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By S. S. Sephton

The La Perla Mine, owned by Altos Hornos de Mexico, S.A., is located approximately 100 miles east of the City of Chihuahua in the State of Chihuahua, Mexico. The mine began operations in 1958. Existing ore beneficiating facilities, consisting of heavy media and jig plants, upgrade the ore prior to shipment by rail to the Altos Hornos steelmaking complex in Monclova. Sizable reserves of two distinct ore types, which were not amenable to treatment in the existing beneficiating plant, or usable in the blast furnaces, have been mined and stockpiled in order to uncover the treatable ores. One of these ore types has been labeled "Diseminado" ore, and is low grade containing approximately 42% Fe with sulphurs ranging from 1 to 2%. The other ore type has been labeled "Pulverulento" ore, and is high grade containing approximately 65% Fe and 0.1% S, but at 80% minus 325 mesh, is too fine grained for efficient sintering. Late in 1968, Arthur G. McKee & Company became involved with the problem of improving the above ores for their efficient and economical use in Altos Hornos' blast furnaces. Early in 1969 representative samples of Diseminado ore were obtained for laboratory bench-scale testing at the University of Minnesota, Mineral Resources Research Center. Concentration tests incorporating magnetic roasting, wet magnetic separation, high intensity separation, high tension magnetic separation, anionic flotation of apatite, cationic flotation of silica and oleic acid flotation of iron were performed. Laboratory testing was concurrently carried on at La Perla Laboratory, which resulted in plans to construct a 10 TPD pilot plant. The pilot plant was operational in September, 1969. Analyzing laboratory results as well as pilot plant results, Altos Hornos and Arthur G. McKee & Company metallurgists agreed that the most promising flowsheet for the low grade Diseminado ores entailed wet magnetic cobbing, desliming, and oleic acid flotation of the non-magnetic fraction. A combined concentrate of

Jan 1, 1972

By Stephen R. Holsinger

During the year of 1972 ASARCO's Silver Bell embarked upon a project that would test a modified grinding circuit using direct digital compute against the performance of its other five grinding lines, implementation and subsequent outcome of that project has been previously reported in a paper by Pitts, et al. The main purpose of the present paper will be to describe major problems incurred during the implementation of this control strategy on the remaining five primary grinding lines and their solutions. Also included will be a discussion of changes made in operating philosophy and methods used since that report was written. Description The Silver Bell concentrator is a 10,436 tonne/day copper sulfide plant supplied by ore from two separate open pits approximately 4 km. apart. As well as receiving ore from these two main pits, ore is received from several exten¬sions of both pits. The ore bodies are variable in rock type and the extensions have increased this variability. Grinding is carried out by six primary, single-stage, 3.20M x 3.66M Allis Chalmers grate discharge ball mills in closed circuit with four Krebs D20B cyclones (Fig. 1). Each of these grinding circuits is controlled by four independent but interactive control elements. These control elements are the fresh ore feed, the water added to the feed end of the mill, the water added to the ball mill discharge sump, and the cyclone feed pump speed. At the present time, instrumentation for measuring primary process variables consists of Merrick weightometers for fresh ore feed, orifice plates coupled with differential pressure transmitters (dp's) for incoming water measurements, back-pressure diaphragm coupled with a dp for estimating ball mill discharge sump level, a magnetic flowmeter for determining cyclone feed velocity and a nuclear density gauge for measuring cyclone feed density. All measuring instruments develop electrical signals that are measured as 2-10 V DC "inputs o the computer interface.

Jan 1, 1977

By C. P. Wa1ters

This paper is written primarily from the standpoint of the miner and consumer of Missouri fireclays in refractory products, principally firebrick used in the linings of blast furnaces, hot metal transfer cars, the recuperators of glass melting tanks, and other allied uses involving contact with liquid metals and reactive slags. For these uses the manufacturing plant requires forecast of certain physical characteristics, as well as chemical characteristics, of the raw clays as a determination of practicability of mining, treatment such as pre -firing or calcining, and final usage area. These methods and results may leave something to be desired from an academic view, but are used my operation as an on-running guide to seeking and mining raw material. The high quality fireclays of Missouri are generally differentiated by the user as the softer, somewhat plastic, "bond" clays and the so-called flint clays suitable for a fired or unfired aggregate portion of a brick composition. The semi-plastic "bond" clays are mined in a producing area of northeast Missouri including Montgomery, Audrain, Callaway, Monroe, and Lincoln counties. In this geographic area the clays are the members of the cheltenham formation and occur in rather large lenses in which the usable members may approach 15 meters in thickness and an area of 6o to 8o acres. Overburden is dependent on the surface topography and, in our operations, varies from 8 meters to 25 meters. It can be up to 36 meters in the prairie uplands, but this is rarely strip mined. There was former mining by shaft and drift methods in which the overlying coal was also worked. These clays are primarily used as a binding material for other clay aggregates and is characterized by its lower refractoriness-cone 3o (165o c) to cone 321/2 (172o c), in terms of pyrometric cone equivalent, and an alumina - Al2o3 - content of plus or minus 31%. Physically, these clays are readily slaked by water to a fine grain size and exhibit a degree of

Jan 1, 1977

By N. Sheikh

In galena flotation, lead ethyl xanthate, Pb(EtX)2, is precipitated and is physically adsorbed in the form of multilayers on the chemically adsorbed first layer of Pb(EtX). 1,2) The solid particles are coated with layers of Pb(EtX)2 in heterogeneous manner over surface areas possessing both anodic and cathodic character (3,4) It is anticipated that the mode of formation of Pb(EtX)2 and the number of ions present, together with stable or unstable by-product species, have great influence on the surface characteristics of galena. The information obtained from a study of these parameters may have important practical applications in flotation. Lead ethyl xanthate is sparingly soluble in water. The metal and xanthate ions released to the aqueous solution during its dissolution undergo various types of reactions depending upon the pH values of the system. Extensive work (5,6) has been carried out to elucidate the mechanism involved in acid decomposition of xanthate ions. Different decomposition paths (7,8) have been proposed for the hydrolytic decomposition of xanthate in neutral and alkaline solutions, The concentrations of lead ions vary due to a number of chemical reactions, regulated by the pH and the temperature of the bulk phase. The hydrophobic character of Pb(EtX)2 may change with the nature of chemical reactions. Some of the numerous by-product species in solution may adsorb on the solid and change the properties of surface film either favourably or adversely. The flotation performance of galena, therefore, may depend to a considerable extent on the physical and chemical stabil¬ity of Pb(EtX)2 layers. Since little work has been reported in literature dealing with the stability of Pb(EtX)2 and its correlation with the flotation of galena, attempts have been made In this study to evaluate the behaviour of freshly formed Pb(EtX)2 in aqueous systems at various pH values.

Jan 1, 1971

By R. T. Swindeman

"INTRODUCTION This paper will discuss the topics in the recently released Conveyor Equipment Manufacturers Association’s (CEMA) Belt Conveyors for Bulk Materials, 7th edition, that deal with safety and maintenance. The most effective way to improve conveyor safety is through designing out hazards in the earliest stages of a project. When the safest way to perform maintenance is also the easiest, mechanics will naturally take the obvious and least hazardous approach. IMPORTANCE OF ACCESS Studies have shown that mechanics spend almost one-third of their time accessing the components to be serviced and approximately 25% of all fatal conveyor accidents occur during maintenance1. Chapter 2 of Belt Conveyors for Bulk Materials deals with design considerations. In Chapter 2 there are recommendations for access at the various sections of a typical conveyor for cleaning, inspection and maintenance. While, in some cases, providing access will increase the cost of the conveyor system there can be significant paybacks in reduced injuries, faster maintenance and easier cleaning. Piping and Conduit One of the significant recommendations in Chapter 2 for access is to restrict the mounting of piping and conduit to the conveyor structure in the Loading and Discharge zones of a conveyor. Piping and conduit are more permanent fixtures than maintainable components and should be located out of the way of components needing frequent maintenance. If electricians and pipe fitters are given convenient locations to make their runs they will natural use them. Flexible conduit and hoses can be used to connect to components from pipe racks and cable trays which further enhances maintenance. Access for Cleaning Another critical aspect of access is for cleaning. Approximately 33% of all fatal conveyor accidents occur while cleaning spillage and carryback under and around conveyors2. In the loading zone having sufficient height under the conveyor for cleaning increases the cleaning interval and allows for safer cleaning. In the discharge zone having adequate access for belt cleaning installation has a direct effect on the ability to install a proper cleaning system and perform maintenance."

Jan 1, 2015

By Marco Aurelio Piangatelli, Massimiliano Bringiotti, Ferdinando De Angelis

"The construction of the three connecting tunnels between the east side and the west side of station C13, adjacent to the Vistula river, has a long and complex history, as they cross beneath one of Warsaw’s main road arteries (Wislostrada). After the attempt to excavate the first tunnel, the face of the tunnel collapsed on the night of August 13, 2012, and a volume of about 10,000 cubic meters of liquid mass (water, silty soil and sand) entered and flooded the west side of the station, which was filled with more than seven meters of material. Fortunately, the workers managed to escape, but machines and equipment were lost. This is the story of how the tunnels were completed, by adopting soil improvement, jet grouting, GRP bars, ground freezing and a careful choice of work steps. INTRODUCTION The Centrum Nauki Kopernik station on Line II of the Warsaw Underground is located alongside to the Vistula river (Figure 1). This station hereinafter is referred to as station C13.The station consist of three parts: the east and west part, and three connecting tunnels 40 meters in length with a large cross section (170 m2).. It includes the rails along the sides and the platform along the centre. The reason for the underground connection of two shafts with a tunnel is due to the presence of a road tunnel (Wislostrada, hereinafter called WS), which must be underpassed by the new metro line (Figure 2).The existing tunnel (Wislostrada) was built many years before, between diaphragm walls, and some foundation structures interfere with the underground tunnel. The challenge was to build the connectors without stopping the traffic of vehicles along the highway above. All works took place in very difficult geological and groundwater conditions due to the presence of Vistula alluvial deposits with poor mechanical characteristics. The soil profile consists of relatively loose sands on top, and silts overlying high plasticity Pliocene clays highly disturbed by the effects of the repeated over-sliding of glacial moraine deposits."

Jan 1, 2016

By E. G. Williams

Underclays and shales are two natural resources associated with the coal measures. Underclays are widely used as the raw material for refractory, semi-refractory, and face bricks. Shales are useful for sewer pipe and lightweight aggregate. This study concentrates on the geologic controls over the properties of shales for use in the manufacture of lightweight aggregate, and the geologic controls over the properties of underclays for use in the manufacture of bloating ladle refractories. The industrial processes are similar in that both involve the trapping of a gas phase to cause bloating. They differ in that the bloating of lightweight aggregate occurs at lower temperatures (1000-1200°C) than ladle bricks (1350°-1470°C). Lightweight aggregate bloats during an initial rapid firing. Ladle bricks are manufactured by slow firing at 1100°C for up to 80 hours. Bloating occurs when these bricks are reheated to 1350°C . Bloating shales are those which expand twice or more times their original volume during fairly rapid firing. Commercial bloating is usually done by rotary kiln processing or sintering; in both cases the firing rate is much faster than the experimental rate (5/8°C/minute). Bloating results from the entrapment of gasses which are evolved during firing. The material must have adequate viscosity to trap gas bubbles throughout the temperature range of bloating. The finished product contains a high percentage of closed pores with a bulk density between 35-75 pounds per cubic foot. Expanded shale finds its most important use as a substitute for heavy concrete aggregate (crushed stone, sand and gravel) where low bulk density is desired property. The extent of bloating in lightweight aggregate is a function of the mineralogy and chemistry of the shale. Figure 1 shows a corner of the three phase diagram, kaolinite-quartz-(illite and chlorite). Several areas on the diagram are outlined which contain predominantly marine shales, fresh and brackish water shales, and underclays. Each sample was rapid fired and the resulting expansion was plotted on the phase diagram.

Jan 1, 1973

By Dan Plazak

Discussion by Dan Plazak The authors are to be congratulated for presenting a simple and practical tool for quick estimates of ground water flow rates. At the risk of detracting from the simplicity of the method, I would like to suggest a slight refinement to the equation as presented. As the authors noted, the Jacob-Lohman equation (1952) assumes confined-aquifer conditions. To use confined-aquifer equations in water-table aquifers, Jacob (1963) presented a correction to convert water-table aquifer drawdowns [Ahw] to equivalent confined aquifer drawdowns [Ah,: 2] [Ah Ah = Ah - w(1) Cw 2H where: Ah, = Drawdown in confined aquifer: Ahw = Drawdown in water-table aquifer and H = Original saturated thickness of aquifer] Substituting the above in equation ( I) of Hanna. Azrag and Atkinson [4nT Ah - 2H ) Qj2.25 T t(2) 1 nr2 S W] Since the above expression predicts lower pump rates, it causes the Jacob-Lohman equation to agree more closely with appropriate analytical and numerical models. Figure 1, adapted from Fig. 5 of Hanna. Azrag and Atkinson, illustrates the closer agreement of the modified equation to their numerical model, as reflected in Qm/QJL ratios closer to unity. The agreement improves particularly as drawdown [Ah] becomes large in relation to aquifer thickness H. An additional advantage of the modified equation is that Hanna, Azrag and Atkinson's correction factor C is much more stable as pit depth increases. Assuming, as did Hanna, Azrag and Atkinson, that their numerical model is an accurate predictor, it is clear that the modification incorporated in equation (2) improves the accuracy of preliminary estimates of ground water flow to a pit.[ ] References Jacob, C.E and Lohman. S. W.. 1952, "Nonsteady flow to a well of constant drawdown in an extensive aquifer." Transactions, American Geophysical Union. Vol, 33, No. 4. pp. 559-569. Jacob. C. E.. 1963, `Determining the permeability of water-table aquifers,' USGS Water-Supply Paper 1536-1, pp. 245-271

Jan 1, 1995

By C. Tsamantaki

Water-bentonite suspensions exhibit a yield stress which is defined as the minimum applied stress for flow. Considerable debate exists about the yield stress being a material property, hence different measuring methods should give similar answers, but several investigators have demonstrated that it is an engineering reality. Yield stress is normally estimated indirectly from the rheogram obtained with a viscometer following an extrapolation to zero shear rate (the dynamic yield stress), but the technique may be prone to errors because of slip flow. Direct measurements, not prone to such errors, rely on an independent assessment of the yield stress, normally carried out with the vane method (the static yield stress). In this paper results are presented and compared regarding the yield stress of 14 different water-bentonite suspensions measured using both the direct and the indirect methods. In order to obtain a wide range of yield stress values, two different types and two concentrations of bentonite at various pH were tested. Rheological data were obtained using a rotational speed viscometer and a yield stress vane measuring device. Measured vane yield stress values ranged from 5 Pa to 83 Pa. Rheological data collected with the rotating viscometer were fitted using various conventional techniques (interpolation, linear and non-linear regression analysis) as well as using a golden section yield stress search methodology based on Herschel-Bulkley model, developed in house. The vane instrument measures directly the yield stress of the suspensions and requires no further data handling. A comparison of the measured values of the yield stress with the direct and the indirect method is then made for the water-bentonite samples. Not very large differences are observed among different techniques and the variations are discussed. Similar data from literature are also analyzed and the comparison shows that differences on the various estimation methods are much larger than the differences observed in this study.

Jan 1, 2006

By G. E. Bandelis

In this paper, solid bed erosion processes and the mechanisms involved when a two-phase liquid-solid mixture flows over a bed of solids in a horizontal annulus are analyzed. There are few such studies available with some authors arguing that the flow rate for erosion is always less than the flow rate for the suspension of solid particles thus necessitating further study and analysis of such phenomena. Tests were run to determine the mechanism of erosion and the erosion velocities of the solid bed formed in a concentric annulus for three flu-ids, water, a low concentration water solution of Carboxyl-methyl-cellulose (CMC), which behaved as a Newtonian fluid and a higher concentration CMC water solution which behaved as a power law fluid. Glass spheres of 2 mm nominal diameter were used as solids. The flow rate at which erosion occurred together with the mechanism of bed erosion were determined visually, while the whole process was video recorded. Our tests indicated different erosion mechanisms and different erosion velocities for the three tested liquids, with the liquid viscosity playing a major role. The first mechanism, observed for water flow, gives erosion from the front of the bed while the second mechanism, observed for the flow of the two viscous fluids, gives erosion from the top of the bed. The role of each of the parameters to erosion is investigated, whether it is turbulent flow with the turbulent fluctuations being more capable to remove the solids from the formed bed or whether it is liquid viscosity, which results in the hydrodynamic shearing force. The results demonstrate that liquid viscosity plays a role in the erosion mechanism and yields different erosion velocities of the solid bed hence, it should be taken into account when modeling flow pattern transitions for solid-liquid flows. Prior studies have not identified this and it could be the reason of the disagreement between theoretical models and experimental and field data.

Jan 1, 2006

By Robert S. Mayo

My subject today is "Drilling-and-Blasting of Small Rock Tunnels". I am sure that there will be some people, primarily college students but perhaps Consulting Engineers, who will say: "Drilling-and-Blasting? I thought that went out with the invention of the Tunnel Boring Machine". A few years ago I talked to Dick Robbins and he told me that they did not try to sell a "Robbin's Mole" unless the tunnel was two miles long. For easy calculations I use 10,000-ft. (3km) A Mole today costs about $2 million which works out to $200 per lineal foot. This is "front money". You cannot place the order unless you are the uncontested low-bidder and then you must wait at least six months for delivery. Of course you can use that time to square up your portals, ready to turn-the-eye, or maybe you have shafts to sink to reach the tunnel grade. But this is "dead-time" and it is a heck of a feeling to have a job and cannot get started on it until you get the Mole. I do not have the American figures on lengths of tunnels but there was an article a few years ago in a British magazine: NEW CIVIL ENGINEER, on this subject. This stated that 75% of the jobs in Britain which were bid were for small tunnels, 10-ft x 10-ft (3m x 3m) or smaller, and that they were less than 5,000-ft (1.5km) in length. This does not mean that 75% of the footage in Britain was less than 10-ft in diameter. It means that 75% of the jobs being bid were less than a mile in length. From my experience in the U.S. I believe that we have substantially the same ratio: so many of the jobs with which I was associated were short, 1,000-ft, 2,000-ft and sometimes 3,000-ft and the bores

Jan 1, 1981

By B. M. Sadashiva Murthy

Increasing quantities of nutrients and pesticides in agricultural runoff in recent years has caused the more developed western nations to become concerned about diffuse agricultural pollution. Thus, the problem of diffuse pollution is of greater importance and concern for any integrated water quality management programme. The above mentioned sources and activities causing pollution to groundwater cannot be completely avoided. The best solution is to entail a firm understanding of the principles of solute transport through porous media. The complex factors that control the movement of contaminants in porous media and the resulting behavior of contaminant plume are usually intricate to assess because of interaction of many factors such as dispersion, advection, sorption, degradation, volatilization, etc. As accurate quantification of solute under field conditions is inherently difficult, the fate of nitrate and phosphate have been analyzed in laboratory column and channel studies. From the analysis of the breakthrough curves, the dispersion co-efficient for nitrates and phosphates through sandy soil was found to be 0.262 m2/d and 0.114 m2/d for column studies, 0.321 m2/d and 0.231 m2/d for channel studies respectively. The column studies revealed that the concentration of nitrates and phosphates increased steadily in the first few days (around 5 days) and then it increased rapidly after 8 to 10 days. However, it attained equilibrium after 20 days. Similarly, the channel studies revealed a steady increase in concentration of nitrates and phosphates in the first few days (1 to 2 days) and rapid increase up to 15 days and there after it attained equilibrium. In the study, one-dimensional analytical model has been developed which include the advection, dispersion, and a characteristic term called ?total elimination rate?. The comparative study of column and channel studies, and model output (in which the term total elimination rate ?K? is considered as zero) exhibited a variation of 30 to 50 % in leaching characteristics of pollutants (nitrates and phosphates).

Jan 1, 2008

By V. D. Ayurov, N. O. Kaledina

The coal mines monitoring should be based on the usage of simple and evident models, letting in the total unity to reflect a structure of organization of mine, system ability of its construction and specific character of its functioning. Such models are offered to create on the basis of extracted hierarchies of objects of mine. All objects of mine are broken on levels of technological hierarchy. For objects of each level of hierarchy there are made balanced models of entering and leaving material flows. Then these balanced models are coordinated among themselves by the help of the inter hierarchical relations. So the structured-qualitative characteristic of the object of monitoring, reflecting in a kind of the table structure of formation of material flows on mine, their intensity and measure of transformation is formed under transition from one level of hierarchy of mine to another. An example of real us- age of hierarchical balanced models on mine is given. The mathematical models of coal mine monitoring should reflect objective laws of their structural organization and functioning. Therefore under the constructing of mathematical models it is important such a way to reflect a structure of the mine organization, that in totality the models has been reproduction of system ability of its construction. The knowledge of the separate parties of functioning of mine, reflected in particular mathematical models, should "to be packed" in combined unity. Or else, they should be connected in system, convenient for recognition by the person. Considering the mine, the expert allocates multi-dimensionality of the objects of making its elements. He always has the knowledge of laws of development of mining work that is connected with pithy stratification of researched object. In dependence of sparial-temporary level of abstracting there are allocated the horizon, wing of a mine field, layer, panel, extracting face, longwall in coal mine. The human thinking on each of these levels operates with a set of concepts peculiar to him, characteristics, definitions.

Jan 1, 1996

By Rong Guo, Qi He, Jinsuo Zhang

"Coal mine ranks No.1 among the 6 major high-risk industries, so it is no doubt that the senior managers in the coal mine bear more stress than other industry executives. The massive working stress results in serious influences on the physical and mental health of the senior managers of coal mine, which might also induce the safety risk in the coal mine. Therefore, to evaluate the senior managers’ working stress and put forward strategies of mitigating the stress is of great theoretical and practical significance in improving the physical and mental health of the senior managers and reducing the safety risk in the coal mine. The research on the working stress has been made in the western countries for almost 50 years, China began the research in the mid and late 1990s, the research is spread across multiple industries including education, medical treatment, computer and aviation. However, there is little research on the senior managers’ working stress of the coal mine, so the available literatures are few. In the paper, we, based on the working stress model of Robbins and through questionnaire survey, analyze the element composition of the senior managers’ working stress of the coal mine; then, we construct the evaluation indicator system of the senior managers’ working stress of the coal mine based on the individual factor, organization factor and environment factor; after that, we determine weights of each evaluation indicators by using the improved triangular fuzzy number weight method, and set up the evaluation model of the senior managers’ working stress of the coal mine on the basis of the fuzzy comprehensive assessment method. At last, in accordance with the model, we evaluate and analyze the working stress for the senior managers of some typical coal mines, and diagnoses the major sources and intensities of the senior managers’ working stress, and then propose corresponding strategies of mitigating the stress."

Jan 1, 2017

By B. K. Parekh, R. H. Mehta

Handling of fine coal is an importance issue for coal as well as the utility industry. Reconstitution in form of pellet or briquet would be desirable if it could be done economically. This paper evaluates the effectiveness of three binders e.g., asphalt- emulsion, corn starch and Brewex, in forming pellets of ultra-fine clean coal. It was found that asphalt emulsion and corn starch were not effective binders for ultra-fine clean coal, however, Brewex provided excellent quality of pellets, which exceeded all the minimum quasi requirements of coal pellets. INTRODUCTION Advanced coal preparation techniques (physical or chemical) are designed to process ultra-fine (minus 200 mesh or 74 pm) coal. The average particle size of the clean coal product is about 25 pm. Currently available dewatering techniques are effective in providing a 25 to 30 percent moisture product of this size coal slurry (Parekh et al. 1989). Handling and utilization of this dewatered ultra-fine size coal is difficult. Reconstitution, which includes pelletization, briquetting, extrusion, etc., of the dewatered coal product provides an ideal answer to handling and transportation problems. However, reconstitution of fines in the coal industry has not been successful due to various reasons. The most important criterion for its unsuccessful application in the coal industry is the cost of the binder (Capes et al. 1965). Various methods of reconstitution, such as briquetting, disk pelletization, extrusion pelletization and roller and die-pelletization, have been tested for fine coal agglomeration (Conkle et al., 1991) (Richards, 1990). Conkle et al., (1992) have provided a detail analysis of each technique in terms of moisture and operating cost requirements. According to Conkle, briquetting requires a product containing very low moisture, whereas pelletization can be done with coal containing up to 30 percent moisture. Currently available dewatering technique provides about 25 percent moisture product, thus, pelletization appears to be an ideal route for reconstitution. In this paper, results are presented on the comparison of three different binders i.e., asphalt-emulsion, corn starch and Brewers condensed solids (BCS or Brewex), for fine coal pelletization.

Jan 1, 1995

By P. J. D. Lloyd

INTRODUCTION The mining of gold commenced on the Witwatersrand in 1884, and reached a mature state by about 1910, when nearly 20 million metric tons were mined. Mining has continued on the central Witwatersrand, the area of the first mining, at a similar rate until recent years. As a result, well over 1 billion tons of gold mining wastes have been dumped along a line over 80 km long, centred on Johannesburg and stretching from Randfontein in the west to Nigel in the east. (see Figure 1). The population of this conurbation, which straddles the line of mining, has grown to about 5 million. In 1922 uranium was first discovered as a constituent of the ore (and thus of the wastes). Investigations in the 1940's revealed that values near Johannesburg were low, but that, at the western and eastern ends of the central Witwatersrand line, and in the new areas centred on the towns of Carletonville, Klerksdorp and Welkom up to 200 km from Johannesburg, values were adequate to permit uranium recovery. Commercial production commenced in 1952, and by 1979 over 100 000 tons of uranium oxide had been produced. There thus arises on the central Witwatersrand nearly 1,3 billion tons of slimes containing an average of about 0,005% uranium and about 1,8% FeS2. The material grading is about 70%-75µm, and the primary mineral is silica (over 85%). The material is contained in dams constructed hydraulically. A typical dam has an area of about 100 hectares and there are over 6000 ha under slimes dams in the area of concern. The treatment of these dams for uranium recovery commenced in 1977. The situation has thus arisen whereby potential uranium ores have unwittingly been dumped in significant quantities within a major metropolis over a period of nearly a century. The purpose of this note is to review the experience arising from the potential hazard over a considerable time.

Jan 1, 1980

By Yoshio Hiramatsu

In Toyoha mine near Sapporo, the ground temperature is very high, perhaps due to inflow of hot mineral water, the origin of which may be rock masses intruded by the igneous activity in the Tertiary period. The rock masses are found about 2 km east of the mine, their temperature being estimated at from 770 to 870 K at a depth of 3 km. In this mine several veins bearing zinc and lead ores have been mined down to the depth of 300 m from the level of the main entry, of which altitude is 551 m above the sea level. Formerly no attempt was made at development of the deeper part of the mine than the 350 m level from the main entry. Let us indicate hereafter the levels in this mine in such a way that the level 300 m deep from the main entry is denoted as 300 mL. Since 1974, however, the 450 mL of Tajima vein has been developed by deepening No. 5 Shaft to, and by driving drifts in the level. Much hot mineral water has been flowing into the drifts from rock. After twenty-eight months from the beginning of development, the total length of excavation amounted to about 1300 m. The success in development so far may be attributed to cooling rock and air with chilled water and to adopting heat-resisting explosives and detonators. This cooling costed very much and for further development in this level a more effective and economical cooling is desired if possible. To obtain fundamental data for designing an economical cooling, the authors investigated, as a preliminary study, the heat balance in the main drift of 450 mL, by measuring the temperatures of air, rock, hot mineral water, the rates of flow of air and water and so on, in August 1976. They are now carrying out continual measurements of the variations in rock temperature, air temperature and humidity with time. The present paper describes the details of the preliminary study.

Jan 1, 1980

By F. C. Nicolli, J. P. Barbosa, P. S. M. Soares, A. V. Estevez

INTRODUCTION Spent household dry cell batteries are frequently disposed of with metropolitan wastestream (MSW) in landfills where they can leak due to handling or aging. Frequent rain and sun cycles, especially in tropical environments, coupled to the action of leachate, dissolve their metal constituents which will eventually reach rivers, lakes and underground water. It is well known that some of these metals are fundamental to sustain life and that is why living organisms concentrate them through their metabolism, but it has been also recognized that beyond specific limits they can definitely be deleterious to the environment. In other words, high concentration of heavy metals are proven to cause chronic or acute diseases and toxic reactions, by acting on mammals' kidneys, liver and nervous system. In addition, some metals frequently used in household batteries, as mercury and cadmium, are said to be carcinogenic (Landis et al. 1995, Silva et al. 1988). Occasionally, plants to treat MSW are coupled to compost production. Household batteries present in MSW can be dragged into these plants, contaminating their products and eventually reaching vegetables making them not proper to consume by people and animals. Alternatively, batteries are also conveyed to incinerators and burned, throwing into the atmosphere volatile species. These can either act as contaminants as toxic vapor species or be delivered by winds and scavenged by rains contaminating soils and water far away from their source. In conclusion, despite the treatment applied to MSW, one can say that battery residues are potentially environmentally hazardous. The careful assessment of the environmental impact caused by residues of household batteries it is not an easy task and it can be referred as a multivariable problem. First, many different household batteries are produced by industry and the potential environmental hazard associated to their disposal depends primarily on the type of battery and manufacturer's technology. These characteristics will dictate how resistant the battery case is to handling and the

Jan 1, 1998

By Donald A. Dahlstrom

INTRODUCTION Solid-liquid separation is an important and many times a critical step (or steps) in a mineral or coal processing plant. Some examples of typical critical separations are as follows: 1. Operating a seven to eight stage CCD thickening circuit to re¬cover almost all of the caustic and alumina trihydrate values in an alumina plant. This must be accomplished with highly scaling slurries, starting temperatures of over 100°C and operate on a closed circuit system. 2. Making a filter step operate to produce a proper moisture for the balling step so that the pellet furnace can produce a proper product. Bentonite addition must be minimized and a good quality filtrate must be maintained to prevent abrasion of the internal parts of the filter. 3. Operating a closed water system in a coal washing plant so that reclaimed water containing less than one percent suspended solids is produced for reuse to maintain beneficiation ef¬ficiency. Refuse solids must also be thickened so that they can be effectively dewatered to meet state disposal requirements. 4. Recovering phosphoric acid values from a slurry containing gypsum and other gangue solids at an elevated temperature. This must be done while minimizing wash water and maximizing P2O5 re¬covery while simultaneously minimizing scale problems. While solid-liquid separations are important to overall performance in the processing plant flowsheet, it is also a significant factor in plant costs. Capital and operating costs for all the solid-liquid separation steps (includ¬ing water reclamation for reuse) in such plants are probably between 10 and 20%. Energy consumption can also be important in these separations. For example, in the copper concentrator, it is the third largest category in energy consumption, after comminution and flotation. Obviously, solid-liquid separation steps in a plant demand proper atten¬tion in design in order to achieve both optimum performance and cost. Simultaneously, knowledge of funda¬mentals are important to improve results in existing plants.

Jan 1, 1986

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