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By J. L. Sepulveda, R. F. Yap, R. Jauregui

INTRODUCTION The Third Theory of Comminution, oftentimes called the Bond Theory, was first published by Fred C. Bond in 1952. Since then, it has been widely used in the milling industry to size crushing and grinding equipment as well as to predict and evaluate their performance. Although two other comminution theories were advanced almost a century earlier, one by Rittinger, 1867, and the other by Kick, 1885, their use was limited because they did not agree with actual commercial comminution systems (Bond, 1961). According to Bond's Third Theory, the work input is proportional to the new crack tip length produced in particle breakage, and equals the work represented by the product minus that represented by the feed. For particles of similar shape, the crack tip length is equivalent to the square root of one-half the surface area, and the new crack length is proportional to [(1/vT - vF)]; where P and F are particles diameter of product and feed, respectively. For practical calculations, the particle diameter can be expressed in terms of the 80% passing sizes, P and F, respectively. Therefore, the work input is given by, [10 _ 10', W = Wi [7P_7F(1)] Where W = the energy input to the mill, (kwh/t) Wi = Work Index, (kwh/t) P = square sieve opening which 80% of the product passes, (microns) F = square sieve opening which 80% of the feed passes, (microns) Equation (1) is the mathematical statement of the Third Theory, or the Bond equation. Numerically, the work index is the energy required in kwh/t to reduce a given material from theoretically infinite feed size to 80% passing 100 microns. The work index of a particular material is determined in the laboratory by conducting the standard Bond grindability test. Bond developed separate tests for the rod mill and the ball mill. The simplified test method described in this paper, which is used at Anaconda Research Center to approximate the work index, is an alternative procedure to the ball mill

Jan 1, 1982

The Illawarra Branch of the Australasian Institute of Mining and Metallurgy over recent years has established an enviable reputation for organising and conducting conferences, colloquia and symposia. And this Symposium must surely enhance that reputation. The subject of the Symposium, Seam Gas Drainage has become a matter of great import- ance to the Australian Coal Industry in a comparatively short time. Our short experience being due to our operations, until recent times anyway, not being unduly affected by gas. The Australian Industry, by virtue of its mines getting deeper and higher rates of extraction being required is now encountering the difficulties with gas which have for so long been part of the coal mining scene in many other countries. Because these gas emissions affect our working conditions by way of creating hazards and affect our production by way of bringing about cessation of operations from time to time this Symposium must be regarded as being very timely out here. I must congratulate the Branch Committee on the general set,yp :o.f the programme of papers. The classification of papers into four groups and the sequence in which the papers were presented provided an excellent flow through from introduction to finality. It is not my intention to further discuss individual papers - already they have been well discussed in this forum and as I understand it outside in the early hours of the morning,

Jan 1, 1982

By Daniel T. Rogers

The Bureau of Mines, in its alumina miniplant project to investigate alumina recovery from domestic, non bauxitic ores, has conducted research on the use of a hydrochloric acid leaching, gas sparging crystallization technology. An important element of this research is the efficient separation of undissolved, siliceous residue from AICI3 -bearing leach liquors by continuous, horizontal, countercurrent vacuum belt filtration. In an effort to calculate material balances quickly and to predict material balances based on different belt filtration configurations, the perfect-mixing-cells-in-series model (PMCS) for calculating material balances around belt filters was used. Because of the porous nature of the solids, the model produced erroneous material balances. Therefore a reliable model, the shrinking voids model, was developed postulating the presence of an unwashable voids liquor volume that decreases with decreasing liquor AICI3 concentration. This volume decrease postulation is equivalent to assuming that dilute liquors flow more freely, causing more voids liquor volume to become washable. Least-squares based computer programs are provided, which are useful not only in producing material balances from plant data but also in predicting balances for untested configurations using the same feed materials.

Jan 1, 1982

By Yoginder P. Chugh, M. Silverman

Premining investigations involve the determination of roof, floor, and coal seam characteristics to delineate areas of troublesome ground control prior to development of a mine. Data obtained can be useful in mine design, equipment selection, estimation of face productivity, development of ground control plans and assessing ground control costs, and in conducting feasibility studies and financial analyses. This paper discusses 1) the causative factors associated with ground instability, 2) how useful data can be obtained for these factors during premining investigations, and 3) analysis techniques to delineate probable areas of ground instability.

Jan 1, 1982

By L. Kraszewski

"IntroductionPortland cement is produced by burning an accurately proportioned mixture (raw feed) of finely ground limestone, silica, alumina, and iron oxide. The kiln discharge (clinker) is a chemically complex mixture of calcium silicates, aluminates and ferrites which when ground and blended with from 2 to 5 percent of calcium sulphate is sold as Portland cement.ManufactureThe most common raw materials are limestone and clay or shale. The raw mix is finely ground and heated to about (2600°F) 1430°C to decompose the calcium carbonate and combine the calcium oxide with silica, alumina and iron to form a partly fused clinker. The clinker is then ground, usually with a small proportion of calcium sulphate (gypsum), which regulates set. Cements with special properties are produced by varying the oxide composition and also by adding small amounts of chemical agents at the clinker-grinding stage or later.Clinker is manufactured by either wet or dry process methods. In the wet process, raw materials are prepared, blended and fed to the kiln as a water slurry (about 30-35% water) . In the dry process, raw materials are prepared, blended and fed to the kiln as a dry powder (about 0.5% water). The processing procedure is as follows:"

Jan 1, 1982

By Edmond Chin

Mineral production was of minor importance to the country's economy, and Bangladesh was not a ranking world producer of any mineral commodity in 1980. Natural gas was the most important mineral exploited, and reserves were sufficient to allow a significant increase in its use. A greater proportion of the domestic heat and electrical energy needs could be filled by increased production and distribution of natural gas. Also, the export of natural gas was feasible either by pipeline to India or in the form of liquefied natural gas (LNG). Some nonmetallic minerals could be developed, as well as coal and peat, given sufficient capital investment. Bangladesh's strongly agricultural economy made fair progress in 1979.80.1 The major harvest of November-December 1979 turned out to be better than had been anticipated after an early drought left poor planting conditions. Secondary crops (harvested in February-March 1980 and June- July 1980) were excellent. There were prospects for a record harvest in fiscal year

Jan 1, 1982

By C. C. White

Much benefit can be achieved by a mining company and a manufacturer if both parties can communicate effectively with each other and assist in product development; mutually beneficial ideas and products will develop. This communication can best be achieved if both parties have the opportunity of having on their respective staffs, specialists whose function is to improve the use of products used in mine development. Although a great wealth of roof control knowledge has long existed among mine operators, until recently there was not an abundance of joint development of roof control material by manufacturers and mine operators. This was perhaps due, in past lean years, to the preoccupation of mine management with other overwhelming problems that arise in mining. Manage¬ment had little time for the lengthy process of product development. Until recently, only a few of the largest mining companies had fulltime personnel devoted to roof control. A bolt manufacturer who is invited to work with an operator's roof control people in developing new material is lucky indeed in that he learns what is wanted and he acquires the benefit of the operator's knowledge and experience. He has only to come up with a satisfactory product at an acceptable price - which is not always easy: The development of new products is usually costly and time consuming, and many new products fall by the wayside. In the case of point anchors, however, this product is here to stay. The advantage of point anchors over mechanical anchors is now a proven fact. This advantage is due to less bleed off and higher ultimate load capacity of the resin point anchors. Point anchored bolts have proven to be superior to fully grouted bolts in many areas. We at Birmingham Bolt Company are fortunate in that we have been involved in joint efforts to develop point anchors with several coal companies from whom excellent advice and direction was received. It is appropriate to mention before continuing that the use of point anchors began in the 1960's in Europe and became widespread only recently in the United States. They fall in two categories, point anchors and combination point anchors. These categories have been defined by the Bureau of Mines as follows: (1) point anchors having 24" of grouted length and (2) combination anchors having 24" or more of grouted length. In the present nomenclature, all are anchored with resin or some sort of cementing compound between the rebar or tensioned element and the wall of the hole. The latest development includes an expansion shell as well as rebar, the purpose of the expansion shell being to facilitate installation without the delay necessary in use of types that require a waiting periond for setting of the resin before tightening. (See figures #1 and #2.

Jan 1, 1982

By Emilio Caballero

El presente texto señala la eficiencia del sistema mecanizado aplicado a las minas subterráneas en Canadá donde se obtuvo rendimientos y costos de producción que permitieron hacer un mejor uso de sus reservas. En ese sentido, se analiza la mecanización y se compara su eficiencia con la Compañía Minera Alianza. Para ello, se revisa el diseño de la mina, selección de equipo, plan de entrenamiento, mantenimiento del equipo y servicio, stock de repuestos, trabajadores en determinada área, supervisión a los equipos y el aspecto laboral.

Jan 1, 1982

By T. G. Carnahan

In-place leaching is being actively utilized to extract uranium, copper, and evaporites from low-grade and difficult-to-mine ores. The purpose of this report is to discuss in-place leaching technology, its advantages and disadvantages, and to give an overview of past, current, and envisioned in-place leaching operations in the U.S.A. Application of in-place leaching to uranium ores has been particularly successful because the conditions for leaching are nearly ideal. The ores are generally in highly permeable sandstone aquifers bounded above and below by impermeable shales. The uranium ore minerals are easily leached with oxidizing acid or base solution and the pregnant solutions are amenable to ion exchange concentration and purification prior to conventional solution neutralization for recovery of yellow cake. Copper is routinely recovered from sub-grade rock by dump leaching at most major copper mines. Acidified ferric sulfate is either sprayed on the surface of the dumps or is ponded on top of the dumps. The leaching solution migrates down through the rock and dissolves copper. The solution is collected from the base of the dumps and copper is recovered by solvent extraction-electrowinning or by cementation on detinned steel cans. True in-place copper leaching is being practiced in the caved stopes of an exhausted underground copper mine. Other innovative in-place copper leaching techniques have been tested commercially and laboratory- and pilot-scale experiments have been conducted to investigate proposed in-place leaching techniques for extracting copper from difficult-to-treat chalcopyrite ores. Solution mining of evaporites is being conducted commercially for the extraction of sodium chloride and potassium chloride. A large-scale experimental project for solution mining of trona has been undertaken.

Jan 1, 1982

By William B. Beerbower

When full length resin grouted bolts were introduced into U. S. coal mines some 11 years ago, it was assumed that safety would be the primary benefit. There is ample testimony to substantiate the safety factor as an important reason for using resin roof bolts. More recently, it has been determined that many coal operators not only concur with the safety consideration but also benefit economically from the safer working conditions resulting from a reduction in roof falls. IMPROVED SAFETY Some typical roof fall reduction experiences with resin bolting follow: ? Western Kentucky Mine - Resin bolting reduced roof falls by 80%. ? Southern West Virginia Mine - Experienced 141 roof falls between 1971 and 1972. In 1973 with resin bolting, roof falls were reduced 85%. ? Western Pennsylvania Mine - In 1973 had 59 roof falls with mechanical bolts. During the same period of 1974 with resin, only 19 falls were experienced - a reduction of 66%. The overall experience with resin bolted roof support has been an average 50% reduction in roof falls. Long term, this can translate into improved worker morale. As a matter of fact, a general mine foreman at the same Southern West Virginia mine mentioned above commented that improved miner morale was noticeable following the introduction of resin bolting. HOW GREATER SAFETY CAN BE COMBINED WITH INCREASED PRODUCTIVITY IN COAL MINING OPERATIONS Experience has also shown that a reduction of roof falls with resin bolting has resulted in increased production for many coal operators. Some typical examples with mines using resin bolting follow: ? Western Kentucky Mine - Resin bolting contributed to increasing production from 7000 to 11,000 tons per day, due to a 80% reduction in roof falls. This was accomplished in part by a reduction of 50% in timbering requirements and the reassignment of three roof fall cleanup crews to coal production. ? Pennsylvania Mine - Following resin bolting introduction the section showed a 50% increase in production. ? Central West Virginia Mine - Continuous miner production averaged 6 cuts per unit shift prior to resin bolting. Reduction of cleanup time, resulting from improved roof support with resin bolting, increased production to an average of 9 cuts per unit shift. PROPERTIES OF RESIN BOLTING SYSTEMS Several of the important properties of resin bolting systems which contribute to its improved performance are listed below: ? Full length anchored resin bolts and many resin anchored, point anchored bolts provide effective resistance to changing rock strata stresses throughout their length, thereby improving stabilization of rock strata over long periods of time. ? Individual layers of rock are anchored into a single high strength beam providing excellent resistance to vertical and lateral strata movement, the most common cause of roof failure. ? Moisture cannot corrode the fully grouted bolt. ? Has twice the pull strength of mechanical bolts (See Figure 1)

Jan 1, 1982

By R. W. Chadwick

Coal was discovered in South Africa in the mid 1830's. The earliest exploitation of modern times was probably in Natal where, in a letter dated 20 August 1842, a Captain Smith mentions that a wagonload of coal brought to Pietermaritzburg, was sold. Until the discovery of diamonds in 1867 and the Witwatersrand gold in 1886, coal mining satisfied a very small domestic demand. By 1889, at least four collieries were operating in the Middelburg-Witbank district, some 120 kilometres East of Johannesburg, and mining had star- ted at Vereeniging, about 50 kilometres to the South of Johannesburg. A map of South Africa is shown in Figure 1 and an enlargement of the main coal mining areas in Figure 2.

Jan 1, 1982

By Neil J. Dahmen, Howard J. Handewith

INTRODUCTION Tunneling machines were introduced into civil con¬struction in the 1800s, but their practical applications were not realized fully until the 1950s. As of 1977, tunneling machines were being used on most tunneling projects of substantial size. As in civil construction, the mining applications of tunneling machines have included long drives such as tunneling for long ore-body develop¬ment drifts, long haulage tunnels, longwall access drifts, and long crosscut developments. Tunneling machines ranging from 1.75-m (5.7-ft) to more than I1.0-m (36-ft) diam are available from several manufacturers; machines for mining applications generally have ranged between 1.75 and 6.0-m (5.7¬and 20.0-ft) diam. Tunneling machines usually operate at or near horizontal grades and are supported by a rail-haulage backup system. However, declines and inclines have been made using alternative backup equip¬ment. Either circular or shaped tunnels may be bored, depending upon the particular equipment used. Using tunneling machines to bore openings through rock offers several advantages over the use of the drilling and blasting technique. These advantages include a rapid advance, continuous operation, uniformity of the opening size, smooth walls, and safe working conditions. Compared to drilling and blasting, boring reduces the disturbance of adjacent rock, thereby reducing the ground-support requirements and the exposure to water inflows. The smooth walls provide natural arching and present minimum resistance to the flow of air. A tunneling machine should be considered a spe¬cialized mining tool that is well-suited to certain mining applications. Properly applied and sustained, a tunnel¬ing machine contributes to realizing substantial econo¬mies in the development of underground mines. All costs herein are expressed in 1977 US dollars, unless otherwise noted. The mention of trade names and manufacturers is not intended to imply an endorse¬ment of any specific equipment. TYPES OF TUNNELING MACHINES As listed in Table 1, five principal types of tunneling machines were being manufactured in 1977. Each type of machine offers a special solution to a specific mining problem. Although each of these types of machines is described herein, most of this chapter is devoted to the full-face rotary rock-tunneling machine that has demon¬strated the greatest production in both civil construction and mining applications. Full-Face Rotary Rock-Tunneling Machines As shown in Fig. 1, a full-face rotary rock-tunneling machine consists of a rotating head that is fitted with rock-cutting tools and is forced into the tunnel face. This head is supported by a bearing on a structural support member that, in turn, is held in place by a hydraulically positioned wall-gripping mechanism. Both torque and thrust are reacted from the cutterhead, through the structural support and gripper mechanism, and to the tunnel wall. The rock cuttings fall to the invert at the tunnel face, and they are removed by means of cutterhead buckets or scoops that transfer the cuttings to a conveyor belt. After advancing the cutter¬head through a preset boring stroke, the tunneling machine is advanced by hydraulically pulling the gripper mechanism in from the tunnel walls, stroking forward, and resetting the gripper to the new forward position on the wall; the machine then is set for the next advance stroke. Full-face machines cut only circular openings. Ma¬chines with diameters as small as 1.8 m (6 ft) and as large as 6.1 m (20 ft) have been used in mining applica¬tions, excavating rock with unconfined compressive strengths in excess of 276 000 kPa (40,000 psi). Although special machines have been designed to turn on a radius of 30 m (100 ft) in both horizontal and vertical directions, the conveyors and backup sup¬port equipment generally limit tunneling machines to turns with a radius of more than 60 m (200 ft). The full-face rotary rock-tunneling machines have provided the fastest and most reliable production of any mining method or machine. These machines can be equipped to place any of a variety of ground-support systems, and they can be provided as "open" machines, as shown in Fig. 1, or as "shielded" machines, as shown in Fig. 2. The open machines have the advantage of allowing the ground to be supported as near to the tunnel face as possible. The shielded machines have been designed principally to allow placing precast con¬crete segments. The shielded machines completely pro¬tect the equipment and personnel at the heading, but they are economically limited to permanent long-life haulageways where full lining is required. Partial-Face Rotary Rock-Tunneling Machines As shown in Fig. 3, the partial-face rotary tunneling machine is equipped with a cutterhead that rotates slowly at about 0.21 rad/s (2.0 rpm). The single-disk cutterhead produces the tunnel section shown in Fig. 4 by means of an oscillating motion. These machines work on the undercutting principle, and they are avail¬able in spring-line dimensional increments of 0.1 m (0.32 ft) from 1.5 to 2.1 m (5.0 to 7.0 ft); the vertical dimensions are 2.4 to 3.2 m (8.0 to 10.5 ft), respec¬tively. Depending upon the tunnel width, the cutterhead disk may be fitted with from 8 to 16 drag-bit cutters. Penetration is achieved by means of hydraulic rams that force the tungsten-carbide bits into the rock mass. The operation is cyclic, with cutting occurring only on the upstroke. On the downstroke, the rotating cutter¬head draws the fallen muck from the invert into the panzer-type conveyor for rearward transfer. This feature allows time to either reposition ore cars or to exit the tunnel with muck trains while the machine is cutting, as shown in Fig. 5. Boom-Type Rotary Excavators The boom-type rotary-excavator tunneling machines commonly are known as boom-type continuous miners

Jan 1, 1982

By P. R. Bremner

In keeping with its mission to maximize mineral and metal recovery from domestic sources, the Bureau of Mines investigated a method to extract silver from tailings not amenable to cyanidation. Results of a laboratory study on the recovery of silver from cyanide tailings containing 1.7 ounces of silver per ton by leaching with an acidic ferrous chloride-sodium chloride solution are presented. The effects of reagent concentration and reaction time on the leaching of silver were investigated. A leaching solution containing 5 wt-pct NaCI, 2.5 wt-pct FeCI2, and 1.2 wt-pct HCI (for acidity) extracted 82 pct of the silver from the tailings during a 24-hour leach. The amount of HCI consumed during the leach was 46 pounds of HCI per ton of feed. Only 23 pct of the silver was extracted by conventional cyanidation. Ninety-nine percent of the silver was recovered from the pregnant solution by cementation onto iron powder.

Jan 1, 1982

By R. D. Studer

"The Snow Lake massive sulphide deposits are hosted by Aphebian volcanic rocks of the Flin Flon - Snow Lake greenstone belt, which in this area were deformed and metamorphosed to the amphibolite facies during the Hudsonian orogeny. All the known deposits occur in felsic rocks a short distance below the contact with an overlying mafic rock unit.The Stall Lake copper deposit occurs close to the top of an explosive felsic volcanic sequence. Rocks stratigraphically underlying the deposit have been hydrothermally altered by the mineralizing fluids. The sulphides were probably deposited from submarine volcanic vent(s) on the flanks of a felsic dome. The ore has been tightly folded, flattened and elongated parallel to the northerly trending lineation of the surrounding rocks during the several periods of deformation. Strike lengths of the lenses range from 60 to 300 feet and plunge-to-strike ratios are as much as 15:1.Ore lenses 1 to 4 typically consist of massive coarse-grained pyrrhotite, chalcopyrite and pyrite. A stratigraphically higher lens differs in that it contains fine-grained pyrite and chalcopyrite, with an increase in sphalerite toward its hanging wall. The No.4 orebody has a distinct footwall alteration zone characterized by brecciation and abundant chlorite-staurolite-garnet schist, with minor kyanite, anthophyllite, cordieriteand anhydrite, as well as sulphide minerals. Chemically, the alteration zone is distinguished from the unaltered felsic rocks by Na20 depletion and Mg-Fe enrichment.IntroductionThe Stall Lake copper deposit is located aproximately 4 miles southeast of Snow Lake in northern Manitoba (Fig. 1). The Snow Lake district has been a producer of gold (Rex, Bingo, Laguana and Nor-Acme Mines) and, since the fifties, of copper and zinc from deposits of Hudson Bay Mining & Smelting Co., Ltd . The district is located near the eastern limit of the Flin Flon - Snow Lake greenstone belt, which contains Amisk metavolcanic rocks of Aphebian age."

Jan 1, 1982

By Beth Robinson

This first of two articles on the Second Mexican Mining Rail Excursion describes some of the more casual experiences participants enjoyed on last fall’s rail tour of major Mexican mining operations. In this article, Beth Robinson, a graduate student at Colorado School of Mines, describes trip organizers, participants, and some of the unique events occurring during the two-week excursion. Next month, part two of the series will highlight various Mexican mines and plants.

Jan 1, 1982

By H. H. Richards, Ray L. Tobie, L. A. Thomas

GENERAL DESCRIPTION Since the beginning of operations, with the exception of a small tonnage mined by slushing, ore extraction has been by full gravity caving. Formerly, a checkerboard sequence of block undercutting was followed with the even-numbered blocks in one panel and the odd¬numbered blocks in the adjacent panel being mined. As these blocks were depleted, the intermediate or pillar blocks were mined (Fig. 1). Following this checker¬board, the mining sequence went through a number of changes, finally evolving into diagonal retreat panel cav¬ing by blocks (Fig. 2). The numbers in Fig. 2 indicate the sequence in which blocks were undercut. Gaps in the numbering sequence indicate undercutting on the level outside the illustrated area. Geology The ore body is a low-grade deposit of chalcopyrite mineralization disseminated throughout structurally weak, highly fractured, strongly altered granitic host rocks. It takes the shape of a gently dipping elliptical cylinder consisting of an ore shell of variable thickness surrounding an interior waste core. Major and minor axes of the mineralized cylinder are 1524 m (5000 ft) and 762 m (2500 ft), respectively, and length approximates 2438 m (8000 ft). Ore is sufficiently fractured to break readily into medium-coarse size. The igneous rock complex containing the ore body is covered by a wedge-shaped blanket of Tertiary con¬glomerate which was brought into place by faulting along the major regional structure of the San Manuel fault. Thickness of the conglomerate cap varies from only 9 m (30 ft) at the east end of the ore body to more than 610 m (2000 ft) at the west. Structurally, the con¬glomerate is much more competent than the igneous host rocks and, when caving, it breaks into massive chunks. Conglomerate boulders seen in drawpoints underground are very coarse. The total rock column over the initial mining area of the 1415 grizzly level was 354 m (1160 ft) of which 122 m (400 ft) was ore, 79 m (260 ft) was leached igneous capping, and 152 m (500 ft) was conglomerate above the San Manuel fault. Diamond Drilling: From 4572 to 7620 m (15,000 to 25,000 ft) are drilled annually from underground workings to delineate the ore body. MINE DEVELOPMENT Haulage Level In the south or main ore body (see Figs. 3-6), with the exception of the draw and transfer raises, all the extraction openings are concreted (Seaney and Tobie, 1965). The haulage panel drifts, which are 18 m (60 ft) below the grizzly drifts, are first driven with pre¬concrete ground support. The drift, which has an arched section, then is concreted using mobile collapsible steel tunnel forms. The haulage drifts leading from the pan¬els to the hoisting shafts are not concreted. After the panel drifts have been concreted, the raise stations from which transfer raises will be driven are constructed and the raise-station ore-drawing chute is installed. The chute is prefabricated of A-36 steel with undercut guillo¬tine gates made of abrasion-resistant 2.5-cm (1-in.) steel plate powered by 20-cm (8-in.) air cylinder installed on each side of the raise station. Transfer Raises The transfer raises are lined with 15 x 20-cm (6 x 8-in.) cribbing and are 1.22 m (4 ft) in the clear. Each cribbing is protected from wear by a high carbon steel angle which is nailed onto the cribbing. The transfer raises are driven from each side of the raise station on an angle of 1.1 rad (63°). Each raise con¬sists of a main and a backover branch. The transfer¬raise driving crew consists of two men working one shift only. Grizzly Drifts After the transfer raise reaches the grizzly level, the grizzly drift can be driven. The grizzly drifts are spaced at 10.6-m (35-ft) centers and are driven parallel to the long axis of the ore body (see Fig. 2). This drift is driven by a two-man crew working on one or more drifts at a time using feed-leg machines. The eight grizzlies in the 42.7-m (140-ft) long drift are spaced at 5.3-m

Jan 1, 1982

By Slavoljub D. Maksimovic

The Bureau of Mines, through a memorandum of agreement with the State of Ohio, tested a method of constructing seals in abandoned mine openings in Ohio by means of pneumatic stowing of crushed limestone and gravel with sand. A total of five openings and one shaft were back-filled. Pneumatic stowing uses compressed air to convey material through a pipeline and into the mine opening. The stowing equipment consists of a power supply, a blower, a feeder to inject the material into the pipeline through an airlock, and a nozzle for directing the placement. Material is ejected from the nozzle at high velocity, creating a high-compaction fill upon impact. Well-graded aggregate up to 2 inches in diameter and having sufficient fines is used as fill. The stowing method is safer and faster than conventional methods, is cost competitive, prevents trespass into abandoned mine openings, and eliminates or reduces the exposure of workers to possible hazards since the discharge nozzle can be kept at the mouth of the opening. The equipment is mobile, and, with some modification, the technique could be widely used for sealing abandoned and active mine openings, even in remote locations and during severe weather conditions.

Jan 1, 1982

By P. A. Korpi

In March of 1978 the first set of Angular Spiral Liners utilized in grate discharge ball mills for wet grinding was put into service in the San Flanuel Concentrator of Flagma Copper Company. The use of these liners has decreased grinding media consumption by 19.2% and grinding energy consumption by 15.8%. In addition a 1.1% increase in grinding throughput is being realized through the use of the system. This paper describes the Angular Spiral Lining System and reviews the test program at the San Manuel Arizona Concentrator. Operating parameters are discussed in addition to the current testing program to futher optimize the system.

Jan 1, 1982

By J. W. White, J. M. Richardson

Mass balance calculations on industrial grinding circuits are routinely performed during circuit design as well as during post-installation evaluation of operating circuits. Though such routine mass balances are straight- forward, a systematic approach can yield substantial long-term benefits. This chapter presents and evaluates several approaches useful for achieving various circuit design objectives and also considers several methods for reconciliation of data collected around an operating circuit. Numerical examples illustrate the basic concepts and techniques.

Jan 1, 1982

By R. V. Ramani, C. J. Bise

INTRODUCTION Classification of coal seams as thin, medium, and thick is both subjective and relative and is more a reflection of the statistical mode of the thicknesses en- countered in a mining region. A geological factor more precisely understood as defined than thickness is the inclination of the seams to the horizontal, commonly referred to as the pitch or the dip of the seam. Thick and steeply pitching seams present far more challenges to the ingenuity of the mining men to devise methods for their extraction than the thick and less steeply inclined ones. In a marginal case, where a thick coal seam may be minable by surface or underground methods, the decision bears a close relationship to the technological developments in the mining scene. Near surface thick coal seams are a special gift of nature and are not as wide- spread as the ones that call for the application of under- ground methods. Recent advancements in surface mining equipment have made it possible to surface mine to great depths. Stripping ratios of 30: 1 (cu yd of over- burden removed per ton of coal) have been achieved. The extraction of thick seams which preclude the use of surface methods is the concern of this chapter. Specifically, the objective here is to outline the special technological factors in underground mining of thick coal seams and to illustrate through several case histories the methods that are practiced in the North American continent. For the purposes of this chapter, a thick seam is defined as one which is not readily minable with currently available equipment without modifications to the extraction method or the equipment. At the present time, a thickness of 4.8 m (16 ft) can be considered as the lower limit, although developments in equipment and methods can be expected to push this limit upwards. The data for the case histories were collected from two extended mine visits in 1975 and 1978, and from other published accounts of the mines visited. Reference is drawn particularly to reports by Bise, Ramani, and Stefanko (1976) and Rarnani and Kenzy (1979). THICK SEAM MINING METHODS The various methods of mining thick seams can generally be grouped into three major categories: full face methods, where the entire thickness of the seam is re- covered in a single pass; slicing methods, where the seam is extracted to its full thickness by horizontal or inclined slices either in an ascending or descending order; and caving methods, where the seam is undercut and the en- tire coal seam is recovered by caving the top coal into the excavated cut, or some variations of the basic sub- level caving method. As shown in Table 1, in each of these methods, the coal is removed by either room-and-pillar or longwall methods. Hydraulic mining and transport of the coal in open flumes or in pipelines are also practiced. Addition- ally, the void created by the removal of coal may be filled with a packing material (stowing) except in the caving methods. The pitch (dip) of the seam, particularly when it is greater than 0.26 rad (150), imposes severe operational limitations on mobile equipment. A more detailed general discussion of the methods can be found in Bise et al. (1976) and Vorobjev and Deshmukh (1964). Specific discussion on the many variations of the three basic methods can be found in the papers presented at the international symposia that were held in

Jan 1, 1982