Search Documents

By H Tran

Local communities are undertaking an increased role in mine project decision making processes, associated with growing pressure and expectations from regulators, financiers and the general public. This is particularly evident in international projects within developing economies. Environmental and social issues are often complex. Solutions demand difficult choices and trade-offs between sustainable development dimensions: economic growth, social equity and environmental integrity. Proactive stakeholder engagement can play a critical role in assisting decision-makers in determining the operational reality of sustainable development by jointly exploring project trade-offs between proponents and affected stakeholders. Good stakeholder engagement has other potential benefits, ranging from reduced financial risk due to regulator delays, legal disputes or negative publicity to increasing market share through good public image. Approaches to consultation and disclosure have changed from short-term regulatory based focuses, to longer-term, strategic tools for relationship building, proactive risk management and new business identification. Approaches to consultation form a significant opportunity to contribute to overall social sustainability. This paper discusses, based on authorsÆ experience, the broader concept of stakeholder engagement and its contribution to social sustainability of the mining industry. The focus is on key considerations for an effective stakeholder engagement. Three case studies outline how broad based, proactive stakeholder engagement can contribute to improved operation and overall sustainability. The case studies include: a greenfields project in Ghana; a long-term mine in Australia; and an open-pit mine in Vietnam.

Jan 1, 2009

Induced microseismicity monitoring is: ?Essential tool in improving the geotechnical understanding about a mine and its inaccessible surrounding rock masses. It helps to determine: -Event timing, location, and magnitude - identification of highly stressed areas -Event source mechanism - the physics of fracturing in its seismic source region -Seismic strain-rate distribution in 5D - identification of stress concentration or migration patterns -Geomechanical interpretation - identification of hidden geological structures or weaknesses in the ground -Strategy for mining e.g. by a block caving method - maximization of mine production and to optimize ground support design -Potential of future seismic hazards -Large event re-entry protocol, to improve mine safety ?The norm as part of mine safety maintenance and long-term mine design practices

Aug 1, 2013

By E. J. Ostrowski

The Bureau of Mines experimental blast furnace, Bruceton, Pa., was operated to determine the feasibility of replacing part of the burden coke requirements with solid fuels injected directly into the smelting zone. This report describes preliminary experiments and information resulting from such an interchange. These experiments were made on an all-sinter burden, using two parts sinter to one part fuel. Anthracite fines were injected intermittently into the smelting zone of the furnace by compressed air through auxiliary tuyeres. The interval between injection periods ranged from 5 to 15 minutes. A maximum of 28 percent of the stock column coke bur-den was replaced. This percentage replacement caused furnace malfunction, but the overall percentage of these experiments indicated that 20 to 24 percent of the coke requirements could be replaced in this way. The metallurgical coke consumption was decreased by as much as 377 pounds (25.3 percent) per ton of pig metal produced. Buckwheat anthracite No.4 was used in this experiment; however, such solids as bituminous coal or coke fines also could be used. The frequency as well as the quantity of solid fuel injected has a direct bearing on furnace performance.

Jan 1, 1960

By Alexandr A. Abramov, Svetlana M. Angelova, Oleg P. Maljuk

In this paper the effect of the ratio of grinding and grindable size classes (ore lumps, particles and grains) and the guantity of intermediate size classes in the feed of the primary wet autogenous grinding "Cascade" mill is studied and the features of the influence of these factors on the mill capacity are shown. On the basis of the experimental data from Lebedinsky mineral processing plant, where magnetite guartzites are processed, a coefficient of the "total ratio of size classes" in the feed of 9x3 m "Cascade" mill was estimated. It helps to get an evaluation of the effect of ore size classes ratio in mill media on the mill capacity and to work out some ways for its optimization in closed or semiclosed technological circuites of primary autogenous grinding. The operation of the specific energy effects for some intermediate size classes, removed from the mill as circulating load, is proposed to improve its efficiency before regrinding in "Cascade" mill and decrease negative influence of such classes accumulation in the mill media.

Jan 1, 1995

Profiled here are the eight RMCMI 2001 Scholarship Award winners. Seven of the eight attended the 97th Regular Meeting and Convention at the Steamboat Springs Sheraton Resort in Steamboat Springs, Colorado in lace June. Each spoke during the Scholarship Awards Monday luncheon and shared their goals with attendees. Thanks to the Patrons of the Institute each student received a 52,000 tuition credit, a plaque signifying the honor and the opportunity to thank the membership. The winners were guests of the Institute Sunday through Tuesday noon and had an outstanding opportunity to meet many in the coal industry that could be potential employers. The state vice presidents host the winners anti pay for their travel and hotel expenses. This is a continuing scholarship and may be renewed for tuition credit again in the senior year it the winner is a full time student and is recommended by the dean of the college or university. A resident from each of the eight member RMCMI states is selected by the Scholarship committee from Board approved institutions. Criterion for the awards is chat the recipient be (1) a full time sophomore student who will use the tuition credit in the junior year. (2) a U. S. citizen and a legal resident of the member state. (3) in good academic standing while pursuing a degree in a mining related field. They must have an expressed interest in western coal as a possible career path. Contributions to this worthwhile project are tax deductible to each individual RMCMI member or his company. Through donations of 5100 or more to the Patrons of the Institute, each new Patron receives a plaque and then a year bar for each additional year contributions are made. Contributions may he directed to RMCMI Patrons, 8057 S. Yukon Way, Littleton, CO 80128-5510. [ ] ARIZONA WINNER Steven W. Shipley University of Arizona The son of Bonnie Fawcett and David Shipley, Steven graduated from Catalina Foothills High School, Tucson, Arizona where lie attained a 3.5 GPA. In his third year at the University of Arizona lie is pursuing, a degree in mining engineering. He is a member of SME, Northwest Mining Association, and a volunteer at San Xavier Mining Laboratory -Miner. He participated in Football, Track, Wrestling and the DECA club while in high school He is the recipient in both the WAAIME and the Phelps Dodge scholarships. Steven volunteers during the academic year at the U of A San Xavier Mining Laboratory. He worked the summer of 2001 at the Black Butte Coal Company in Point of Rock, Wyoming. [ ] COLORADO WINNER Brian Gerdes Colorado School of Mines Graduating as the Valedictorian of Horizon High School with a 4.0 GPA, Brian is well prepared for the rigors of academia. In high school he was recognized in rite President's Education Awards Program and as a Radio Shack Tandy Scholar. lie belongs to SME, is a member of Fellowship of Christian Athletes, and volunteers at a local elementary school tutoring children. He enjoys intramural Sports. His father is Bruce Gerdes of Broomfield, Colorado Brian is holding a 3.75 GPA while studying for a mining engineering degree. He would like to go into reclamation scaring that his dream would be to reclaim an old mine and restore it to a function, al role in society again.

Jan 1, 2001

The South American mining community leads the world in recognition and adoption of the latest grinding mill relining technologies. Clearly this region has recognised the benefits of improved relining performance as a defence against the Global Economic Crisis. Five of the six RUSSELL TWIN 8 Mill Relining Systems, ordered to date, are either operating or being manufactured for South American mine sites. The sixth is for Sino IronÆs Cape Preston Project in Western Australia. This paradigm shift, towards very short relining times, aims at displacing mill relining activities from the critical path of concentrator maintenance shutdown schedules. The first uses of the RUSSELL TWIN 8 Mill Relining System have seen some of the highest semi-autogenous grinding (SAG) mill new liner placement rates ever observed by Russell Mineral Equipment (RME), as would be expected with two, independent Machines working simultaneously in the one mill. The liner replacement rate, averaged over the whole reline is good (27 per cent improvement over single Machine System in the same mill) but short of full potential. This modest result can be attributed to reline crews unfamiliarity with RUSSELL TWIN 8 System operation. This paper compares a sample of conventional (single machine) RME Mill Relining System performance data with one ædouble headÆ mill liner handler by others. Early performance of EscondidaÆs new RUSSELL TWIN 8 System performance is then compared with recent single machine data from Escondida, immediately prior to RUSSELL TWIN 8 commissioning. RME and Escondida look to constantly improving RUSSELL TWIN 8 System performance as full operational familiarity is achieved. The preparation of this paper revealed a need to develop a global industry standard for mill relining data collection. A web based platform to develop data collection templates is also discussed in Appendix 1.

Jan 1, 2009

By Frédéric Flament, Mary Goldman, Gille Barbery

"Pilot plant testwork on autogenous and semi-autogenous grinding was conducted at the Centre des Recherches Minérales du Québec (CRM) to investigate the effect of ball addition, solids concentration in the feed, specific energy consumption and feed size distribution.A mathematical model that incorporates breakage and surface mechanisms of loss of mass was developed to predict load size distribution of autogenous and semi-autogeneous mills. An empirical model was used to predict product size distribution of autogenous and semi-autogeneous mills.IntroductionAutogenous (AG) and semi-autogenous (SAG) grinding circuits offer potential capital and operating cost savings when compared with conventional grinding. Empirical relationships have been developed to calculate specific power consumption for SAG mills using pilot plant test results (Mular et al., 1989; Barrett and Allan, 1986). Empirical models can be obtained from plant practice data but are applicable only to the particular ore and set of operating conditions used in the model. Grinding mechanisms have to be taken into account in order to develop a model that can be applied to different ores and mills.In AG and SAG, the ore itself acts as grinding media, with the finer sizes being crushed by the coarser ones. Surface mechanisms of loss of mass, in which the inner core of the particle remains intact, also take place. In SAG milling, the effect of balls also has to be taken into account. The role of each one of these mechanisms changes with particle size and the model has to account for such changes."

Jan 1, 1991

By H. Y. Bassett

THIS paper attempts to show the various major operations used in nonferrous tube mills and does not necessarily represent the current practices at the two plants of the Wolverine Tube Div., of Calumet and Hecla Consolidated Copper Co. Inc. The operations now utilized in the production of tubing are quite varied and show the continual interest and advance metallurgically and mechanically to produce the best quality at the lowest cost. Naturally, any advance or lowering of cost is dependent on large production runs, which mean longer machine running time with maximum worker productivity. Standardization of tube sizes and elimination of unnecessary alloys are therefore very necessary, otherwise the development of costly new production machines would be of little benefit to the consumer as a whole. It should be noted that in such a short presentation it is impossible to describe in any detail the major operations. The short descriptions, it is hoped, will enable all who are interested to obtain a general idea of present, day practices in the copper tube producing field. Fig. I is a flowchart or sequence of operations showing the steps usually employed in tube production. RAW MATERIALS Quality starts with the raw material, and it is absolutely necessary that raw material be controlled and held to rigid specifications. Little need be said, however, regarding the raw materials for the casting of bars or billets, as the material always is purchased and checked against the various governing A.S.T.M. specifications. CASTING Casting, as is indicated in Fig. I, is quite varied and for convenience can be classified into two groups: The casting of solid billets, and the casting of hollow tubes, each of which is the base for subsequent operations. In the melting of the raw materials various types of furnaces are used, and the selection of any one is dependent on many factors outside the scope of this paper. However, as of general interest, a short description will be made of each. Induction Furnace (Low Frequency) This is popular in the melting of copper-base alloys, Fig. 2. In the induction furnace the principle of an inefficient transformer is used for the production of the necessary melting heat. The primary coil is.arranged around a transformer core made of high permeability silicon steel sheet. The secondary circuit is through a melting channel around the primary coil and opens into a hearth which contains the main charge. The heat is generated only in the melting channel, but because of the stirring action there is a constant flow of metal around the channel, which when mixing with the charge causes very rapid melting. The volatilization losses are very low as the furnace atmosphere above the charge is heated only by the heat from the melting metal. The stirring action, however, makes the use of any slag cover difficult. Such furnaces- must be started with molten

Jan 1, 1951

By R. W. Brown, J. L. Loper

The process of sand erosion has been harnessed to perform a useful function—the directed perforating of oilfield tubular goods and formation rock. In this process the sand is carried by a liquid medium and the resrrlting slurry is forced through a nozzle at an extren~ely high velocity. Upon impingement against casing or formation, the fast-moving sand particles erode a perforalion in the casing followed by cavity formation in the medium behind the casing. The many factors influencing the rate and extent of this penetration are inherently difficult to establish experimentally. Equations describing submerged jet behavior have been applied to these conditions. The following paper develops the equations for predicting the performance of the perforating jet. INTRODUCTION The cutting power of high-velocity sand has been well known in the petroleum industry for many years. Sand carried by a high-velocity gas stream causes severe erosion. This effect has been the cause of numerous production problems in certain areas. It was logical to expect that this phenomenon would one day be adapted to useful industrial applications. Development of mobile, high-pressure, high-horsepower pumping equipment in the last decade has made this possible. The sand erosion process involves pumping a fluid containing abrasive solids through a set of nozzles at a high differential pressure (2,000 to 4,000 psi). The pressure conversion into kinetic energy imparts high velocity to the sand particles which, upon impact with the formation face or casing wall, erode the material in an organized pattern. This paper attempts to establish the effects of the various parameters and to predict the performance and limitations of the process. The following factors will be considered: nozzle differential pressure, perforation size, sand concentration, nozzle stand-off distance and backpressure. In general, this discussion is presented along theoretical lines and is substantiated by field results and data where reliable field data exists. POWER EQUATIONS The penetration rate of a high-velocity hydraulic jet is proportional to the power or energy of the jet at the point in quesion. The power of the submerged jet will decrease with distance (or penetration) into the cavity created by the jet. As the cavity is extended into the formation, the power of the jet diminishes to a value equal to the threshold cutting power of the particular formation which is defiined as the lowest power level at which the formation will be abraded by the jet. Obviously, when this point is reached, no further cutting of the formation can occur. It is assumed that this power level for a particular formation is a function of the hardness of that formation. The following equation can be written for the hydraulic process of formation abrasion.

of our good friends and fellow members: Patrick J. Quealy J. C. Roberts Gilbert Morton Sidney W. Farnham who left the world better for their having lived in it

Jan 1, 1930

By Stephen C. Tadolini, Thomas M. Barczak

An underground investigation was designed and implemented by the National Institute for Occupational Safety and Health (NIOSH) to evaluate the effectiveness of pre-driven longwall recovery rooms supported with pumpable concrete roof supports. The pumpable roof supports were instrumented with hydraulic pressure cells to measure support loading. Displacement of the support was also measured to assess the support reaction to ground movements. Movement up to 12.5 m (40 ft) into the roof was measured with sonic probes and multipoint extensometers. Borehole pressure cells were installed in the panel fender and outby pillars to evaluate load transfer. Shield loading was also assessed to determine the impact of the panel fender yielding on face loading. The results show that the yielding of the panel fender produced uncontrollable convergence that caused yielding of the shields and stiff pumpable roof supports when the face was approximately 3.12 m (10 ft) from the recovery room. The shields continued to yield until the face advanced into the recovery room. Convergence in the pre-driven recovery room typically reached 15-20 cm (6-8 inches) once the pumpable supports yielded and shed load through a declining residual load capacity prior to being cut out by the longwall shearer. Roof deformations occurred beyond the 10 m (32 ft) horizon with total deformation ranging from five to 10 cm (2 to 4 inches), indicating that standing support was necessary to help control the span as the longwall face advanced into the pre-driven recovery room. Stability improved as the pumpable supports were cut out, thereby reducing the span between the shields and the outby pillars. In the end, the shields were successfully recovered under stable ground conditions. However, a disconcerting discovery was the load shedding of the outby pillars as the recovery room was mined into. This behavior was not expected with a relatively narrow room that is fully supported with standing support. If this mechanism is indeed occurring, the width of the room and performance of the standing support becomes even more critical. Premature failure of the support can lead to excessive convergence on the shield line that cannot be controlled even with modern shields, resulting in instabilities that can lead to catastrophic results. Understanding the load transfer mechanics associated with the advancement of the longwall face into the pre-driven recovery room and the widening of the room to recover the longwall shields is providing valuable information for the refinement of the NIOSH Ground Reaction Curve for standing and intrinsic roof support system design.

By Walter E. Lewis

Ventilation-air-heating plants are in use at several iron mines on the Menominee Iron Range, Iron River, Mich. Many of the mines are exceptionally wet; during winter months, when fresh air is forced down the wet air shafts into the mine, it is usually necessary to heat the fresh air to above freezing so that the openings will not be closed by ice. The problem of ice closing air shafts has been solved by the practice of reversing the air current before the openings have become completely closed. This procedure involves drawing warm air from the mine up the usual down-draft air shaft. The M. A. Hanna Co. has found, at some of their mines on the Menominee Range, that it is more convenient to heat the below-freezing air rather than reverse the air current. Intermittent freezing and thawing of ice that has formed in the air shaft tends to loosen the wall rock, requiring major shaft repairs. The openings are kept free of ice the year around by heating the fresh air during periods of below-freezing weather. Table 1 shows the monthly minimum moan, maximum mean, and mean temperatures for the Iron River area. Under normal conditions of temperature, the furnaces are fired in November and usually are not allowed to go out until April of the following year.

Jan 1, 1951

By Robert D. Schmidt

This Bureau of Mines report describes the development and application of a computer model for simulating the hydrological activity associated with in situ leaching. The model is intended to provide uranium resource developers with a description of the flow behavior of leachants and ground water during the development, production, and restoration phases of a leaching operation involving an arbitrary pattern of injection and recovery wells. Different aquifer environments are modeled, using a closed-form solution to the partial differential equation that describes three-dimensional changes in piezometric head as a result of pumping from leachant injection and recovery wells. The computer program can model a maximum of, 50 arbitarily located wells. Numerical techniques involving difference quotients and Taylor expansions about time points are used to derive time, velocity, areal sweep, and fluid volume parameters associated with leaching hydraulics. These parameters are output by the program in graphic and tabular formats. Other numeric methods insure that the program running time is minimized without significantly affect-ing the accuracy of results.

Jan 1, 1980

By Sungdon Kim, Chuljoo Kim, Joonsoo Kim, Hosung Yoon

"This study was performed to establish the optimum leaching condition through the experiments for the oxidized roasting followed by leaching of fluorocarbonate rare earth mineral (bastnasite) using sulfuric acid, and to establish the optimum condition for the recovery of cerium by ion- sieve method from the leached solution of bastnasite. The optimum conditions for the sulfuric acid leaching by the oxidizing roasting of bastnasite concentrates are that the roasting temperature is 550 -600 °C, the roasting time is 2 hrs, the concentration of sulfuric acid is 6N, the leaching temperature is S0°C and the leaching time is 3 hrs. At this time, the leaching yield of total rare earth oxides in the leaching of bastnasite concentrates is about 91%. When cerium is recovered by ion-sieve method from the solution obtained by the sulfuric acid leaching of the roasted bastnasite, the optimum conditions are that the TREO(total rare earth oxide) of leached solution is below 70 g/l, the flow rate of leached solution is 9.6 ml/min and the ratio of height to diameter of resin packed in column is above 9. In these conditions, the recovery of cerium of yield 85% and purity 96% is possible.IntroductionThe rare earths are generally called to Lanthanides, which are totally 17 elements including Scandium of atomic number 21 and Yttrium of atomic number 39 in addition to 1 S elements from Lanthanum of atomic number 57 to Lutetium of atomic number 71. Since the separation to each element is very difficult because of their similar chemical properties, the rare earths were used as the mixed rare earths and the mischmetal without separating to each element. However, the high purity rare earths are recently produced commercially to each element so that they are used as the materials for high tech. industry such as electric and electronic materials, catalysts, optical materials, special alloy and 8Uperconductive materials, and their consumption is rapidly increasing. Therefore, the purpose of this study is firstly to prepare leached solution by using the series of process of the oxidation-roasting and leaching by sulfuric acid. Next, second purpose is the separation and recovery of cerium by ion-sieve method."

Jan 1, 1998

By W. J. Wenger

Results of the analyses of 151 crude oils from 123 oilfields in the Williston basin are presented. Production statistics are compiled by geographical areas within the basin, by geological age of the producing zones, by States, and by field for the nine fields with the greatest production. Characteristics of the oils are tabulated, and correlations are included to show relationships between properties of the oils and some factors in the environment. Properties of the oils in some instances imply that there has been migration and mixing of the oils between productive formations of different geological time periods.

Jan 1, 1968

By W. P. Sykes

SEVEN years ago, Dr. S. L. Hoyt1 presented a masterful discussion of the hard metal carbides and cemented tungsten carbide. His lecture summarized most of the data then available in the field; many of the details relating to the production, properties and structure of cemented tungsten carbide had never been previously published, for, in fact, this information was only then being acquired. During the intervening years, the cemented carbides have assumed a place of real importance in the commercial field, with an ever widening variety of applications due to constant improvement in the control of production technique and the unremitting search for new compositions that might better the performance of the alloys in one or another type of service. The most recent summary of production methods, properties and guides for the successful application of the cemented carbides appears in the 1936 edition of the Metals Handbook. Of necessity, much detailed information relating to structures and physical properties is not included in such a summary, and it is with such details that the present discussion is concerned. STEPS IN PREPARATION While the past 25 years have contributed a valuable store of working knowledge in the field of powder metallurgy, many of the process controls are derived entirely from experience and are not as yet finally explained by a specific physical or chemical phenomenon. For most of the cemented carbides such difficulties are aggravated by our lack of knowledge not only of the complex alloy systems involved, but also by the possible changes in structure and even in composition which may accompany variations in temperature, heating time, and furnace atmospheres. The principal steps in the preparation of these cemented carbide alloys are fairly well known, but to supply a foundation for discussion these steps are briefly outlined as follows: 1. Formation of the metal carbide in the form of powder by heating a mixture of carbon and the metal powder or oxide for several

Jan 1, 1938

By J. W. Pamplin

Some 40 miles east of Tampa is the center of the Florida pebble phosphate deposits. These are of Pliocene age and consist of several members of the Bone Valley formation.1 Physically the phosphate-bearing formation, locally called the "matrix," varies in thickness from nothing to about 25 ft., usually averages 10 ft., and consists mainly of clay, quartz and phosphate pebble, the latter particles varying in size from colloidal to one inch diameter. The pebbles range in color from dark gray, through white, to cream, and in hardness from that which must be broken with a hammer to that which can be scratched with the fingernail. They assay from 66 to 80 per cent bone phosphate of lime (called "B.P.L." in the trade) or tricalcium phosphate (Ca³P2O8) and are secondary apatites. They form from 15 to 50 per cent by weight of the matrix. Over the matrix lies a varying thickness (5 to 40 ft.) of overburden composed of quartz and clay. Briefly, mining consists of two operations2 in open pits about 200 ft. wide by 1/2 mile or more in length. In the first step overburden is removed by large electric draglines at the rate of about 700 cu. yd. per hour. This is followed closely by hydraulic mining of the matrix at the rate of about 300 cu. yd. per hour. This matrix yardage is moved from an improvised sump in the floor of the pit to an ore-dressing plant, sometimes a mile distant, by one, two or three "pit-cars" (pumping units) working on a 12-in. pipe line called the "rock line" to distinguish it from the high-pressure water line feeding the hydraulic giants, which is known as the "hydraulic line." This 700 yd. of overburden and 300 yd. of matrix—1000 yd. in all-amounts to about 1500 tons of material moved each hour. The overburden is dumped into the mined-out pit alongside the pit being mined. The matrix moves to the ore-dressing plant at the rate of about 7% long tons in 4000 gal. of water per minute—a ratio of approximately one to two by weight, or one to six by volume. This constitutes the "mill heads," but in the pebble district it is called the "matrix."

Jan 1, 1938

By J. W. Pamplin

Some 40 miles east of Tampa is the center of the Florida pebble phosphate deposits. These are of Pliocene age and consist of several members of the Bone Valley formation.1 Physically the phosphate-bearing formation, locally called the "matrix," varies in thickness from nothing to about 25 ft., usually averages 10 ft., and consists mainly of clay, quartz and phosphate pebble, the latter particles varying in size from colloidal to one inch diameter. The pebbles range in color from dark gray, through white, to cream, and in hardness from that which must be broken with a hammer to that which can be scratched with the fingernail. They assay from 66 to 80 per cent bone phosphate of lime (called "B.P.L." in the trade) or tricalcium phosphate (Ca³P2O8) and are secondary apatites. They form from 15 to 50 per cent by weight of the matrix. Over the matrix lies a varying thickness (5 to 40 ft.) of overburden composed of quartz and clay. Briefly, mining consists of two operations2 in open pits about 200 ft. wide by 1/2 mile or more in length. In the first step overburden is removed by large electric draglines at the rate of about 700 cu. yd. per hour. This is followed closely by hydraulic mining of the matrix at the rate of about 300 cu. yd. per hour. This matrix yardage is moved from an improvised sump in the floor of the pit to an ore-dressing plant, sometimes a mile distant, by one, two or three "pit-cars" (pumping units) working on a 12-in. pipe line called the "rock line" to distinguish it from the high-pressure water line feeding the hydraulic giants, which is known as the "hydraulic line." This 700 yd. of overburden and 300 yd. of matrix—1000 yd. in all-amounts to about 1500 tons of material moved each hour. The overburden is dumped into the mined-out pit alongside the pit being mined. The matrix moves to the ore-dressing plant at the rate of about 7% long tons in 4000 gal. of water per minute—a ratio of approximately one to two by weight, or one to six by volume. This constitutes the "mill heads," but in the pebble district it is called the "matrix."

Jan 1, 1938

By Peter Cunningham

Teck Facts and Figures ? Employ more than 14,000 around the world, includes: ? Over 8,600 in Canada (close to 7,800 in British Columbia) ? Close to 2,000 in Chile ? Over 2,700 in Peru ? Approximately 600 in U.S. ? Hundreds more in exploration around the world ? Total capital expenditures ? $1.8 billion for 2012

Aug 1, 2013

By Robert J. Ramsey, Robert D. Jeppson

Introduction The Utah Copper Division of Kennecott Copper Corporation will present its contribution to the A. M. Gaudin Flotation Symposium in four parts. The first two segments will discuss briefly the historical development of the Utah Copper Division and describe the ore body. The third segment will review the facilities and operating procedures for the Utah Copper division's newest plant, the Bonneville Concentrator. The fourth segment describes flotation process circuitry, including molybdenite recovery procedures, at the Arthur and Magna concentrators. (The Bonneville concentrator is a crushing and grinding unit and does not have its own flotation plant). I. Historical Development Progress through change has been the key to the successful operation of the Utah Copper Division for more than 70 years. As the original low-grade copper operation, the pressures to remain competitive and profitable have dictated the timing and extent of each change. Much of the change has occurred through technological and economic advancements in the area of ore concentration. The Utah Copper Division of Kennecott Copper Corporation had its origin June 4, 1903, when the Utah Copper Company was organized. The initial mill of 300 tpd (tons per day) capacity was built at Copperton a few miles below the present Bingham Canyon mine site. From 1904 to 1910, the Copperton mill was expanded to a capacity of 1,000 tpd, and all of the known apparatus for gravity concentration were evaluated.

Jan 1, 1976