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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

By H. Bernatowicz

The Nucoalyzer (On-Line Nuclear Analyzer) is a new method recently made comercially available for monitoring the elemental composition of a process stream. Applications for the Nucoalyzer include grading ore on a conveyor belt before or after it has gone through primary crushing, monitoring the slurry being transported in a pipe to the concentrator, and monitoring the stream of concentrate or the tailings from the concentrator. The Nucoalyzer may also be used at the smelter to monitor elements in the feed that include copper, sulfur, iron and silicon where a uniform feed is desired. The advantage of the Nucoalyzer over most other techniques are (a) it performs the measurement directly on the process with no sampling necessary, (b) the analysis can be performed on streams of particulate and slurries, and (c) the system does not have to make contact with the material it measures. These Nucoalyzers are operating at coal fired power plants and several more are about to be placed in service. This paper discusses the principles of operation of the Nucoalyzer, our experience in the field, application of minicomputers to data reduction, and potential applications in the ferrous and non-ferrous industries.

Jan 1, 1982

By Richard H. Sibson

Changes in seismic style along the San Andreas fault system appear to correlate with strike-parallel variations in regional heat flow. Large earthquake ruptures are associated with colder segments of the fault zone, while comparatively high levels of microearthquake activity and small to moderate shocks occur along the hotter segments, often in association with aseismic creep. Such behavior is consistent with geotherm and strain rate dependent profiles of shear resistance versus depth, constructed for quartz-bearing rocks. These profiles suggest that the transition from frictional to quasi-plastic behavior with depth should increase by a few kilometers passing from the hotter to the colder portions of the fault zone, giving rise to long-wavelength concentrations of distortional strain energy near the base of the seismogenic zone in the cold segments. Some support for this hypothesis comes from the observed deepening of microseismic activity passing northwest and southeast from the central Californian active region to the currently locked, ~400 Ian segments of the San Andreas fault that ruptured in 1906 and 1857.

Jan 1, 1982

By D. S. Burgess

About 100 tests of flammability were carried out with H2, CH4, and CO in mixtures with air using the Federal Bureau of Mines 12-ft-diameter spherical pressure vessel as an explosion test chamber. The LeChatelier approximation gave slightly conservative estimates of the limits of downward flame propagation (onset of significant pressure rise) using the following measured values for the individual gases, in percent: H2, 8.5; CH4, 5.25; CO, 13.0. No justification could be offered for using the LeChatelier law for upward flame propagation limits and indeed reported data suggest that calculated values would not be conservative. The observed upward limit for methane-air mixtures (4.85 pct) was less than the widely accepted lean limit for this fuel; those for lean hydrogen-air mixtures and carbon monoxide-air mixtures did not differ from the lowest accepted values. Diffusional effects on the propagation of single, binary, and ternary fuel mixtures are discussed.

Jan 1, 1982

By A. L. Hinde, A. A. Bradley, K. H. Stanton, P. J. D. Lloyd, G. K. Schymura

As part of its programme to develop methods for concentrating minerals underground, the Chamber of Mines of South Africa in conjunction with Lurgi evolved and tested a I MW centrifugal mill. The machine is capable, in the autogenous mode, of grinding more than 25 t/h of minus 75 plus 44 mm hard quartzite to a minus 3 mm product containing over 50 per cent minus 75 µm material, and as a ball mill of grinding more than 60 t/h of minus 19 mm quartzite to a minus 3 mm product containing over 35 per cent minus 75 µm material. Extrapolation of computer models of the process suggests that these limits can be comfortably exceeded. Long-term continuous testing is in hand, aimed at proving that this I m in diameter by I m long mill is in every respect the equivalent of a conventional 4 m by 6 m ball mill.

Jan 1, 1982

By Moffatt G

Approval to operate Longwall No. 7 at Appin Colliery in 1979 was granted pursuant not only to the usual statutory provisions, but also that an adequate means of drainage of methane gas be incorporated. A methane drainage plant was subsequently designed and constructed some two km from the pit top at the site of the Colliery main upcast and down cast shafts. Modern plants of this nature are completely automated using programmable controllers and computers to provide control and management reporting functions. Complex mechanical systems are installed to extract the gas and in handling and storage processes. Gas utilisation plants exist which can generate power from methane. These plants are further enhanced by harnessing waste heat produced in the generation process. The installation, automation control and monitoring of these plants incorporates many facets of engineering from high voltage power supply and computer technology to complex gas handling procedures used in gas compression storage.and generation of electric power.

Jan 1, 1982

By Edmundo Tulcanaza

Increasing exploration costs and the need to reduce the time-frame from exploration to production presently demand formalized procedures to assist exploration planning and reserve analysis. Exploration planning is basically required to select the most cost-effective drilling/sampling program, and to flag the needs for data of greater relevance and reliability. Reserve analysis is necessary to determine the sensitivity of grade-tonnage relations, and to clarify the economic implications of mining when alternative technical choices are to be considered.

Jan 1, 1982