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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 Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By Bond. F. C.

Introduction Although autogenous grinding applications date back to the early years of mineral processing when tumbling-type mills were first used, there has been a renewed interest in this type of grinding over the past few decades. A review of the historical aspects of autogenous grinding is provided which follows the development of the various types of mills and processes. Wet and dry grinding operations using either autogenous or semiautogenous mills, including the Aerofall mill, are described. Em¬phasis has been placed on the test work required to evaluate the feasibility of using autogenous or semiautogenous grinding for a spe¬cific application. The factors which must be considered to establish the appropriate design criteria for these types of grinding circuits are discussed in detail. Successful application of these grinding circuits results from a thorough test program, a careful analysis of the test results, and a practical engineering approach to the design of the process flowsheet and milling equipment. Certain pebble milling applications, where the pebbles consist of the ore or rock itself, can be classified as autogenous pebble mill grinding, or better known as ore-pebble mill grinding. In the generic sense, pebble milling includes those grinding operations using a tum¬bling mill in which the grinding media consists of various types of pebbles, either ore or other pebbles, instead of metal balls. Conse¬quently, this subject is covered separately in Chapter 5, Pebble Mills. Operating data from many of the installations located throughout the world are included in this chapter. Our appreciation is extended to those companies who contributed this important information. History of Autogenous Grinding F. C. BOND Autogenous grinding, the grinding of ore by itself rather than by special metallic or nonmetallic grinding bodies distinct from the ore, has an old and honorable place in the art of mineral processing. It antedates the use of the rod mill and is only a few years younger than the development of the continuously fed rotating tumbling mill in which it is accomplished. Its history71, 72 is inextricably bound up with that of gold mining. Every important early development of autogenous grinding took place on ores of gold, except in one case of silver ore. In common with many other features of modern mineral processing, autogenous grinding would have been greatly retarded if the winning of gold had been so small a part of mining in the past as it is today. Gold unlocked the secrets of rock-on-rock grinding. In years past gold ores were crushed in stamp batteries and then flowed over copper plates covered with mercury which amalgamated the gold. The McArthur Forrest cyanide process for the dissolution of gold in a dilute alkaline cyanide solution appeared in 1890. With finer grinding it recovered more gold than could be won by amalgama¬tion alone. The first response was to stamp in cyanide solution and finer mesh screens guarding the stamp discharge. However, the fine screens greatly restricted the tonnage stamped, and they wore out rapidly. Rotating tumbling mills70, 73 were the obvious choice to grind the coarse product from stamps. Many of these were Gates tube mills; more than 1,000 of these were sold between 1907 and 1913. They were made in 6, 51h, and 5 ft diam, with lengths of 20 and 22 ft. The 5 x 22-ft size was the most popular. The term tube mill did not refer to the mill shape, but included all tumbling mills using grinding pebbles. The flint or chert rocks found along the coasts of Denmark and Normandy are probably the most durable and resistant to abrasive and tumbling wear of any on earth. They were much used in the early tube mills. The early linings were all silex or Danish flint blocks, cemented into place with portland cement. Local hard stones or cast iron blocks were also used later. The cast iron liners carried lifters and were not bolted to the shell, but were held in a tight circle against it by steel wedges hammered home between them. The first ribbed metal liners designed to hold grinding pebbles pounded into them by the mill action were developed at El Oro, Mexico, about 1905. These and the Forbes, Komata, and Tonopah types, as well as the Osborne type developed in South Africa within about five years, were very successful in reducing liner wear. They sometimes lasted two years or more, and equalled or exceeded the life of silex lining at a lower cost. Some of the types were patented. The pioneer name recorded in the history of autogenous grinding is that of Kenneth L. Graham. In 1907 at the Geldenhuis Deep Ltd. mine near Johannesburg, he ran a comparative test on two tube mills. One mill used pieces of the common gold ore, or banket, and the other used the customary imported Danish pebbles as grinding media. He published the first account of autogenous grinding.'' In a test lasting 81 days the ore pebbles showed a definite saving over the Danish pebbles. Graham's historic paper also describes the invention of B. Chew of the Crown Deep Mines of the first trammel to be attached to a grinding mill discharge. The first known account of autogenous grind¬ing printed outside of South Africa was the description of the Graham test in Mining and Scientific Press's The use of grate- vs. overflow¬discharge was also described at this time." The news of this successful test appears to have had immediate effect, for by the end of 1908 the number of tube mills on the Rand had increased from 72 to 120.76 Presumably, most of these mills were now using autogenous grinding. The ore was principally a hard conglomerate of quartz peb¬bles cemented by silica, and was eminently suitable for use as a grind¬ing media. It is unfortunate that in South Africa the new autogenous process continued to be called by its old name of pebble milling, since the name does not distinguish between grinding with ore and grinding with extraneous pebbles. For increased clarity the former is called rock pebble milling. The first historical stage in the development was what is now designated as secondary autogenous grinding, or rock pebble milling. It consists of grinding a feed all passing 'A in. or finer with ore pebbles about 2 to 4 in. in diameter. The next stage to develop was that now called intermediate autogenous grinding, where larger amounts of larger pebbles about 5 in. in size were used to grind ore which had been crushed to about rod mill feed size or somewhat finer. It was developed on the Rand and in other places where coarse stamp screens were used to increase mill tonnage. In many cases the early secondary and intermediate autogenous mills have now been replaced by conventional ball mills and rod mill-ball mill circuits, with the object of further increasing grinding capacity. The third stage to develop is designated as wet primary autogenous grinding, in which run-of-mine ore or a primary crusher product is all fed to the autogenous mill together. It developed many years later indepen¬dently in the United States and in South Africa. The fourth and most recent stage is that of dry primary autogenous grinding, which developed in the United States and Canada. All four stages originated in gold mining. The first use of commercial autogenous grinding outside of the Rand apparently took place five years later in 1912 at the Santa Gertrudis gold mine at Pachuca, 55 miles northeast of Mexico City. This was followed closely by an installation at Goldfields, Nevada, in 1914-1915. The first name associated with autogenous grinding

Jan 1, 1985

By K. Terhaar, K. Prisbrey

The objective was to determine operating protocols for a lead and silver column flotation cleaner, being considered for installation at Sunshine Mining & Refining Co.’s Kellogg mill in Northern Idaho. In this column, tetrahedrite and galena were floated, and pyrite and other gangue minerals, including siderite and silica, were depressed. The procedure was to develop a dynamic simulation using a linear transfer function approach. The transfer function related cause and effect variables such as the influence of collector and depressant dosages on recovery and grades. Bench-scale flotation tests followed by pilot column testing provided the data used to develop a linear transfer function. The transfer function was derived using a spreadsheet. This model was applied to simulate how the flotation column could be made to respond to disturbances in feedgrade. The results were that xanthate, zinc sulfate, and sodium cyanide were found to be suitable controls for grade and recovery. Simulations showed the best sequence of manipulations of these reagents, and these sequences were useful for future operator training. The conclusion was that linear transfer functions were able to simulate lead and silver recovery, develop scenarios for operator training, and provide a foundation for future on-line control.

Jan 1, 1999

By Louis I. Briggs

Evaporite rocks, salts of the more soluble ions in seawater, are deposited when seawater becomes concentrated by excess evaporation in marginal basins having restricted circulation with the open sea. The simple conceptual model of an evaporite basin (Figure 1) has a single inlet connecting the basin to the open ocean, the restriction to open circulation being caused by a bar or reef near the mouth of the inlet. As the oceanic water flows from the inlet into the basin it tends to form a thin layer of brine that spreads out over the surface of the basin brine, and as it does so, the salt concentration In the surface layer gradually increases owing to continued evaporation of the water. When the solubility products of the salts in the surface layer are exceeded, the salts precipitate and settle to the bottom of the basin, The order of precipitation follows their order of least solubility. Normally the carbonates separate first, followed by the sulfates and chlorides and the more complex double salts. In nature there is considerable overlap in their ranges of precipitation. The water flow in the basin, also shown schematically in Figure 1, diverges from the inlet essentially perpendicular to the lines of' equal salinity in the surface brine. The distribution of salts deposited in such a model will parallel the isosalinity lines, being arcuate and concave toward the inlet (1)

Jan 1, 1967

By Yngve Jordal, Sverker Hartwig

The IVAR Project is Norway's largest sewer project presently under construction. One part of it is an 8.1 km (26 575 ft) long, 3.5 m (11 ft 6 in.) diam tunnel, bored in one year. The tunneling project has been running smoothly; however, not everything has been easy. A few of its features are highlighted in this paper: The use of 9 0 (12 cu yd) muck cars on a double track back-up in such a comparatively narrow tunnel. Employing only four men in the tunnel, including the loco driver. A quartzite dyke - 96% quartz with 350 MPa (50 000 psi) and more compressive strength - turned out to be 640 m (2 100 ft) thick instead of the predicted 90 m (300 ft). Passing under a gaseous swamp with only same 5 m (16 ft 5 in.) of rock cover on top.

Jan 1, 1991

By Ian Farmer, Nigel Glossop, Philip Garritty

Performance of tunnel boring machines - usually expressed in terms of penetration per revolution - is determined by relations between machine thrust and torque, head diameter and rock strength. It can be shown that for a given machine power and head diameter there is only one rock strength at which the machine performs optimally. In weaker rocks the available torque limits performance and in stronger rocks, the available thrust. Performance data shows that, historically, machines have been designed to cut weaker rocks. Optimization of thrust is required to improve performance in stronger rocks.

Jan 1, 1987

Although out of the ordinary, MINING ENGINEERING is pleased to present a few photographs that typify Western US mining activity during WWII. The photographs, taken by R.G. Zellers at Colorado mining operations between 1942 and 1946, were primarily intended for equipment manufacturers. The pictures capture scenes of hardworking miners during a time when output was essential to fuel a wartime effort. In addition to providing a rare opportunity to step back 40 years and view the equipment and activity common at the time, Zellers' skill with photo composition and his ability to obtain clarity in dimly lit underground operations gives viewers an acute awareness of the dampness and dirt of the mines, and a sense for the miners' limited working space.

Jan 6, 1983

By Alan Weakly

The Powder River Basin (PRB) of northeastern Wyoming and south-central Montana is the largest strippable reserve of compliant coal (0.6 or less lb of sulfur per million Btu) in the United States. In mining parlance, it is the Mother Lode! Wyoming’s portion of the Basin provides one fifth of the coal produced in the United States annually, and the state has been the nation’s top coal producer for the past six years. Based on 1993 production statistics and estimated strippable compliance coal reserves, Wyoming may remain the nation’s top producer for the next 76 plus years. Past articles on the PRB have focused mainly on production. After living and working in the Basin for the past five years, I have developed an appreciation for the area and its major industry that transcends size and production data. In the fall of 1993, MINING ENGINEERING’s publisher and I decided to present a series of articles that would cover many other aspects of the Basin. Ten topics were outlined, and specific individuals, selected for their knowledge and experience, were asked to contribute articles about their specialties. Although Montana has four large mines producing coal from the PRB, this series of articles will concentrate on the Wyoming portion of the Basin.

Jan 1, 1994

By Richard Addison

Introduction A staccato beat rings out across the Australian bush. The stamps are at work, slowly pulverizing gold-bearing ore for a patient, grizzled prospector. This is the State Battery near Marvel Loch, a tiny 22.6-t/d (25-stpd) capacity gold plant operated by the Western Australia government. It is one of many such plants located in remote corners of Western Australia and the Northern Territory. They are miniature custom mills operated for the benefit of prospectors and small mine operators. These state batteries are so called because they are operated by state governments and because the principal item of machinery in the plants are stamp batteries. They are fascinating anachronisms - throwbacks to times past, a century ago, to the hard and heady existence of gold rush days. Marvel Loch plant Typical of these operations, and perhaps a trifle more sophisticated than most, is the plant near Marvel Loch, a small center for mining and farming about 400 km (250 miles) due east of Perth. Marvel Loch. The name evokes visions of a pristine lake set in a cool, green, misty glen of the Scottish highlands. Actually, it is probably a hot, dry salt pan somewhere nearby. Just where the "loch" is difficult to discern. There is no sign of it on the map.

Jan 7, 1986

By Stephen Davies

Mining at Hudson Bay Mining & Smelting's Snow Lake Division is now in the mature stage of the mines' lives. These base metal copper-zinc mines have been operating for more than 30 years. Stopes had small pillars called "crown pillars" left between them. These pillars range up to 100 m (330 ft) long by 40 m (130 ft) wide and 15 m (50 ft) thick. Hudson Bay have been successful in removing these pillars by using cable slings, electric scoops, longhole mining and diesel load-haul-dump equipment. The different shapes and sizes of these pillars has required that many innovative techniques be tried in order to find acceptable economic mining methods. An overview of the different methods will be outlined. Background Hudson Bay Mining & Smelting's Snow Lake Division is located in Manitoba some 800 km (500 miles) north of Winnipeg and 200 km (125 miles) by road from the Hudson Bay smelter complex located in Flin Flon, Manitoba (Fig. I). There are currently three producing underground mines and one open pit operating at the Snow Lake Division, producing copper and zinc concentrates. Pillar recovery is intensive and vital to continued production of these mature mines. The crown pillars are producing more than 50% of the current mine production. This figure will increase in the future. The two mines at which crown pillar recovery is taking place and that will be discussed are the 400 t/d (440 stpd) Rod Mine and the 1000 t/d (1100 stpd) Stall Lake Mine, both five-day-a-week producers. Crown pillar mining has been applied for some time at these mines. However, it was only recently that the crown pillars became vital to production. This has necessitated a number of innovative techniques to recover them.

Jan 1, 1992

By R. M. Schuster, F. M. Benavides

Samuel G. Bonasso The article by Benavides and Schuster provides an in-depth study of three transportation alternatives for a four-mine complex in western Kentucky. The examination of aerial tramways was not undertaken as an alternative solution. We feel this is an oversight since industrial ropeways are competitive with both belt conveyance and overland truck hauling. In fact, there are aerial tramway systems, operational today, which are 100 years old and operating with some original components. Maybe unfortunately, there are no fixed rules or guidelines whereby the most cost-effective means of transporting and handling material for a given situation may be selected. Each situation must be evaluated so the variables that influence the selection of the desired alternative can be clearly defined. However, there are in-place industrial ropeway systems that have proven economically feasible over a range of 1.6-50 km (1-31 miles). The more difficult the terrain, the greater the cost efficiency.

Jan 1, 1983

By Steve Kral

After years of cutbacks, closures and layoffs, the mining industry is beginning to expand again. Metals prices are up and so is investor interest and confidence in the industry. Gold prices began to slowly move up in January2003 and finally cracked the $12.86/g ($400/oz) mark in early December. Silver, too, is up, as is copper. So the mining industry is feeling good about itself again and many companies are expanding existing mines and developing new ones. One side benefit to mining’s improved outlook is that companies are beginning to send more of their employees to industry conferences. Such was the case at the109th Northwest Mining Association meeting in Spokane, WA. The meeting was held in early December and attracted 1,100 attendees and 84 exhibiting companies. That is up from 2002’s attendance of about 900 people and 65 exhibitors. The theme was “A Partnership into the Future.”

Jan 1, 2004

By Steve Klein

The Transbay Downtown Extension (DTX) in San Francisco involves the construction of a large-span tunnel with a width ranging from 43 to over 60 feet. The tunnel will be excavated using the New Austrian Tunneling Method (NATM), also known as the Sequential Excavation Method (SEM), in highly fractured and sheared rock. This paper discusses the FLAC (Fast Lagrangian Analysis of Continua) analyses that were completed for conceptual design of the tunnel support system and to develop surface settlement estimates for the evaluation of measures to protect existing buildings located above the tunnel from damage.

Jan 1, 2008

Coal has been mined in the central Appalachian coalfields of southern West Virginia, western Virginia, and eastern Kentucky for more than a century. The dwindling reserve base consists in large part of coal that would have been considered unmineable by earlier generations. Nearly every current operation is working on a property where coal has been extracted in the past, from seams either above, below, or both. NIOSH is conducting research aimed at helping mine planners prevent hazardous conditions due to multiple seam interactions. To-date, more than 300 case histories have been collected from underground mines, mainly from central Appalachia. This paper focuses on several of the more challenging situations that have been encountered, including: • Room-and-pillar development 20 ft (6 m) beneath full extraction workings at a depth of 1,000 ft (300m) of cover (Virginia) • Pillar recovery 45 ft above full extraction workings at 900 ft (270 m) of cover (Virginia) • Near-simultaneous room-and-pillar mining with pillar recovery with 40 ft (12 m) of interburden and 1,500-2,000 ft (450-600 m) of cover (Kentucky) • Longwall mining directly beneath main entries in overlying seams (West Virginia) • Some of these operations have been highly successful in overcoming the challenges, others less so. The lessons learned from their experience will help ensure that these and similar difficult reserves can be mined safely.

Jan 1, 2006

By Steven B. Van Kouteren

Value added is a loosely used qualifier in the industrial minerals industry. What exactly does it mean? What does it take to make a value-added product? More importantly, how does a company maintain the value added to its products? Most people agree that value-added means the value of a mineral minus the cost to produce it. This is misleading, though, because producing value-added products does not necessarily guarantee the profitability of the products. In some cases, firms that make value-added products are not very profitable. It is important not only to understand how to make a value-added product but also how competitive forces in the market can erode the value added. Therefore, when looking at value-added opportunities, producers must understand how to obtain the value added and how to maintain that value. Four major factors limit a company's ability to add value to its products. This paper examines these factors. It uses the ground calcium carbonate and mica industries as examples.

Jan 1, 1990

By Robert L. Hautala

In 1978, Congress established a series of mineral institutes for a limited time to assist financially in mineral industry education and research. Later, education and research were separated, establishing generic institutes for specialized research that transferred the former research budget from the base mineral institutes program. One of the generic institutes is the Mine Systems Design and Ground Control (MSDGC). It provides for research funding specifically for mining. Other parallel generic institutes were established in pyrometallurgy, comminution, waste management, etc. MSDGC was established at Virginia Polytechnic Institute and State University(VPI). The founding group included the University of Alabama, University of Kentucky, University of Nevada Reno, University of Idaho and University of Alaska reporting through VPI to the US Bureau of Mines (USBM) and Congress. Other universities also participated on an ad hoc basis.

Jan 1, 1996

By Tim O’Neil

The US Bureau of Mines and the Santa Cruz Joint Venture (SCJV) are working on a production-scale, in situ copper recovery project in Arizona. The research is being conducted at the Santa Cruz site, located 11.2 km (7 miles) west of Casa Grande, AZ (Fig. 1). Santa Cruz is a 1.36-Gt (1.5 billion st) porphyry deposit averaging 0.6% Cu, of which 65% is oxidized, concealed beneath post-mineral alluvium at depths of 365-914 m (1500-3000 ft). The Bureau and the SCJV will evaluate in situ mining as a method for extracting copper from this previously unmined deposit. Research objectives of the Santa Cruz project are: - to demonstrate that an in situ research test can be operated in an environmentally acceptable manner ; and * to operate an in situ well pattern long enough to develop information on copper recovery and economic feasibility data. In situ mining of copper from undisturbed deposits has not been attempted commercially. There are several potential advantages with in situ mining: less exposure of miners to hazards, lower pital costs for the mine and plant and lower operating costs. In addition, there are no pits, dumps or tailings, which means less environmental impact.

Jan 1, 1991

By Kal V. S. Sastry

Comminution is an important unit operation of mineral processing in terms of large tonnage of materials processed, huge energy consumption, and high capital and operating costs. Substantial research efforts have been directed towards the development of process-engineering tools including mathematical models, process scale-up procedures and computer simulation packages (Austin et al., 1984; Rajamani and Herbst, 1990; Kawatra, 1992). The simulation systems are found to provide significant cost savings by enabling efficient circuit design and development of control systems. Most of these simulators are built from scratch, do not make use of software modules from other sources and are designed with minimal usage of formal software methodologies.

Jan 1, 1995

By V. Svalbonas

The process of grinding mill quality assurance brings with it associated truths and myths. When the mill quality assurance process took its first great steps forward in the late 1970s and early 1980s, they were in response to the conditions existing, in our industry, at those times. In the late 1990s, we are faced with different conditions, and problems, and need fresh solutions. Yet the system developed for the early 1980s has produced no new approaches for the late 1990s, becoming instead more hardened and entrenched. This lack of development and actual resistance to research is now costing the industry significant amounts of money in late start-ups and useless scrap loss.

Jan 1, 1999