficiency may be attributed to energy dissipation by particle interaction phenomena in compressed beds.

Jan 1, 1993

By J. H. Adkins

The goal of the project reported in this paper was to make recommendations for changes in roof bolting procedures and bolter design that would reduce roof bolter accident rates. To that end the project identified the major hazards associated with each task a roof bolter operator per-forms, normalized task accident frequencies. by the time spent on each task, reviewed equipment design features and operating procedures related to the identified hazards and developed recommendations for equipment and procedural changes that would reduce, if not eliminate, the hazards. The research steps completed during the project were: a. A review of related literature and re-search including roof bolting time studies b. The collection and analysis of 1975 accident data pertaining to roof bolter activities c. The normalization of the accident frequency data on the basis of time/task data d. A review of roof bolting equipment e. The development of a field data collection instrument f. The selection of mines for bolter observations g. The collection of field data h. Data analysis and problem specification i. The development of recommendations for corrective measures

Jan 1, 1978

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 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 L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd's (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine's high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar's top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar's top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By Kent D. Gastreich

As the number of prime surface coal mining sites diminishes with time and with the potential for increased coal demand, areas that were previously not mined because of their difficult mining conditions, marginal economics, and controversial environmental nature, will have to be considered if surface coal mining is to continue in the future. Such sites include areas near major rivers, streams, and lakes. The name most commonly given to these areas by regulatory agencies is "wetlands". The exact definition(s) to the term "wetland" will be discussed later in this paper. Most coal operators are now more than familiar with permitting requirements under the Surface Mining Control and Reclamation Act of 1977 (SMCRA). The SMCRA permanent regulatory program recognizes the fact that coal mining operations will periodically encounter streams, wetlands, etc. Although discussed in various sections of the permanent program regulations, the relationship of coal mining operations to wetlands is particularly evident in the performance standards regarding hydrologic balance and protection of fish, wildlife and related environmental values.

Jan 1, 1980

By R. T. Hukki

The paper describes a two-stage industrial hydraulic classification system including hydrocyclones for primary classification and newly developed cone classifiers for secondary classification. Performance data based on large scale grinding and classification tests are presented.

Jan 1, 1981

By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

By H. Habenicht, E. Urschitz

Three mining methods for full extraction in flat coal seams - i.e., longwall (L W), shortwall (SW), and rib pillar extraction (RPE) - are com- pared with each other in view of the introduction of a mechanized, self-advancing roof support. Features are shown according to which RPE appears most attractive under certain conditions. In this presentation, the mining methods are outlined and discussed. The new support (Alpine Breaker Line Support, ABLS) is described, and its employment is explained. A trial operation in RPE using the ABLS has been conducted successfully in a South African coal mine. The results and improvements are stated.

Jan 1, 1988