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By Norman Weiss

The purpose of this paper is to relate progress in the engineering field called Minerals Beneficiation to undergraduate and graduate education in our colleges and universities. So mach has been written recently about the education of scientists and engineers that I've had to struggle against the temptation to repeat what has been often said before, particularly because man like Robert Mehl of Carnegie Tech, Antoine Gaudin of MIT, H. M. Davis of Penn State, and John Vanderwilt of Colorado School of Mines have recently written many quotable paragraphs on the subject. In the matter of progress, the rapidity of proliferation in the area of physical and chemical beneficiation is convincing proof that this is indeed a lively art, and having been a millman for 35 uninterrupted years I prefer to lean on personal observations to some extent, rather than on statistics and charts v culled from the journals. I propose in this paper to talk first about the progress we've made in our branch of the mining industry in the last 3 decades--particularly in the last few years--and try to show that future progress will depend upon the quality of education and capacity for research possessed by our colleges. The second part will deal with some of the problems the colleges encounter in trying to fulfill these responsibilities.

Jan 1, 1959

By D. D. Carda, S. Dunn, W. E. Dunn, T. Storbeck

Gold chlorination which will be presented in this paper is limited entirely to the use of gaseous chlorine in a fluidized bed reactor at elevated temperatures in the absence of aqueous water. The only water used in this process is that which is used in scrubbing the effluents. Low temperatures are defined as those below 500°C. All experimentation for this paper was carried out on a laboratory scale, with all calculations extrapolated frpm this. Some of the data is drawn from work with other projects which have been operated at pilot and larger scale.

Jan 1, 1998

By T. B. Jarboe

The 150 MWe circulating fluidized bed (CFB) power plant described in this paper is the product of recent Bechtel experience in the industrial power and steam market, where the economic feasibility of fluidized bed projects has been proven. The inherent ability of CFB boilers to burn low grade fuels makes it possible to also burn low ranking coals or coal wastes, including anthracite culm, coal mine rejects, and coal wash plant waste slurry. The impact of fuel characteristics on plant design is addressed. This includes solids handling, steam generation, sulfur capture, and auxiliary load requirements.

Jan 1, 1988

By Keith P. Bingham

A total transportation system providing for delivery of finished products to customers and raw materials to the plant is one of the key elements contributing to Dravo Lime Co.'s success. Dravo Lime operates one of the largest noncaptive lime and limestone plants in the US at Maysville, KY, where products produced can cost as much to deliver as they do to manufacture, depending on the product itself, its destination, and mode of transportation. With delivery costs comprising a significant portion of total costs for its products, Dravo Lime has made total transportation services an integral part of its business strategy. Prices for most products produced at Maysville are quoted f.o.b. destination. In short, Dravo Lime's Distribution Department relies on the approach that says to customers: "Leave the delivery to us." The 0.9-Mt/a (1-million stpy) Maysville facility, located about 121 km (75 miles) east of Cincinnati, OH, on the Ohio River, primarily produces sulfur-absorbing Thiosorbic lime used by coal-fired electric utilities as a scrubber reagent to remove SO2 from stack gases. The product is currently supplied to eight power plants in the Ohio River valley region which extends through portions of Kentucky, Indiana, Ohio, West Virginia, and Pennsylvania. The plant ships approximately 907 kt (1 million st) of lime and a comparable amount of limestone fines each year. In addition, it arranges for delivery of about 181 kt (200,000 st) of coal to the plant. Lime is transported by covered hopper barges, covered rail cars, covered dump trucks, and pneumatic trucks. It is transferred from barges to storage silos and trucks at a terminal on the Ohio River near Marietta, OH. At another terminal near Aliquippa, PA, lime is transferred from barges to silos, trucks, and rail cars. Future plans call for rail-truck transfer terminals. To produce lime from limestone at the Maysville plant, about 181 kt (200,000 st) of coal is burned annually in three coal-fired kilns. In transporting this coal from the mine to Maysville, Dravo takes maximum advantage of opportunities for reloading carriers with limestone for outbound backhaul movement. Coal, shipped from southeastern Kentucky mines in rail cars, is transferred to barges at a river terminal in West Virginia. After it is unloaded at Maysville, the barges are cleaned with a front-end loader and then reloaded with limestone. Limestone also is backhauled in connection with truck deliveries of coal. The economics of using backhaul shipments has permitted Dravo to enter into markets in which Maysville limestone otherwise would not be competitive. The company plans to backhaul limestone in empty open hopper cars in the future. As part of its ongoing assessment of the distribution system that supports the Maysville operation, Dravo Lime recognizes that a new environment has developed between shippers and carriers because of regulatory changes in the railroad and trucking industries. Although dry bulk shipments by barge to and from Maysville have always been unregulated, other forms of transportation have not. With railroads and motor carriers now approaching varying degrees of deregulation, contracts and agreements are replacing tariffs, and rate negotiations are becoming a daily activity. Dravo Lime plans to take advantage of these transportation changes, all of which may be beneficial to both its plant and its customers.

Jan 10, 1982

By R. S. C. Rogers

Semiautogenous milling was the subject of 18 papers presented at the SME Fall Meeting in Albuquerque, NM. The process design and scale-up of SAG mills was discussed in several papers. These included a review of testing methods for the sizing of SAG mills by Barrett and Allan. Austin et al. presented an improved simulation model for SAG mill grinding, with model validity demonstrated against pilot plant SAG mill test results. Rogovin et al. presented some pilot plant results relating to the charge motion, hold-up, and particle size segregation in SAG mills. A number of papers described improvements achieved in operating SAG mill circuits. These included presentations by Dailey concerning mills at the National Steel Pellet Co., by Brown and Pipke concerning the Island Copper grinding circuit, by Lovering of Afton Mines Ltd., and by Bartrum et al. of Kidston Gold Mines Ltd. Bush and Hegerle of Amax also discussed SAG mill operations at the company's Henderson mill. They addressed the topics of SAG mill grinding economics. Other papers addressed the topics of SAG mill control (Vanderbeek et al. and Pennington), grinding media (Meulendyke et al.), liners (Kjos), and fabrication (Carr and Utley). The 1985 AIME Annual Meeting in New York featured four technical sessions concerning size reduction technology and practice. Papers involving fundamentals included one by Hinde describing simulation model usage for quantifying the classification and concentration of heavy minerals in grinding circuits. Another, by Rogers et al., discussed simulation models that were used as a design aid in evaluating screens versus hydrocyclones for fine grinding circuits. Arbiter discussed dimensional considerations in grinding mill design, and Schwechten and Schonert proposed the use of a high pressure roll mill as a means for reducing energy requirements in fine grinding. Internal classification in tumbling mills was the subject of a paper by Hogg and Rogovin. The mechanism of internal classification were identified and their effects illustrated. Simulation for the process design of ball mills was exemplified in a paper by Rogers, Austin, and Brame concerning phosphate grinding in laboratory to large diameter mills. The incorporation of liberation into size reduction models was the subject of papers by Austin et al., Herbst et al., and Choi et al. Gupta used simulations models to further investigate zero order fines production in ball and rod mills. Of the published literature, many articles on size reduction were published in various publications in 1985. Articles concerning ball mill grinding included an analysis of interstitial gas effects in dry grinding by Cottaar et al., an analysis of mill grinding noise by Watson, studies of factors affecting breakage rates in dry milling by Gupta et al., and in wet milling by Tangsathitkuchai and Austin. Austin and Klimpel presented two papers concerning ball wear and ball size distribution effects. Koka et al., and Gotsis et al., described recent developments concerning the Szego mill. A model for hammer milling circuits was described in papers by Gotsis et al. and Gotsis and Austin. Heim et al. reported on wet grinding in Perl mills, and Kuga et al. presented findings concerning pulverizing in a screen mill. Studies of grinding effects on particle surfaces were published by Meloy and Hofmann et al., and Scieszka proposed a technique to investigate pulverizing properties of coals. Austin and Rogers discussed factors of powder technology of significance in industrial size reduction in mills and crushers. (Space limitations preclude listing of references. A complete list is available from author.)

Jan 5, 1986

By A. Tsouchlaraki

The physical relief is a variable of utmost significance in the course of the analysis, evaluation, design and management of the visual environment. In this sense, there are cases where it would be desirable to isolate the relief from the other elements making up the visual environment and to investigate its visual value separately. Such cases include the technical works that cause major and permanent alterations to the relief, as well as the works, of which the spatial arrangement depends on the morphology of the ground. Slopes, distances, hypsometric difference values, viewsheds are elements of the relief that affect the visual value of a landscape, but also the ability of a particular landscape to accept and absorb new activities. This paper describes an evaluation and prediction model of the visual value of the landforms. A questionnaire survey takes place in order to investigate the public preference, using a sample of 221 persons in the broader area of Athens. The means used to demonstrate the various forms of reliefs are 32 perspective digital relief images. Following this and in order to investigate the elements of the relief that affect its visual value, 20 quantitative indices were developed and examined so as to select the best ones. Linear prediction models of the visual value were investigated from the previous elements, through a regression process. The applications of the proposed model in physical environment planning are many, such as the prediction of visual impact by arranging works that alter the relief of a region, as it happens in mining, metallurgical and waste disposal sites.

Jan 1, 2005

By Charles F. Herbert

Placer mining in Alaska had a desultory inception at Sumdum in 1869 and, during the years 1880 to 1896, received thin nourishment from the readily exhausted gold placers at Juneau and the longer- lived deposits in the Circle, Fortymile, and Rampart Districts. Finally, in 1897, the great Klondike discovery in Canada's Yukon Territory brought thousands of gold seekers to the Far North. These men, in successive, far-reaching gold rushes, found the large, rich gold deposits at Nome, Fairbanks, and Iditarod, as well as many smaller placers. The hectic rushes during the years from 1898 to 19 10 were comprised of hundreds, sometimes thousands, of "stampeders," as they were called. Since very many Alaskan creeks, from the crest of the Brooks Range in the north to the Pacific Ocean in the south, carry at least a few flakes of gold, there was ample seed to start a new stampede. Of the 100s of reported "strikes," nearly 50 became more or less important placer mining districts; by 1910 they had produced about 184.2 fine Mg (6.5 million fine oz) of gold; the all- time peak production year was 1906. A few of the miners and prospectors prospered, but the total value of all gold produced failed, by a wide margin, to cover the aggregate costs incurred by them in the undeveloped Territory of Alaska. On the other hand, the traders with their readily salable supplies of food, tools, and hardware and the freighters with their teams of dogs or horses and boats of all kinds profited throughout the era of stampede.

Jan 1, 1991

By William L. Meyer

With high expectations and promises, computers were introduced to the mining industry in the late 1950s and 1960s. Applications spread in the 1970s when capabilities began to catch up to early promises, and people within the industry began to see the benefits of using computers in business and engineering. The 1980s have seen a wide acceptance of computers, both large and small, and the use of computers is accelerating throughout the industry. Computers have changed dramatically over the last 25 years. The power of some microcomputers matches the capabilities of main frame computers from less than 10 years ago. The variety of software on these micros exceeds what is available on many larger computers today. Changes over the next 10 years are likely to be greater than those of the last 10 years. Making use of available computer technology while trying to cope with the rapid changes in this industry will take much effort within mining. To take advantage of existing computers, consider and evaluate: project needs and applications; professional, administrative, and support personnel; computer software and technology; and computer and related hardware. Note that this list is in the order of decreasing importance. These categories are also presented in the order for which organization and solutions should be provided to get the most out of computers. To start with a computer and go up the list has cost many mining companies needless expenditures, wear and tear on personnel, and frustration. It is assumed for this article that the need for computer tools is self evident. You can do without a computer by using the appropriate techniques (manual methods) with any required shortcuts plus a lot of personnel time. Indeed, several companies are doing that as a "solution" to what to do with their excess engineering personnel in these lean times. However, the future may be better served if some of these people could be prepared to handle the applications and technology presently available. People and Problems To present the factors related to turning the computer into a productivity aide, it may be enlightening to examine some common people problems that reduce productivity. The most valuable resource any mining company has is its mining professionals. And, particularly with computers, people either make the successes or create the barriers, not the machines. The following list of people problems (syndromes and complexes) has been amply observed during some 23 years of the author's experience with computers and their applications in mining. Home Grown Syndrome - "Nobody else can really understand my problem or provide `exactly' what I need to solve it." While there may be some truth to this syndrome, it would be an unusual project that cannot be handled with existing solutions for at least 75-80% of the needs. Having a base to build on will also speed the development of any remaining needs. It does not seem justified to start all over for a new application, even if "our problem is unique." Not Invented Here Syndrome - This "justification" for the continued existence and budgets of some groups has various forms: "It wasn't done here so it obviously can't do the project the way I think it should be done"; and, "If it wasn't invented here, it obviously

Jan 11, 1983

By L. R. Duncan, W. H. McCracken

Magnesium is the eighth most abundant element in the earth's crust and the third most plentiful element in seawater. It is found in more than sixty minerals and in brines and seawater as a magnesium salt. One of the more important magnesium minerals is magnesium carbonate, MgCO3, with a theoretical maximum magnesia, MgO, content of 47.6%; this carbonate form represents the world's largest source of magnesia. The next most used sources for magnesia are magnesia-rich brines and seawater, where it occurs as magnesium sulfate and magnesium chloride. Other commercially important magnesium-bearing minerals are dolomite, CaMg(CO3)2, which serves the aggregate industry as well as the chemical industry; brucite, Mg(OH)2, which is used in the production of both caustic and sintered magnesia; olivine, ([MgFe]2-SiO2), which serves the refractory and heat storage industries; talc, (H2Mg3(SiO3)4), which serves several industries as a filler and as an ingredient in cosmetics; and serpentine, (H4Mg3SiO9). All these minerals are of commercial value because of the desirable characteristics given to them by magnesia and by its placement in the crystal structure. These minerals are the starting raw materials for a wide range of products. These include the production of magnesium metal, and several grades of magnesia used for the production of both sintered magnesia for refractory manufacture and lighter fired caustic magnesia, the latter is used in fluxes, fillers, insulation, cements, decolorants, fertilizers, chemicals, in the treatment of waste water including pH control, and in the removal of sulfur compounds from exhaust stacks. Magnesite is one of several major minerals possessing a dual character, i.e., non-metallic industrial mineral because of its principal end use as a calcined/dead burned raw material in non-metallic end products and magnesium metal. A large percentage of the magnesia used to produce magnesium metal comes from seawater1 brines. The total primary metal production worldwide is about 350 ktpy. This report reviews magnesite solely as an industrial non- metallic raw material. Magnesite for most refractory purposes must have a content of sintered magnesia above 85%, with a preference for those sources that have the potential for 90 to 98% magnesia values. Dolomite, used in large tonnages because of its physical properties in the construction industry, also has a use in the chemical and the refractory industry. In its calcined form, dolomite is used as a slag addition in the making of steel, and it is used in a high-purity calcine to precipitate magnesium hydroxide from brines or seawater. The precipitated magnesium hydroxide is intermediate for the production of deadburned clinker for refractories and the production of magnesium metal. In many cases the high-purity magnesium hydroxide obtained by precipitation with calcined dolomite, or in some cases limestone, has supplanted much of the world's production of magnesia from the natural mineral magnesite. This is especially true of the United States and in major industrialized areas of Europe and Japan. The best known of the minerals directly and widely exploited for its magnesia content is magnesite (MgCO3), one of the calcite group of rhombohedra1 carbonates, which includes calcite (CaCO3), siderite (Fe2CO3), and rhodocrosite (MnCO3), among others. The members of this group enter into a wide range of substitutional solid solutions when the positive ions have similar radii. The radii of magnesium and iron ions are within 6% of each other; hence magnesite and siderite form a complete series, of which the mineral breunnerite (ferroan magnesite) is a well known end member. The radii of calcium ion is 36% larger than that of magnesium ion and only limited substitution exists at each end of the MgC03-CaCO3 series. Dolomite is not a member of the calcite group, but it occurs when calcium and magnesium ions alternate in equal number in an ordered structure among carbonate ions. The result of this relationship is that calcite and dolomite are commonly found intermixed with magnesite. They occur, commonly, as identifiable crystal entities, which can be separated to a varying degree from the magnesite by benefication techniques. Magnesite, when pure, contains 47.8% MgO and 52.2% CO,. The pure mineral is found occasionally as transparent crystals resembling calcite. Magnesite principally contains variable amounts of the carbonates, oxides, and the silicates of iron, calcium, manganese, and aluminum. Although the genesis of natural magnesite deposits can be complex, it is distinguished in nature in two distinct physical forms, namely crystalline, (with a wide range of visible crystal sizes) and cryptocrystalline, sometimes referred to as amorphous, where the crystal size is not detectable to the eye and will range from 1 to 10 pm. The two types not only differ in crystal structure but in the sizes of the deposits and in modes of formations. The crystalline form has a Mohs hardness of 3.5 to 4.0. The color range is from white to black with shades of yellow, blue, red, or gray. The color is not a significant indicator of purity but in a given deposit an experienced person can often roughly grade the magnesite by observing color and crystallinity. Macrocrystalline deposits occur in relatively few, but generally large, deposits on the order of several million tons. The ore shows a marble-like crystal structure and belongs to the sedimentary or metasomatic groups of origin. The cryptocrystalline variety of magnesite frequently occurs in many small deposits, although there are exceptions. Cryptocrystalline magnesite is typically massive with no cleavage and is sometimes descriptively called bone magnesite. The fracture is usually conchoidal, with a hardness of 3.5 to 5.0. The color is mainly white, but it can have tints of yellow, orange, or buff. Accessory silaceous minerals such as serpentine, quartz, or chalcedony are usually present. Calcium minerals are normally absent or in low concentration, this contrasting with the macrocrystalline form where calcium minerals are almost invariably present and some- times in high concentration. The specific gravity of cryptocrystalline magnesite ranges between 2.90 and 3.00, whereas the value of pure crystalline magnesite is 3.02. In actuality, the normal specific

Jan 1, 1994

By Sandy Clamage

"Seeing the elephant" was a term often heard from 1849 to 1859, the period of the great California Gold Rush. The phrase was used by those who dreamed of instant wealth that was believed could be found in California's virgin earth and mighty rivers. When news of fabulous gold strikes throughout the Mother Lode reached far away places, thousands of gold seekers left home and family to join in this incredible adventure. The barren, placid state of California soon became one of the most bustling, populous areas in the West. Before the Rush Before 1848, Mexico ruled California. It was then a remote region, isolated from the urbane world of the East by 2900 km (1800 miles) of scorching desert and impassible mountains. By sea, it was 29 000 km (18,000 miles) via Cape Horn. Spaniards and Mexicans settled the coastal lands on tracts donated to them by the Mexican government. The inland area was the domain of Indians and a few settlers who also had been granted estates by the government. Among those settlers was Captain John Sutter. He had been fortunate enough to obtain about 11 leagues of land (spanning 100 km or 60 miles in length) from Governor Alvarado of Mexico. Sutter was somewhat of a visionary and possessed a gifted tongue. He managed to receive most of what he desired with minimal trouble. Enchanted with Sutter's personality, Alvarado promised him Mexican citizenship. Alvarado also commissioned him to become "representative of the government and officer of justice on the Northern frontier." With uncanny foresight, Sutter began building a fortress on his land in the Sacramento Valley. It later became known as Sutter's Fort. It was more than just a fort, however. It also became a stopping place and home for many foreigners and Americans who traveled to California, and it eventually became the mecca for overland pilgrims. Sutter engaged in many projects. These included a cattle ranch, fur trade and trapping, and the construction of an irrigation system. With the influx of immigrants over the years, the community around the fort grew rapidly. Lumber was much in demand and Captain Sutter agreed to supply the necessary capital for a community sawmill. The site chosen was some 72 km (45 miles) northeast of the fort on the south fork of the American River in the town of Coloma. During the course of its construction, Sutter's carpenter, James Marshall, discovered a problem with the tailrace on the mill. First Strike Early in the morning on Jan. 24, 1848, Marshall was checking the tailrace. He noticed an object shining in a foot of water at the bottom of a ditch. He reached down and picked it up. His heart began to thump wildly as he realized he was holding a gold nugget. Sutter planned to keep the find a secret for as long as possible. He knew very well that he could become an extremely wealthy man. The ever loyal Marshall remained faithful to his employer and the pair did their best to obtain more

Jan 3, 1985

Hoses provide a variety of critical functions in nearly every mining operation. Care must be taken when selecting, installing, and maintaining hoses to insure maximum hose life. This article provides some tips on selecting, installing, and maintaining hoses, and gives some background on hose construction and safety. In the mining industry, hoses are used for everything from hydraulic systems that power drills and hammers, to suction systems that carry away dust and debris. Hoses are used on such a large scale that a detailed discussion of specific applications would require many volumes. Dayco Corp. has, however, provided these basic and brief guidelines on how to select and maintain hoses for mining applications. Hose Construction Industrial hoses are classified as much by construction as they are by application. So, it is a good idea to start any discussion of hoses with a description of their basic construction. There are four basic components to hoses: the tube, reinforcement, cover, and couplings. The tube is the inner core or lining, and is usually made of smooth rubber or plastic. Rubber or plastic composition is scientifically determined to be compatible with the materials passing through it, such as fluids, chemicals, gases, or substances. Other considerations in compounding rubber or plastic for the tube are temperature ranges and stresses. In practice, the required compound can be precisely chosen from various charts listing them according to application. There are, however, applications where the compounding of tube rubber or plastic must be customized. The reinforcement hose layer immediately surrounds the tube. As its name implies, this layer gives the hose strength and provides protection for the tube component. Reinforcement must be able to contain the inner forces-pressure and shock-transmitted through the tube, and such external forces as twisting, bending, and crushing. So it is usually made of high tensile strength cord, fabric, or wire that is wrapped, wound, or braided around the tube. The cover is the outermost layer, so it may encounter oils, acids, water, dust, abrasion, and other potentially damaging influences depending, of course, on the application. An appropriate cover material must be selected, usually a rubber compound. Couplings, even though they are a hose attachment, are as vital as any other component and should be considered part of the hose. Couplings are usually made of carbon steel or malleable iron; however, bronze or stainless steel might be used if corrosive chemicals are involved and brass or aluminum might be used for flammable fluids. A coupling is usually in two parts: an insert that fits into the hose bore and contains the threaded male, female, or other type of connection, and the ferrule or collar that fits over the hose and secures the coupling in place. There are, however, some one-part couplings that fit securely into the hose without a ferrule. These are usually limited to low pressure applications and are secured to the hose by clamps or metal bands. Some fittings can also be permanently "built-in" as the hose is made. Couplings can be classified as reusable or permanent. The reusables can be removed when the old hose expires and used again on a new hose. The permanent types are crimped or swaged on at the factory or in the field (using special equipment) and must be discarded with the hose. Although couplings are most commonly thought of as having threaded fittings, there is another type called "quick connect." These either have a crow-foot or jaw-type design that permits a fast, "snap" connection. This can

Jan 6, 1982

By John H. Bassarear

How best to grind? This article compares conventional crushing and grinding circuits to semiautogenous circuits, reviews recent installations of each circuit, and describes seven methods of using autogenous or semiautogenous techniques. In addition, steel and power consumption data are reported and benefits of a dedicated digital computer are discussed. Size reduction is the most expensive operation in most mills. Crushing and grinding are usually the greatest part of capital costs. They often make up 60-70% of total mill operating costs. Picking the proper crushing and grinding method must, therefore, be done with great care, so the best circuit is chosen. In the past 10 years, there has been a strong trend toward using autogenous or semiautogenous grinding in mills throughout the world. Autogenous grinding is favored when the ore is quite competent and a fine grind is needed. Semiautogenous grinding is applied when fine crushing could cause severe problems or when ore varies in hardness or competency. The taconite iron ores of northern Minnesota fit the description for an autogenously ground ore and many are successfully ground that way. Most im¬pressive are the nine lines of 11 x 4.9 m (36 x 16 ft) mills powered by 9 MW (12,000 hp) motors at Hibbing Taconite. Porphyry copper and molybdenum ores fit the description for semiautogenous grinding. Recent installations include Highmont in British Columbia, Canada, and Quintana in New Mexico. Sampling Many mills have been built based on inadequate sampling data or insufficient tests. With the cost of many mills exceeding several hundred million dollars, it is mandatory that geologists, mining engineers, and metallurgists work together to prepare representative testing samples. Simple, repeatable work index tests are usually sufficient for rod and ball mill tests. However, pilot plant tests on 50-100 t (55-110 tons) of ore are frequently necessary for autogenous or semiautogenous mills. Preparation and selection of the test sample is of utmost importance. Procedures for autogenous and semiautogenous mill pilot plant tests are relatively simple for those experienced in running them. Reliable and repeatable results can be obtained if simple, fundamental procedures are followed. The accompanying figure shows a typical conventional crushing-grinding circuit with three stages of crushing followed by ball mills or a rod mill-ball mill combination. This flowsheet is typical of many mills that have been built throughout the world. Also shown is a typical semiautogenous (SAG) circuit with feed coming directly from a primary crusher, and the semiautogenous mill product going to the separation process. Many similar semiautogenous circuits are built today because they are simple and flexible, lend themselves readily to high tonnages, and usually cost less to build and operate. With some ores, it is possible to send mine-run ore directly to the semiautogenous mill, eliminating the need for a primary crusher, as at Benquet's Dizon mill in the Philippines. An autogenous or semiautogenous circuit, as the name implies, uses the ore itself to assist in the grinding process. Generally, for this type of circuit, the ore must be competent so large ore chunks can be used to grind smaller ones. If ore alone is used, then the circuit is autogenous; but if balls are added to supplement the charge, then the circuit is termed semiautogenous. Ore competency is not always a criterion, to break up large pieces and wash away cementing material. Autogenous and semiautogenous mills have been very successful on relatively soft uranium ores where the values are frequently found in the material that cements together the sandstone grains. Autogenous Circuits When an ore is hard and competent, it is always possible it might make good "pebbles" that could be used for fine grinding media. Seven methods have been tried using combinations of autoge-

Jan 6, 1982

By Luis Fernandez-Deza, Kaveh Talebi, Seth Pollak, Soto Vardakos

INTRO PAGE

Jun 13, 2021

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada's Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By Steve O. Green

Introduction The US coal mining industry is in the midst of a serious economic crisis. To overcome increased operating costs, operators must take advantage of technological advancements in mining systems and equipment. A relatively new concept available to underground operators is that of extending the cutting depth for continuous mining machines (continuous miners). This system is referred to as deep-cut mining, and may be described as mining cuts to depths greater than the standard 6 m (20 ft). Productivity is increased through the reduction in continuous miner tram time since the machine stays longer in each working face. To successfully implement a deep-cut mining system, three problems must be overcome: Exposing personnel to unsupported roof, methane removal, and controlling respirable dust. Remote control cutting The solution to the unsupported roof problem is remote control. Recent advances in remote operation, especially in radio remote controls, have resulted in what is generally accepted as a safe and reliable tool. By installing remote control on the continuous miner, the operator and helper can now be stationed in a supported area. This allows for improved visibility while cutting coal. With this in mind, the limiting factor now becomes the position of the haulage equipment operator. North River No. 1 uses Joy 21SC shuttle cars and Jeffrey 4110 diesel ramcars. These cars are pulled against the bumper of the continuous miner. Thirteen meters (43 ft) can be cut with the shuttle car and 14 m (45 ft) with the Ramcar before the operator of the car is inby permanent supports. By loading away from the bumper, about 1.5 m (5 ft) can be added, bringing total cut depths up to 14.5 and 15 m (48 and 50 ft), respectively. Methane removal The second problem involves methane removal, and the line curtain-to-face distance. In other words, how to maintain the line curtain to allow sufficient ventilation to remove methane and not expose workers to unsupported roof. The solution, in most cases, is to use a blowing-type ventilation system. Unlike the exhausting system that, to be effective, must be installed within 3 m (10 ft) of the working face, the blowing system can be effective at greater distances from the face. Numerous studies have documented the advantages of a blowing-type ventilation system over an exhausting system. Furthermore, a blowing system can be significantly improved by in¬creasing the velocity of the air delivered without increasing the volume of air. With this greater ventilation efficiency, it is possible to secure ventilation plan approval from the Mine Safety and Health Administration (MSHA) for a greater curtain set-back distance. This eliminates the need for personnel to travel inby permanent supports to extend the face brattice. The distance from the face MSHA will approve, however, is on a mine-by-mine basis. MSHA will evaluate the particular mine's face methane liberation rates, along with the specific proposals regarding face ventilation volumes and velocities. Dust control The third problem is controlling respirable dust. This is actually a result of the solution to the second problem. In the blowing system, while mining on the slab cut, both the continuous miner and haulage units are located on the dusty side of the line curtain (Fig. 1). In the early 1970s, underground operators virtually abandoned this system, in order to comply with the 2 mg/m3 (0.87 gr per cu ft) federal dust standards. The blowing face curtain did offer increased face ventilation efficiency. With that, greater curtain-to-face distances were also

Jan 8, 1987

By G. D. Bates

INTRODUCTION The Magmont mine is a joint venture of Cominco American Inc. (operator) and Dresser Minerals, Inc. The mine-mill operation is located approximately 160 km (100 miles) southwest of St. Louis, MO, in what is commonly referred to as the "Viburnum Trend.” The Magmont mine is designed for a production rate of 3810.2 t/d (4200 stpd) on a 5-day week, three shifts per day basis. Initial production began in 1968. The mine is open stope, room-and-pillar, and essentially horizontal along the trend of the ore body. Briefly, the main geological structure can be described as a brecciated graben bounded by reverse faults. The ore body in cross section is shaped like a bell curve with some lateral extension at the lower part. Presently outlined ore is 609.6 to 762 m (2000 to 2500 ft) in width and 2133.6 m (7000 ft) in length. The ore varies in thickness from 4.87 m (16 ft) on the fringes to an average of 27 m (90 ft) in the high ore areas bounded by the reverse faults. Lead is the primary metal with zinc and copper secondary. MINE DESIGN The basic design of open stope, room-and-pillar mines has been described by several writers and need not be repeated here. (Anon., 1970; Bullock, 1973; Casteel, 1972; Christiansen et a]., 1970; and Lane, 1964) This discussion covers the mining sequence as applied to the particular problems at the Magmont mine, the use of equipment, and deployment of the work force. In the upper portion of the Magmont ore body is a layer locally called the False Davis shale. This layer lies below the true Davis shale, is normally interbedded with dolomite, is of varying thickness, and if mineralized, is included in the top pass of the mining sequence. In thick ore areas this layer will be 2.13 to 2.43 m (7 to 8 ft) in thickness and will occur in the upper portion of the pillars. Due to its incompetency the presence of this False Davis layer is of primary concern in mine planning and operation. Mining areas are divided into three basic groups by ore thickness. First are areas of ore up to 6.09 m (20 ft) in thickness. These areas are below the False Davis shale and are mined single pass with drill jumbo. Second are those areas up 13.71 to 15.24 m (45 to 50 ft) in height. The first 4.87-111 (16-ft) Pass is taken at the top of the ore and the back and pillars secured. Benching the lower portion(s) in 4.57 to 4.87-m (15 to 16-ft) passes is then done with either a drill jumbo drilling horizontally or a crawler drill drilling vertically. Normally these areas are below the Table 1. Productivities per Manshift False Davis shale. These areas may also be mined by back slashing, or overhand benching, where the first 4.87-m (16-ft) pass is taken at the base of the ore and successive 4.87- m (16-ft) passes are taken upward. A minimum amount of back slashing is done at Magmont since it requires repetition of roof control on each pass and roof control is the single largest stoping cost at Magmont. Ore left to provide a working platform oxidizes and is coated by oil spills thus reducing metallurgical recoveries. The third mining area is over 15.24 m (50 ft) in height UP to a maximum of 40.23 m (132 ft) and will encompass the False Davis shale. These areas are mined by first driving +15% inclines to the top of the ore body. The top pass is mined and the back is bolted and roof mats installed as a matter of standard practice to minimize roof problems as mining progresses downward. Once the back and pillars on the top pass are secured, benching begins on successive passes with either the drill jumbo or crawler drill. Pillars on all successive passes below the top pass are secured as necessary. While benching progresses below the top pass, the pass at the base of the ore body is mined leaving a sill of 4.57 to 7.62 m (15 to 25 ft) in thickness to be removed with the crawler drill in a retreating manner. Rooms are mined on a 1.57 rad (90") grid pattern to insure alignment of pillars where multiple passes are taken. Pillars are designed on a 17.98-m (59-ft) spacing with rooms up to 10.66 m (35 ft) in width. Heading widths are wide enough for the mobile equipment to turn without additional allowance for curves. The result is a flexible layout which provides a maximum number of headings available for high extraction rates and grade control. PRODUCTION Incentive Bonus Incentive bonuses play an important part in the mine production at Magmont. Production crews are trained to perform only one of the mining functions of drilling, blasting, mucking. or roof bolting. This specialization, or functionalization, is augmented by development to open all possible stoping areas as early as possible in the life of the mine. This insures that each crew will have enough headings to perform its specialty. The incentive bonuses increase exponentially as output increases. The lucrative incentive bonus coupled with the specialization of the production crews and proper mine development have combined to give the high productivities shown in Table 1. Development crews perform all mining functions in their area. The incentive bonus is paid on a per foot basis, Crews on different shifts working the same heading share equally in the bonus proportional to their contract hours worked.

Jan 1, 1982

By Roy M. Fleming, Christine B. New

The federal standard for limiting exposures to miners from radon daughters was reduced from 12 working level months (WLM) per year to 4 WLM per year in 1971. However, even at that time some researchers were concerned that the new limit would eventually be shown to result in an excess risk for lung cancer mortality in miners. The National Institute for Occupational Safety and Health (NIOSH) is now engaged in a comprehensive review of this topic. As the lead agency in this project, NIOSH has developed a work plan and established a work group to implement this plan. The procedure and specific considerations are outlined in the work plan for developing a comprehensive ionizing radiation standard recommendation for all miners, underground and surface. Such a recommendation will not only include estimates of health risks at various levels of exposure, but also appropriate recommendations for medical monitoring, sampling and analytical methods, sampling strategies, posting, engineering controls, personal protective equipment and recordkeeping. The work group has twenty-four members. Eleven members are NIOSH personnel representing six divisions of the Institute. The remaining thirteen members represent other federal agencies, specifically the Bureau of Mines of the Department of the Interior, the Mine Safety and Health Administration and the Occupational Safety and Health Administration of the Department of Labor, the Office of Radiation Protection of the Environmental Protection Agency, and the Bureau of Radiological Health of the Food and Drug Administration, Department of Health and Human Services. Several factors contributed to the decision to initiate this project. First, the efficacy of the current standard was considered. An initial study group was formed in the spring of 1980 by NIOSH to identify and evaluate articles that contained information on lung cancer mortality risks at and below the present permissible exposure limit. The conclusion drawn from their evaluation was that a two-fold excess risk of lung cancer mortality at and below 120 cumulative working level months (CWLM) of exposure to radon daughters is evident. This composite indication from selected studies was of sufficient magnitude to justify further evaluation. The study group, however, recognized that other studies and information must also be considered in a quantitative risk assessment which would form the basis for recommending an acceptable and feasible permissible exposure limit. A second factor in the decision to pursue further evaluation was the gaps in the current standard that had previously been identified by the Mine Safety and Health Administration (MSHA). These included lack of medical monitoring of underground miners and absence of regulations for surface miners. The seriousness of the health hazard relative to other hazards in mining was also considered. Along with exposures to silica and asbestos fibers, radiation was judged to be one of the major health hazards in mining, These combined factors constituted the justification to develop criteria and recommendations for improved mandatory health standards. The development process begins with a survey and review of the available world-wide information on the topic, including data and information developed by NIOSH. The end product of this review is to be a document that will contain an evaluation of the collected information and support for any recommendations that are made. To develop this document, the present work group has been divided into five task groups with the following emphases: Health Effects, Medical Aspects, Monitoring, Environmental Exposures, and Engineering Controls and Work Practices. The Health Effects task group is to evaluate the evidence from epidemiologic and animal studies of adverse health effects associated with all forms of ionizing radiation encountered in mining and milling operations. The "weight-of-evidence" of the results of all relevant and useful studies will be summarized, with the reasons for emphasizing the cited studies. The critical cells or tissues and the factors that should be considered in estimating the dose to these areas from various types of radiation will be identified. An evaluation will be made of the possible impacts of smoking and exposure to diesel exhaust on the determination of health effects related to radiation. The implications of biologically redundant dose in terms of the time between tumor initiation and death will also be analyzed. The Medical Aspects task group is to review the generally-accepted and the state-of-the-art medical technology for the detection of adverse health effects from ionizing radiation exposure. An evaluation will be made of the accuracy of urine and fecal analysis, wholebody counting, chromosome analysis and nose blows, as well as their usefulness for early detection of adverse health effects. Early detection is of little utility to the affected individual unless subsequent medical care can improve the prognosis. Recommendations for screening tests will be made after carefully considering the accuracy of the diagnostic procedures and the usefulness of early detection. Required recordkeeping and transfer rights will also be addressed. The Monitoring task group is to consider the state-of-the-art technology for the monitoring of occupational radiation exposures. Instrumentation, sampling strategies and analytical procedures will be reviewed for both personal and area sampling. The implications of the associated levels of confidence for non-compliance decisions will be evaluated. The discussion will also include an evaluation of the feasibility of replacing present monitoring systems with recent technology and the impact that a possible lower permissible exposure limit would have on monitoring requirements. The Environmental Exposure task group is to investigate the field procedures and mathematical methodologies which have been used to quantitate exposure levels to the various kinds of ionizing radiation. The magnitude and direction of the possible biases in past exposure assessments will be estimated. Specific attention will be given to the controversies concerning the quantification of the biological dose equivalent and the feasibility of recommending standards for mixed radiation exposures which use "rem" as the unit of measurement. The Engineering Controls and Work Practices task group will analyze the advantages and limitations of

Jan 1, 1981

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984

By Carl Eschman

Productivity from large walking draglines is primarily dependent on operator skills. This machine may be in operation three shifts a day, 364 days a year, and its output is directly related to coal uncovered and mine profitability. Dragline operators must have highly developed manual skills and be knowledgeable in mine planning and working strategies. When using equipment costing more than $25 million, some formal training is usually required before an operator is allowed to assume complete con¬trol; however, dragline operators rarely receive any structured training before operating these giant excavators. A form of apprenticeship is usually followed where an operator candidate progresses from a groundsman to an oiler position. As an oiler, he is permitted to operate the dragline for short periods under supervision. After apprenticeship, the operator is considered sufficiently prepared to operate the largest, most powerful machine at the mine. The apprenticeship training method has obviously provided the surface mining industry with skilled dragline operators; however, conditions are arising that require a realistic and effective training tool that can be accessed by mining companies. New mines -either planned, under construction, or recently opened in the West-do not have access to a pool of experienced operators and oilers as do Midwest mines. As coal mining activities increase in both the West and Midwest, demand for trained dragline operators could be required in a short amount of time. Also, the more productive techniques along with sound basics of strip mining are sometimes lost in the informal "OJT" training method. Modern draglines are the pacemakers of the strip mine, and are simply too expensive to be used as a training device where lost productivity and susceptibility to damage can directly affect mine output. The Dragline Training System is a logical first step in formally educating or retraining operators. The program, started by the US Bureau of Mines and continued by McDonnell Douglas Electronics Co. under contract with the US Department of Energy, was installed and evaluated at DOE's Carbondale Mining Technology Center in Carterville, IL, last year. It is now being operated by Southern Illinois University at Carbondale. System Description The Dragline Training System addresses specific environments and work practices encountered during an actual mining operation at a midwestern US surface mine. This area was chosen because of its high number of strip mines using large walking draglines. Most draglines in the region are Bucyrus-Erie 1370, 1450, and 2570 models, so the dragline trainer was patterned after the company's 1370 machine. Operating and emergency controls are sufficiently standardized for most large walking draglines, and peculiar dynamics and responses from any specific dragline can be programmed into the computer system. The computer simply prompts the user to select the manufacturer and any peculiar response or rate changes needed. A 46-m3 (60-cu yd) bucket is simulated, but for closer simulation, various bucket configurations can be provided. Dragline Trainer The dragline trainer uses the TV-model simulation technique. A scaled model of the dragline is positioned in a model mine. A television camera is positioned at the operator's theoretical eyepoint, and the view captured is projected into a large screen in full scale. The screen is positioned in front of the operator seated in a full-size cab at Bucyrus-Erie controls. By manipulating the controls, the trainee can operate the model dragline and observe its reaction in the television display. In addition to housing dragline controls and consoles, the wooden, oversized cab contains the digital computer, terminal, video recorder and monitor, power switch box, air conditioner, and has enough room for the instructor and five student observers. The 50:1 scaled dragline model contains servo-controlled functions for hoist, drag, swing, delta swing, and longitudinal and lateral position. The delta swing provides bucket lag during swing and a realistic pendulum action when the swing is terminated. The over-responsive second order servo system is designed to provide hoist, drag, and swing rates exceeding present draglines. In all cases, position servos are used for better control and sta¬bility. The normal rate operation of an actual dragline is computed for the specific machine and presented to the servo amplifiers as iterative position commands increasing or decreasing

Jan 6, 1982