Search Documents

By T. C. Rao, N. Suresh, M. Vanangamudi

Experiments have been conducted on a 76¬mm water-only cyclone treating cowl below 2 mm size having different proportion of -0.075 mm fines. The effect of cone angle of the cyclone (compound cone) has been analyzed on water split, separation of -0.075 mm fines, separation of coarse solids, and on separation behavior of individual size ructions of the feed. It has been noted from the results that there exists a linear relationship between the flow rate of water in the feed and flow rate of water reporting to overflow stream. This relationship is independent of the cone angle and, with changes in the fines content in the feed, only the intercept of the straight varies, keeping the slope constant. Further, it has also been found that the percentage of fines below 0.075 mm reporting to the overflow is directly proportional to the percentage of feed wafer reporting to the overflow. This relationship is found to be independent of the cone angle and the fines in the feed. The performance of the water-only cyclone has been analyzed by plotting partition curves for different experimental conditions. In addition to this, the Rosin-Rammler equation has been fitted to the partition curves of the hulk feed and individual size-fractions. Tile analysis of the results has revealed that the specific gravity of separation increases with increase in the cone angle of the cyclone while it decreases with increase in percent fines present in the feed. Mathematical relationships have been developed to relate the variation in SG50 with cone angle and percent fines in the feed.

Jan 1, 1990

By M. Georgescu

In the framework of the rapid development of contemporary society, the use of underground space represents an emerging opportunity because it can contribute to decreasing pressure of occupation of ground space while in the same time it can increase the economic value of the useful mineral substances, also through underground establishment of dangerous surface activities, contributing thus in environmental protection. This paper underlines the most important benefits of underground space utilization with the objective the exploit the potential of this real resource.

Jan 1, 2006

By K. Hanumantha Rao, K. S. E. Forssberg

The limestone deposit of Gâsgruvan, Sweden contains around 89.2 per cent of calcite. Conventional and column flotation studies have been carried out using sodium oleate and modified fatty acid collectors in the presence of sodium silicate modifier. The effects of collector and silicate concentrations, and particle size on the flotation response of calcite have been investigated. The results have been discussed in comparision to the performance of conventional and column flotation cells. The feed size is found to be a significant factor to achieve a high grade calcite concentrate either by conventional or by column flotation. The loss of selectivity at fine particle sizes is examined using a modified DLVO theory.

Jan 1, 1995

By R. A. Jankowski, R. Gong, R. Bhaskar

Full-scale laboratory studies were performed to evaluate the effect of passive barriers and spray airmovers on airflow patterns at a continuous miner face and in reducing underboom boilout and operator exposure to dust. Tasks included baseline data collection, smoke trace tests, detailed performance testing using tracer gas, and data analyses. Data collection was performed using a state-of -the-art gas analysis system and custom-developed software. The data were used to calculate operator exposure at various locations and to develop gas profile maps to analyze airflow patterns. The effects of the proposed underboom airflow modification devices on operator exposure and airflow patterns were examined. Based on the results obtained from the study, the base case condition, with no barriers, appears to result in the lowest operator exposure.

Jan 1, 1992

By Jean-François Bougard

It has already been said and written many times that the construction of tunnels in urban areas has proved more and more indispensable, if not vital. The subsequent development of tunnelling for 15 or 20 years has brought large and sometimes spectacular improvements in the fiel of execution methods. But, particularly in urban areas, the difficulties to be overcome have been and still remain very important. As a matter of fact, most of big cities, have been settled near rivers or lakes; therefore the geological nature of the subsoil is soft ground and the level of ground water is high. As a consequence, tunnelling is often faced with many well-known difficulties and dangers issued from decompression or loosening of the ground, sometimes accompanied by a lowering of the water table. This may compromise the safety of the excavation itself and of the workers, and also cause settlements which many severely damage nearby buildings or structures. This is the reason why it is always necessary to work out specially adapted methods and even to develop new ones. In most cases, each work is a particular one and has to be specifically studied since numerous data are to be taken into account, related not only to geology, cohesion and behaviour of the different layers concerned, ground water, the activity of it, but also to

Jan 1, 1979

By Robert J. Timko, R. Lincoln Derick

This report presents a Bureau of Mines evaluation which examined the extent of gas leakage from sealed areas in underground coal mines and measured the effectiveness of commonly used structures for controlling the quantity of gas exchange from these areas. Evaluations were performed on five abandoned areas in two operating underground mines within the same coal seam. Methane and oxygen levels inby the seals were monitored to determine if the gas leakage was unidirectional or bidirectional. Results from this and prior work dealing with minimizing gas leakage were then applied to a newly abandoned area. Also discussed are bulkhead site preparation, construction plans, techniques for monitoring the status of the atmosphere within the sealed volume, and a modified European method for controlling gas migration from the sealed volume to the active mine.

Jan 1, 1991

By M. Imazu, Ö. Aydan

"A new rock classification, Rock Mass Quality Rating (RMQR) proposed by the authors is briefly explained and its application for the design of tunnels and large underground openings support systems is described in this paper. RMQR is an index to evaluate the state of rock mass. The preliminary support design may be based on the concept to prevent structurally controlled failures of rock mass. RMQR can also evaluate the yielding characteristics of rock masses so that it is possible to design a support system to prevent or control the deformation of surrounding rock through the utilization of the proposed rock mass classification and some analytical and numerical techniques.INTRODUCTIONThe first three authors proposed a new rock classification system named Rock Mass Quality Rating (RMQR) (Aydan et al. 2014). RMQR quantifies the rock mass state, and using the classification system together with intrinsic geo-mechanical properties of intact rock, possible geo-mechanical properties of rock masses can be estimated. The instability modes of underground openings can be structurally controlled because of the orientation of discontinuities in rock mass and the in-situ stress induced (Aydan 1989; Aydan et al. 1993, 1996; Aydan and Genis, 2010). The authors (Aydan and Ulusay, 2013; Aydan, 2016) have also tried to apply this system to the design of a tunnel support system. The authors presented initial guidelines for the RMQR classification system earlier (Aydan and Ulusay, 2013).In this paper, the authors provide further guidelines and establish some criteria for support system design that considers loading conditions and types of underground structure according to RMQR. Guidelines are alsos proposed for selecting the proper size of rockbolts, rock anchors, shotcrete, steel ribs and concrete liner. These guidelines are based on past experiences with various underground structures, theoretical understanding and the anticipated behavior of rock mass under given in-situ stress conditions and excavation techniques. The authors also utilized some databases developed for large underground caverns and tunnels excavated in squeezing or rock-bursting ground conditions."

Jan 1, 2016

By William S. Brack

The mining industry is very capital intensive. Competition for shareholder investment has never been greater. Among our challenges is the need to meet shareholder expectations (return on investment), and one area of focus lies in our ability to deliver value on our capital investments. Our ability to meet project expectations affects not only the economics associated with the immediate project but also directly influences our ability to finance future projects. Moreover, our ability as individual companies to attract investor capital is influenced by how we have done overall as an industry, and our industry track record has not been that stellar. With safety, scope, cost, schedule, and quality as guiding principles, our project managers today must address high expectations on every project. This paper examines some items that should be on today's project management checklist. PROJECT ENGINEERING Herbert Hoover, himself a mining engineer, said Engineering is a great profession. There is the fascination of watching a figment of the imagination emerge through the aid of science to a plan on paper. Then it moves to realization in stone or metal or energy. Then it brings jobs and homes to men. Then it elevates the standards of living and adds to the comforts of life. That is the engineer's high privilege.

Jan 1, 2004

By J. M. Dean, R. L. Grayson, N. P. Reddy, Y. J. Wang

This 1988 survey provided a cross section of computer software use by the US coal industry. From 1250 mailed questionnaires, 171 responses were received, representing a 4.9% sampling of the 3496 US coal mines. Since the distribution of responses by regional stratum matched well with the distribution of mines by stratum, the survey was given statistical validity. A statistical description details the computer use by region, mining type, organizational size, the level of application within the organization, the type of software application used, the variety of computers on which the software was used and the percent of users within engineering, management and general applications categories. Overall, 63.5% of the respondents use computers at some level in their organizations. All of the mines producing more than 454 kt/a (500,000 stpy) used computers. However, for those producing more than 9 kt/a (10,000 stpy) but less than 91 kt/a (100,000stpy), only 44% used computers. Generally, the software applications were still data-processing oriented. The application sophistication grew with increasing mine size. Computer system models and configurations for different sized organizations are recommended based on the configurations used in the industry at the time of the survey.

Jan 1, 1994

By S. Majumder

The longwall face supports are provided to prevent the caving of the immediate roof at the working face. The thickmess of the immediate roof may be obtained from the working height and the average bulking factor. A strong bed, capable of withstanding maximum bending moment among the beds within the above roof, can be identified. The load on the strong bed may be obtained by finding out the overlying beds that will rest on the former during sagging. The limiting span, L, for this unsupported loaded bed, as a beam with built-in ends, can be calculated. The support density required at the supported span (face to goaf edge), a, during the first fall, is now estimated with a view to have a safety factor 2 at the face end, when bending failure occurs at the barrier end. However, the above value is inadequate to provide necessary safety margin against bending failure at the face end, during the subsequent cavings as cantilever, for L/a less than 20.6. d For the latter case a separate expression is derived.

Jan 1, 1986

By Charles D. Litton

Introduction The successful performance of a mine-wide fire detection system depends upon its ability to detect the presence of a fire rapidly and reliably. It should be sufficiently rapid so that there remains enough time to either safely evacuate a mine or successfully extinguish and control the fire or both. It should be reliable so that false alarms are minimized without sacrificing sensitivity or time response. Systems should also be durable enough to withstand the mine environment over long periods of time while maintaining a high degree of re1iability. Maintenance and calibration of the system should be simp1 e and not require the expenditure of excessive time and manpower to keep it operational. Sensitivity, reliability, and durability of mine fire sensors and their electronic support systems have been the subject of significant Bureau research within recent years. The bulk of this type of information, along with the reports from which the remainder of this paper is derived, can be found in the Bibliography. It suffices to say that there exist commercially available systems for use in mines. This paper addresses a strategy for the use of such sensors and systems. Many factors influence the design of a mine-wide fire detection system, such as the potential sources and modes of ignition, the types of combustibles involved and the quantities avai1able for fire growth and flame spread. Mine air ventilation influences the growth and spread of a fire and also serves to transport the products-of-combustion to other areas of a mine remote from the fire. Fire sensor alarm thresholds place limits on the sizes of fires that can be detected. The time to respond to a detected fire, either in terms of evacuation or control and extinguishment, place additional time constraints on the design of fire detection systems. These factors interact with each other, often in complex ways, and it is the definition and understanding of these factors and their interactions that hold the key to the design of adequate and re1iable fire detection systems. Our level of understanding of the problem has increased dramatically within recent years and it is the intent of this paper to discuss the progress that has been made and how the information can best be used to improve the level of fire detection in underground mines.

Jan 1, 1986

By Ian M. Loomis, Malcolm J. McPherson

A fuel-rich fire exists when the quantity of fuel exceeds the oxygen available to support complete combustion. The transition of a mine fire from Oxygen- rich to Fuel-rich combustion, at this time, represents a point-of-no-return in the fire fighting efforts. Once the transition has been made to fuel-rich the fire can spread at a rate 6 to 10 times as fast as in the oxygen-rich state. Utilizing available technology the fire fighters are left to seal the fuel-rich fie as the only technique to extinguish it. In order to minimize the devastating effects of a fuel- rich fire it is important to understand the means by which the fire becomes fuel-rich and by what means the fire may be returned to an oxygen-rich state. This paper covers: the development mechanism of a fuel- rich fire in a coal mine entry and the physical and chemical effects of applying water to a fire. Experiments concerning these matters have been conducted in a 30 cm square wind tunnel constructed at the Virginia Polytechnic Institute and State University as a model coal mine entry. This paper also addresses the results obtained in testing application of a water mist to fuel-rich fires in this model. These tests have centered on the development profile of a fuel-rich fire in a duct and demonstrate a means of regaining control of a fuel-rich fire by returning it to an oxygen-rich state.

Jan 1, 1995

By D. R. Lowe

Silicic volcanic ash and its alteration product, bentonite (montmorillonite and cristobalite), occur widely within sedimentary phosphate deposits. The genetic link, if any, between volcanism and phosphate deposition has yet to be satisfactorily established The phosphatic member of the Santa Margarita Formation (upper Miocene) in northwestern Ventura County, California, consists of approximately 15 per cent volcanic ash and bentonite. Ash and bentonite units greater than one foot thick can be subdivided into a lower zone of airfall origin and an upper zone of mixed ash and terrigenous detritus washed in from surrounding land areas. This upper zone also contains up to 20 percent phosphate pellets. These thick volcaniclastic units are typically overlain sharply by thin beds (less than 6 inches thick) of relatively pure, silica-cemented pelletal phosphorite. The shelf basin in which the phosphatic sediments accumulated was characterized by quiet, reducing but not euxenic bottom waters; Abundant decaying organic debris, mainly diatoms, served as a reservoir of readily available phosphate. The sudden influx of a large volume of silicic volcanic ash into this basin and its subsequent alteration to montmorillonite had two profound effects on thechemistry of the interstitial and bottom waters: (1) during alteration magnesium was removed from these waters and incorporated into the montmorillonite lattice, and (2) the pH of these waters increased as alteration progressed Both of these effects led to the crystallization in the bottom waters of carbonate fluorapatite. This accumulated as thin phosphorite beds immediately above the volcaniclastic units.

Jan 1, 1972

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By S. L. McGill, P. B. Comba

Many gold producing states are requiring rinsing of spent heaps to render the leached ore compatible with the environment and are placing reatrictiona on effluent pH, cyanide, and metals concentrations. Parameters which may affect efficient rinsing of heaps have been investigated by the U.S. Bureau of Mines in laboratory column studies. Rinse profiles for cyanide, metals, and pH as a function of the volume of water passed through the spent ore are presented. Fresh water rinsing of spent ore in columns showed that rinsing of cyanide was not affected by the rinse solution application rate. Interrupted rinsing with a rest period between applications reduced the volume of water required to meet effluent cyanide concentrations. Chemical oxidation of the effluent with recycle of the rinse solutions reduced effluent cyanide concentrations and the volume of waste solution generated. Water rinsing was not successful in reducing the alkalinity of the cement agglomerated ore. A dilute acid solution had to be applied after cyanide removal to reduce effluent pH levels to specified limits.

Jan 1, 1991

By John R. Alderman

INTRODUCTION Uranium production in Saskatchewan commenced in 1953 at the Eldorado Nuclear Mine in the Beaverlodge area of North Saskatchewan, followed over the next few years by nineteen other smaller mining operations. Production rose from 453 600 to 2 721 600 kg per annum by 1958. Over supply and the closure of the U.S. market to Canadian uranium led to an equally rapid decline to an annual production which varied from 453 600 to 907 200 kg during the years 1964 to 1976. As the market strengthened there was a resurgence of the Saskatchewan industry and production from one underground and two open pit mines reached 2 809 000 kg in 1980 for a return of 232 000 000 dollars (Table 1). [ ] Future output will be increased by a potential 4 536 000 kg from the Key Lake mine now under construction and scheduled to commence production in 1983. Three other projects are in an advanced stage of planning and are currently in the regulatory review process. Several other prospects are being evaluated. Figure 1 shows the location of uranium mines in Saskatchewan and table 2 illustrates the present status of uranium mines and deposits in Saskatchewan. At present, 1400 workers are employed directly by the uranium mining industry. Saskatchewan has diverse mining conditions ranging from the underground mine of Eldorado Nuclear Limited 1870 m deep with a low grade ore deposit averaging from 0.17-0.20% with extraction by carbonate leaching through an open pit operated by Gulf Minerals with a moderate depth of 110 m and 0.37-0.5% grade to the shallow open pit of Cluff Mining 30 m deep with a very high average ore grade of 7.0%. Extraction at both open pits utilizes acid leach extraction processes, the latter after gravimetric separation. Key Lake mining has commenced development of the next mine to come into production which will have two open pits to a depth of 80 and 175 m with an average 3% grade to be followed by the deep open pit of Canada Wide Mines Ltd. to a depth of 205 m with average grade of 1.5%. One existing mine will transfer its operation from the present open pit to another of slightly higher average grade. From the Saskatchewan point of view, these different operating conditions present varied problems associated with occupational health and safety and concomitant solutions to reducing worker exposure to radiation. REGULATION OF URANIUM MINES In 1946, the Government of Canada enacted the Atomic Energy Control Act which established the Atomic Energy Control Board (AECB) and vested in that board powers to regulate all aspects of uranium mining. In the early years, the AECB relied heavily on provincial occupational health and safety agencies for the surveillance of uranium mining and milling operations, particularly with regard to "conventional" occupational health and safety, the AECB having at the time no specific regulations covering the mining industry. There has been considerable dialogue between federal and provincial governments since, traditionally, occupational health and safety is a provincial jurisdiction. Saskatchewan is strongly opposed to the intrusion of the federal government into the provincial domain, particularly as there is little justification for a federal presence in occupational health and safety in uranium mining and milling. The province does, however, accept the need for federal involvement in certain aspects of the regulation of nuclear industry, for example, international sales, research and security of reserves for domestic use.

Jan 1, 1981

By A. De Daio, D. W. Fuerstenau, P. C. Kapur

Because of its pronounced tendency to undergo visco-plastic flow, coal ground in the energy- efficient high-pressure roll mill invariably comes out as a briquetted product which must be disintegrated in a conventional type of grinding device such as the ball mill. The aggregates comprise comminuted coal particles that are apparently extensively cracked and weakened during passage of the coal through the high-pressure rolls. Their deaggregation in a ball mill in the second stage is accompanied by further grinding of the coal with a relatively small expenditure of energy. Thus in spite of energy dissipation in briquetting in the roll mill and in disintegrating the briquettes in the ball mill, significant energy savings are possible in the fine grinding of coal in a two-stage serial grinding mode, as compared to ball mill grinding alone. It has been determined that for the best results the energy invested in the two grinding stages must be apportioned in an optimal manner, the energy consumption in the ball milling step is comparable to, even in excess of, the energy spent in the high- pressure roll mill; and 20 to 50 percent energy savings might be possible in grinding coal to minus 200-mesh size by this two-step process, as compared to conventional milling.

Jan 1, 1995

By Cathleen J. Webb, Arden D. Davis

Begun in 1992, an abandoned and inactive mine inventory in the Black Hills National Forest has resulted in a hazard ranking of the most severely affected sites. For purposes of the inventory, the Black Hills were divided into three main regions: the northern Black Hills, with primarily gold mines; the central Black Hills, with pegmatite mines and some gold mines; and the southern Black Hills, with primarily uranium mines. A computerized data base has been developed along with standard forms for field crews. The forms were designed to provide maximum compatibility with other abandoned mine inventories on various federal and state lands. Measures recommended for the most hazardous and heavily affected sites include regrading of unstable slopes, fencing of open shafts, and remediation of the most severe acidic mine drainage problems. Specific examples are illustrated where remediation and safety measures were recommended or have been implemented. The vast majority of the sites, however, needed little or no remedial action.

Jan 1, 1995

By N. E. Green, W. E. Bennett, G. P. Harrison

INTRODUCTION The art of raise boring has seen rapid development and use in underground mines both for primary development and production during the past 10 years. The intensive full scale tests by a major mining company in the Precambrian rocks of the Sudbury district of Ontario, Canada, have lead the way to the general acceptance of the raise drills. The rock successfully raise bored in the Sudbury area has compressive strengths varying from 30,000 to 50,000 p.s.i. As a result of the intensive study, it became evident to many mining companies that raise boring could be adapted to many mining systems. Many companies have modified their planned mining methods to incorporate raise boring. The result has been a more rapid development of their mine and naturally, an economic advantage to them. In conjunction with the acceptance of the raise drills, the range of equipment has broadened in terms of size, capability and efficiency. The concept of raise drilling and the general arrangement of the equipment has seen little change. At some locations there have been reported successes of drilling pilot holes up

Jan 1, 1997

By R. H. King, R. E. Cameron

Mine managers need better information to locate new opportunities and then to assess the effect of their decisions. We are developing the Intelligent Decision Support System (IDSS) and its two major modules, the Pattern Recognition System (PRS) and the Knowledge Representation Subsystem (KRS) to automatically provide the information from monitoring system sensors. This paper focuses on the KRS Algorithms that extract key information from lists of time-study type event sequences. Because a static system cannot operate effectively in a dynamic mining environment, we are prototyping an index- based technique that allows the system to adapt itself to this dynamic environment.

Jan 1, 1991