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By Slawomir Bock

One of the fundamental issues having a great influence on the safety of operations performed in underground mines is the support of workings which should assure their stability in a given geological setting and mining-technical conditions during the life of their use. The process of designing and planning support for roadways may use analytical and empirical, as well as numerical methods. In view of the rapid increase in the data processing capacity of computers and the development of more and more enhanced software for numerical modeling, it seems to be advantageous to use numerical methods in searching for design solutions ? especially in the case of non-typical mining and geological conditions and in cases that need to take into consideration a great number of factors. The assessment of rock mass quality is of basic significance in designing mine supports. Fractures, dividing the rock mass into blocks, have an essential influence on the stability of the rock mass and underground construction. This paper presents examples of the fundamental underground investigations on the rock mass conducted by the Department of Extraction Technologies and Mining Supports of the Central Mining Institute in Poland, such as: penetrometric tests, introscopic examination, laboratory testing of rock samples and ground penetrating radar. Moreover, this contribution also presents examples of applications of three-dimensional numerical modeling for the purposes of stress-strain analysis of the rock mass around mine workings and computer-aided support design by means of FLAC3D (developed by Itasca) and Cosmos/M (developed by Structural Research and Analysis Corporation).

Jan 1, 2009

By Christopher J. Bise

INTRODUCTION A mine's power system represents the driving force behind all of the extraction and auxiliary operations, because the production and transportation of mined material and the operation of equipment, such as fans and pumps, are all dependent upon a power source. Once, compressed air was the primary power source in mining; today, however, electrical power dominates. This chapter reviews the basic theories of com- pressed-air and electrical power and provides several examples of applications in mine power systems. COMPRESSED-AIR POWER General Compressed air has been, and should continue to be, an important source of power in mining operations. It has been used as a means of blasting coal in conventional mining and is widely used to operate stopers, mucking machines, and other air tools in both coal and hard-rock mines. Such applications demonstrate the suitability of compressed air for applications requiring linear motion, but it is also utilized for its reliability and safety. A compressed-air system is composed of a compressor, a receiver, a distribution network, and the air-operated machines. A compressor takes in air at normal atmospheric pressure (free air) and compresses it to a higher discharge pressure. The discharge pressure must be high enough to overcome the friction in the distribution system of pipes and hoses and deliver the compressed air to the machines at the pressure recommended by the manufacturer. The most common type of compressor used in mines is the reciprocating compressor in which the air is compressed by a piston in a cylinder. An air receiver is a container or storage tank that is located in the distribution system between the compressor and the machines. It stores compressed air, when the full capacity of the compressor is not being used, and gives a more steady flow of air to the machines. The distribution network consists of pipes, Oalves, elbows, tees, and hoses that transmit the compressed air from the receiver to the machines in the mine. It is essential that the distribution system be designed with the proper sizes and lengths of pipe, hose, and other components to keep pressure losses well within allowable limits. The air-operated machines consist primarily of drifters, stopers, pluggers (sinkers), slushers, and several other pieces of equipment. Compressor Operation Every compressor is made up of one or more basic elements; a single element, or a group of elements in parallel, comprises a single-stage compressor. Many compression problems involve conditions beyond the practical capability of a single compression stage. Too great a compression ratio (absolute discharge pressure divided by absolute intake pressure) may cause excessive discharge temperatures or other design problems. It, therefore, may become necessary to combine elements or groups of elements in series to form a multi- stage unit, in which there will be two or more steps of compression. The number of stages commonly used in reciprocating compressors is as follows: Pressure Number of stages 0- 150 pig 1 80- 500 psig 2 500-2500 psig 3 2500-5000 psig 4 When pressure is referred to as pounds per square inch gage (psig), it means the pressure (in pounds per square inch) above barometric pressure as measured by a gage. The sum of barometric pressure and gage pressure is the absolute pressure (in pounds per square inch). The gas is frequently cooled between stages to reduce the temperature and volume entering the subsequent stages, thereby reducing the work required for compression. The basic reciprocating compression element is a single cylinder compressing on only one side of the piston (single-acting). A unit compressing on both sides of the piston (double-acting) consists of two basic single-acting elements operating in parallel in one casing. The reciprocating compressor uses automatic spring-loaded valves that open only when the proper differential pressure exists across the valves. Intake valves open when the pressure in the cylinder is slightly below the intake pressure. Discharge valves open when the pressure in the cylinder is slightly above the discharge pressure. In Fig. 1, Diagram A shows the basic element with the cylinder filled with air at atmospheric pressure. On the corresponding theoretical pressure-volume

Jan 1, 1986

By John Chadwick

In the drive for faster development advance rates, railbound systems can be advantageous, in the right application. Skanska BS, the contractor to mine-owner VSK Mining, is driving an adit access deep into the Volouské Urchy Mountain in Slovakia to access the Gemerská Poloma talc deposit, some 70 km (43 miles) west of Ko?ice, the economic and cultural center of eastern Slovakia. Skanska began driving the Stolna Elisabeth drift in February 2007, extending 25 kg/m (50 lb/yd) rails (with 750 mm or 30 in. between sleepers) into the mountain to provide the main haulage to bring talc ore and waste out of the new mine once it gets into production. It will also be used for personnel transport, and is now almost complete. As a second egress and for ventilation, a 2.5-m- (8.2-ft-) diameter shaft of about 300-m (984-ft) depth will be sunk from surface down into the orebody. Stolna Elisabeth is a 4.2-km- (2.6-mile-) long drift mined at a cross section of some 11 m2 (118 sq ft). The finished profile is 3.2-m- (10.5-ft-) wide by 3.6-m- (11.8-ft-) high. It has been developed on a very slight incline so that water runs down it to the portal. The portal is 450 m (1,476 ft) above sea level and the end of the tunnel is 475 m (1,558 ft) above sea level. Much of the rock through which it has been developed is magnesite with a compressive strength of 150 MPA. The drilling jumbo is an Atlas Copco twin-boom Boomer 282 Raildrill drilling 42- to 45-mm- (1.6- to 1.7-in.-) diameter holes to a maximum length of 2.5 m (8.2 ft), depending on ground conditions. This is a drill jumbo that Skanska has used successfully on a number of projects during its life. The condition of the rock through which this drift has been driven is extremely variable. Five classifications of A1 and A2 and B1, B2 and [ ] B3, where A1 is the most competent rock and B3 is the worst, have been established. The amount of support used, of which more later is determined by the classification of ground conditions. At the time of the author?s visit, the project it was about 40 days ahead of schedule, thanks to high advance rates. For instance 6 m/day (20 ft/d) was the plan for the advance to be achieved in B1 rock, but sometimes the crews have achieved 8 m/day (26 ft/d) ? which is the average for the whole project ? in those conditions. In A1 the best advance rate achieved is 12 m/day (39 ft/d). The loading unit is a Häggloader 9HR-B on crawler tracks with a pony track to run on the rails. This allows the unit to be towed up and down the drift along the rails. To load from the face, the Häggloader lowers the foot of its onboard conveyor system to the floor and the backhoe pushes material on to it. This digging arm-conveyor combination produces uninterrupted, optimal volume loading of the haulage unit, without spill ? a technique offering load capacities as high as 180 m3/h (6,356 cu ft/hour). Further-more, as an electrically powered machine, it offers great

Jan 1, 2009

By R. Heuberger

South Africa?s Mining Industry operates in an uncertain world. The industry is buffeted by problems regarding retrenchments, unemployment, low productivity, rising costs, volatile exchange rates, commodity prices and government regulations. With these uncertainties facing the decision makers, what does applied technology offer the industry?s management? Modelling languages with risk analysis have been available for decades, but in the past their use has been very expensive, requiring large mainframe computers and lots of computer time. Today, PCs and laptops have incredible power, which makes interactive risk analysis affordable and viable for all enterprises, whether small or large. ?What is risk analysis? ?What is Monte Carlo simulation? ?What kinds of computer models need to be created to use risk analysis and Monte Carlo simulation effectively? This paper addresses these questions and illustrates, with a practical example in the mining industry, how risk analysis techniques can be used to support decisions in the climate of uncertainty.

Jan 1, 2005

The coalmining industry in most countries has, at various times, suddenly been confronted with a decrease in demand for coal. Readjustments made to reduce productive capacity to match the reduced demand for coal have often resulted in hardship to mineworkers, loss of capital investment to mineowners and loss of coal reserves. The consequences of the methods of readjustment adopted in the past have contributed to the low esteem in which coalmining has come to be regarded by a large section of the gel).eral public in many countries.Four sections of the coalmining industry, representing different periods of time, different countries, and different methods of industrial organisation have been examined:Great Britain and the U.S.A. during the inter war years, the post war European Coal and Steel Community, and post war Great Britain under the National Coal Board. An assessment has been made of the methods adopted in each case to reduce productive capacity to match a reduced demand for coal. Finally the fundamentals of the problem-as it affects the coalmining industry of any country-are considered, and possible solutions put forward.During the 1950's, surplus productive capacity was a problem in many coalmining countries, including Australia and New Zealand. As a rule no techniques are developed beforehand to solve problems introduced by a sudden decrease in demand for coal. The situation is usually tackled by a series of short term decisions based on expediency, which often in the long run only aggravate the problem.

Jan 1, 1961

By Raymond Philippe

Background ?Hatch Water has worked with most major mining companies in Chile on water supply and optimization projects ?Many(copper) mining projects are located in Northern Chile, which hosts the driest desert on earth Water supply options in Northern Chile ?River ?Availability/Rights ?Well(Brackish) Water ?Availability/Rights ?Municipal Wastewater ?Water rights ?Volume ?(Desalinated) Seawater ?Infinite source ?Cost depends on distance and altitude ?Recycling/Reutilization ?Water losses All these options present water quality issues

May 1, 2009

THE furnace described in this paper was installed and has been used by the writer for the melting of rich gold-bearing material, and has given complete satisfaction. It may, therefore, prove of interest to other metallurgists who are faced with a similarproblem. A furnace of this type should be most suitable for the reduction of "cyanide preeipitate," "charcoal ash," "anode sludge," &c. formerly the melting was done in fire-clay crucibles, using coke or crude oil as fuel, but this method is costly, slow, and unsatisfactory. A small water-jacketed blastfurnace was also tried on this material after briquetting it, but was discarded in favour of the electric furnace. The furnace may be described as a simple form of the Heroult resistance furnace, with an adjustable upper electrode and a fixed bottom electrode. This is in accordanee with Stansfield's classification of electric furnaces. The heating is mostly due to the resistance offered the current by the molten slag. The melting chamber consists of a steel drum 14 in. in diameter and 21 in. high, and is lined with silica bricks specially made to fit the shell. The internal diameter is G in. The bricks contain 92 % silica, are made from calcined quartzite bonded with fireclay, and are similar in composition to those used for lining steel converters. They are set in a mixture of finely-ground quartzite and clay, and well dried out. The shell is bolted to a steel plate, which forms the decking of a truck running on rails, so that the furnace can be moved under ahood for carrying off the fumes. A cast-iron slag spout is bolted to the shell below the tapping hole, which is 4 in. from the bottom of the shell. Opposite this another hole 1 in. in diameter is drilled through the shell and lining for the passage of the lower...

Jan 1, 1918

By Fred H. Perfect

The thermite process, now approximately 80 years old, has heretofore been practiced in ceramic vessels. This paper outlines the successful production of several million pounds of alloy produced in zuater-cooled copper vessels. THE high temperatures necessary to produce fluid high-alumina slags have probably discouraged earlier workers from attempting to utilize water-cooled vessels. The following is a report of the successful application of this method, together with a report of the physical and chemical problems which had to be solved to make the process practical. About 1955, when Reading Alloys was more concerned with the production of ferrotungsten than with most other types of alloys, it was found to be exceedingly difficult to predict the reaction rate that one would experience from one tungsten ore to another . This difficulty was increased for two reasons: the physical state of the ore as received was seldom uniform, and the size of the mass to be reduced is a very important factor in developing the proper mix for the final reaction rate and temperature. For example, the reduction of ferric oxide with aluminum would theoretically produce temperatures of about 2600°C, and, in spite of the losses due to radiation, the energy is quite sufficient to produce a very fluid slag and iron effecting a very clean separation. On the other hand, the reduction of ferrous oxide with aluminum, while it is not as energetic as the first reaction, nevertheless should give temperatures above 2000"C, which should cause separation of alumina and iron, but, in practice, this is not the case and good fluid slag is not obtained. The heat of reaction, calculated from the thermo-dynamic heats of formation of the constituents of the reaction, offers a convenient guide to the feasibility of a thermite reaction. Thus, the values listed in Table I can be used as an approximate guide for estimating the results which may be expected in the reduction of various oxides and combinations of oxides using aluminum as the reducing agent. As a generalization, those reactions releasing approximately 600 cal per g of total mass represent the lower limit at which to expect separation of slag and metal. On the other hand, those reactions which release 1100 to 1200 cal per g are too violent to be used successfully without the addition of considerable inert material, such as lime, fluorspar, and so forth. Reaction rate is also a very important factor. This, of course, is governed very largely by the sizing and intermingling of the oxidant and reductant and by the starting temperature of the reactants. This is why those figures shown in Table I can only be a general guide. Further, it must be understood, that by adding powerful oxidizing agents, such as nitrates, chlorates or metallic oxides, such as molybdenum trioxide or chromium trioxide, certain problems respecting low temperatures are not necessarily solved. This is true in the addition of oxidants, such as sodium chlorate. By adding too high a concentration of sodium chlorate as a booster for a particular reduction, explosions are generated. Generally, they are of an intermittent type so as to give a Roman Candle effect to the reduction itself; with each surge resulting from a minor explosion more and more material is thrown

Jan 1, 1968

By Erle C. Donaldson

A study of subsurface waste disposal in the United States shows that in eight States a wide variety of industrial wastes are being injected into formations ranging in age from Precambrian to Recent. More than 30 wells ranging in depth from 300 to 12,000 feet are used for waste disposal into subsurface formations which include unconsolidated sand, sandstone, vugular limestone, and fractured gneiss.

Jan 1, 1964

By Tatsunori Matsumoto

In order to investigate the variation of the load-displacement relations of piles in a site, vertical load tests were carried out on 25 small diameter steel pipe piles. Out of 25 piles, 15 piles were close-ended pipe piles attached with a blade at the pile toe and 10 piles were open-ended pipe piles. To carry out these load tests in a short period, two types of rapid load test methods, Spring hammer load test and Falling mass load test, were employed. The load tests of all piles were completed in two days with 180 blows in total. In order to estimate a static load-displacement of a pile, non-linear damping method was used. It is shown from the test results that the performance of the pipe piles such as the ultimate pile resistance and the initial pile head stiffness is improved by re-driving.

Jan 1, 2006

Open pit mine plans define the complex strategy of displacement of ore and waste over the mine life. Various mixed integer linear programming (MILP) formulations have been used for production scheduling of open pit mines. The main problem with these models is the inability to solve real size mining problems. We present a mixed integer linear programming formulation for the open pit long-term production scheduling problem. To reduce the number of binary integer variables in the formulation, we aggregate blocks into larger units referred to as mining-cuts. We present a semi-supervised clustering algorithm which aggregates smaller blocks based on rock types and grade ranges into mining-cuts without introducing an error in the grade estimated values. The cluster-based aggregation algorithm is fully integrated into an MILP mine production scheduling prototype software. We have implemented, verified, and validated our models. The efficiency of the new platform is evaluated through iron ore mine case study.

Dec 6, 2010

By Robert M. Valentine, Randall L. Lentell

Interpretive geotechnical investigations were performed for the construction of approximately 15,250 m (50,000 ft) of wastewater interceptor tunnels in Austin, Texas. Subsurface materials consist of claystone (or clay shale), siltstone, altered volcanic tuffs, chalk and limestone. Special conditions include multiple crossings of Onion Creek with cover ranging from about 1/2 to 3 m (2 to 10 ft.), and the potential for mixed-face conditions consisting of water-bearing gravels overlying clay shale or siltstone. The methods and results of the geotechnical investigations are discussed and the inferred construction conditions are compared to those encountered during tunneling. A suggested format is presented for interpretive geotechnical reports for underground construction projects.

Jan 1, 1985

By Erle C. Donaldson

A study of subsurface waste disposal in the United States shows that in eight States a wide variety of industrial wastes are being injected into formations ranging in age from Precambrian to Recent. More than 30 wells ranging in depth from 300 to 12,000 feet are used for waste disposal into subsurface formations which include unconsolidated sand, sandstone, vugular limestone, and fractured gneiss.

Jan 1, 1964

By George K. Burke

Jet grouting was first introduced into North America in the late 1970's, soon after the conception of the system in Japan. However, it was not until the mid-1980's that the technique began to gain acceptance through the renewed interest and efforts of specialist U.S. geotechnical contractors. Since this time, innovations by both American users and other users worldwide have further advanced the technology, with a corresponding increase in the range of practical applications. This paper discusses the evolution of jet grouting in North America, from simple single-fluid methods to more complex triple-fluid methods, to the energy-intensive double-fluid and SuperJet methods. Case histories, illustrating a variety of jet grouting methods and applications, are cited to support the newest United States applications.

Jan 1, 2002

Tougher physical mining and processing conditions as well as economic circumstances are causing many mining companies to re-evaluate the effectiveness of their mine planning. The recent wave of inexpensive computing power provides the engineer an opportunity to go beyond the simple one-off manual mine design and production schedule. The basis of a good mine plan is a good reserves model. Concepts of mine model development are explored. The only advanced computer methods that have found consistent success in optimising mine planning functions are economic pit limits determination and material scheduling. Three pit 'optimisation' methods are examined; the moving cone, the Lerchs-Grossman 3-D graph algorithm, and the new network flow algorithm. A comparison between the methods in terms of total profits generated and computer time involved are made. The results show that both algorithms produced a higher profitability than the simple moving cone, but at much higher execution times. However, the profit increase never exceeded much more than 1 % of the base moving cone case. Finally, a method for the development of long range material production schedules with forward waste planning is presented.

Jan 1, 1989

By R. G. B. Pickering

The purpose of this report is to chronicle the interaction between Anglo Platinum and Sandvik Tamrock during the development and application of Long Hole drilling as a viable, practical method of narrow reef mining. Appendix I identifies the people from Anglo Platinum and Sandvik Tamrock who participated in the project. Experiments in long hole drilling and blasting were conducted in a number of mines during the late 1950s and early 1960s. Some of the advantages envisaged were: ? The face would require no daily examination ? No barring down or immediate face support required ? No sockets to clean ? Only a few well drilled holes rather than large numbers of short blast holes ? Avoidance of poorly-marked holes, inaccurate drilling and improper charging ? Better supervision and better face advance. Holes 50 mm in diameter were drilled parallel to the stope face by pneumatic rock drills from advance strike gullies. Holes were charged with amon gelignite cartridges. All holes were traced with detonating cord to ensure propagation of the explosive along the entire hole length. Rock was blasted onto a rolling scatter pile. A number of reasons have been cited for the failure of this mining method. The two most important were: ? problems in controlling hole deviation which was exacerbated by the fractured ground ? difficulty of charging up holes in fractured ground, due to relative movement of rocks along fracture surfaces over the hole length. The idea of drilling a few long holes parallel to the face rather than a large number of short blast holes perpendicular to the face was an attractive option to mechanize narrow reef mining. This particular project started in 1998. Initial trial The Union Section of Anglo Platinum purchased a Solo H 606 RA long hole drilling rig from Sandvik Tamrock in 1997. In order to evaluate long hole drilling as a basis for face mechanization. The Solo series of drill rig was selected because it combines high drilling performance and good drilling accuracy with high versatility with a compact size. The mining method proposed by a consultant to Anglo Platinum was based on the current trackless mining method practised by Union Section. The trackless mining method is better described as a trackless development and transport method. Reef strike drives were installed approximately 40 metres apart on dip using trackless equipment. Reef was mined using hand-held pneumatic rock drills and scraper winches and the excavation was supported by sticks and crush pillars at the top of the panel. Rock scraped from the panel was loaded by Lad Haul Dumps (LHDs), trammed to an inclined haulage, and loaded into trucks for transport to the surface. It was intended that the Solo would drill from the bottom advance strike drive to the top advance strike drive. Blasted rock would still have to be scraped from the panel. It is also difficult to understand how well the support would have survived. In the initial motivation it was stated that drilling accuracy was of the utmost importance. Previous work by Sandvik Tamrock had shown that, in other parts of the world, 20 metre long holes could be drilled with a deviation not exceeding 100 mm provided that: ? holes were collared with low thrust pressure ? operating percussion pressure was kept low -often not ? exceeding 10 MPa ? drilling was to be with tubes, not rods. In early 1998 initial trials with the Solo H 606 RA were inconclusive; attempts were made to drill 40 metre long holes from one advance strike drive to another. Difficulty was experienced in measuring hole straightness and accuracy, there was no survey of the holes, and no blasting was done to expose hole barrels. A long hole mining system was then defined, with a corresponding face mining system, which led to the design of a test programme for the Solo drilling rig. Earlier experience at Telfer The proposed face mining system was based on a mining method that is practised at Telfer Gold Mine in Western Australia. Telfer has a series of domed reefs that dip at angles of less than 20° to more than 40°. The ore bodies are all thin, with widths from 0.6 metres to 0.8 metres. Figure 1 shows the face layout. Experience at Telfer indicated the following: ? The spacing and size of the dip pillars should be based on the rock mechanics of Union Section ? The length of the face is determined by the ability to correctly position the blast holes. Positioning of blast holes is a function of hole straightness and hole direction. Blast holes in Telfer were drilled 64 mm diameter with retrac bits from Sandvik. Extension steel was T38 by 1.8 metre lengths. A guide rod was fitted immediately behind the bit. To achieve greater accuracy' operators drill at relatively low percussion and thrust pressures. Results show that very good accuracy is achieved in 20 metre stopes and only fair accuracy in longer stopes. In 34 metre long stopes 10-15% of the holes are not sufficiently accurately positioned and have to be re-drilled. All blast holes are

Jan 1, 2006

By Cyril Wells, Paul E. Busby, Donald P. Hart

EARLY workers in the field have established that the diffusion of nitrogen follows normal diffusion laws. Concentration-penetration data from layer analyses of reasonably pure iron specimens nitrided in an atmosphere of CO-NH, are given in the literature,' and were utilized to calculate diffusion constants for nitrogen. The diffusion of nitrogen in y-iron may be represented approximately by the equation D - 1.07x10-' e-:".'m'"'r The diflusion coefficients, D, given by this relation represent mean values, since carburization occurred simultaneously with nitriding. The rate of diffusion in a-iron has been determined from calculations based on data obtained by measuring the internal friction of samples at each of several frequencies and temperatures.' The expression for the diffusion of nitrogen in a-iron derived from these experiments is D= 1.4x10-1 e-it tudent Increasing carbon content decreases the diffusion of nitrogen in a-iron.3 Oxygen, either dissolved or combined, has been shown' to exert a marked influence on the diffusion of nitrogen from a nitriding atmosphere. D values increase rapidly with increasing amount of dissolved oxide in the metal, but diffusion is completely inhibited if preliminary oxidation has produced the slightest trace of iron oxide on the surface. Diffusion coefficients have also been calculated from permeability and solubility data,' but these values are of questionable accuracy since the boundary conditions assumed may not actually exist. Experimental Method The method selected for determining the diffusion coefficient of nitrogen in a-iron in the present investigation was the so-called fractional saturation method. This method consists of heating the specimen in an atmosphere of H2-NH8 mixtures of known composition for a predetermined time. Following this treatment, the specimen is analyzed for nitrogen; and the diffusion coefficient is calculated from a known relationship of the fractional saturation, time of diffusion, and dimensions of the specimen. This relation is based on a solution of Fick's second law and assumes that the surface of the specimen immediately reaches a given concentration which remains constant during the diffusion anneal. Tabular solutions are available4-5 and provide the relationship between fractional saturation and the quantity \/Dt /a, where D is the diffusion coefficient in cm2 per sec, t is the time of the diffusion anneal in seconds, and a, in the present experiments, is the half-thickness of the slab specimen in cm. The method is adequately discussed elsewhere.9-10 Material The material for this study consisted of 3/32 in. thick (0.24 cm) carbonyl iron plate. The major impurities in this iron are oxygen and carbon, which are readily removed by long time high temperature treatment in an atmosphere of purified hydrogen. In order to reduce the diffusion anneal time for saturation runs, some of this material was rolled to a thickness of 0.015 in. (0.038 cm) prior to the hydrogen treatment. Hydrogen treatments were carried out in a hydrogen-purified Armco iron tube heated in a global

Jan 1, 1957

Mt Waruwari, host to the Porgera underground gold mine in the Highlands of Papua New Guinea (PNG), is a prominent ridge of sediments and dioritic intrusives rising approximately 600m above the surrounding valley floor. Consequently, mining occurs in a low stress environment where structural and material considerations are foremost in the design of ground support and reinforcement. Three major rock types are encountered in the underground mine. These are locally referred to as "black sediments" (slightly indurated mudstones and clayey siltstones), "altered sediments" (derived from black sediments by thermal metamorphism), and "intrusives" of dioritic composition. The rocks are moderately to highly fractured, the black sediments being generally weak and occasionally sheared while the altered sediments and intrusives are quite strong but often brittle. Mining takes place from a series of stopes extracted by either slot and mass blast or a modified vertical crater retreat method (VCR), which are backfilled with a cemented rock fill. Ground support and reinforcement has been relatively conserv- ative to date and generally relies on pattern reinforcement of all development with cement grouted "rebar" bolts and mesh, extending to pattern cable bolting with occasional mesh in stope backs. In addition there has been considerable reinforcement of stope hangingwalls in some cases. A programme of broad scale structural mapping, monitoring with extensometers, material property testing, and stress analysis has been devised to allow optimum design of ground support and reinforcement in the future. Considerable work remains to ,omplete this task.

Jan 1, 1991

By Gordon King

The City of Atlanta Water Department required the construction of a two sludge tanks and a wet well at the existing Chattahoochee Water Treatment Plant. Western Summit won a design build contract to build the project and proceeded to evaluate the wet well size options, and construction procedures. The final design required the construction of a 1 million gallon tank. The proximity of the underground tank to adjacent roads and existing plant structures ruled out laying back the excavation. As a result a tied back sheet pile retaining wall was initially proposed.

Jan 1, 2003

By Robert S. Warfield

Eight samples of high-volatile bituminous coal, each large enough for a study of its coking properties by the Bureau of Mines, were obtained from a location along the Kukpowruk River, Arctic Northwestern Alaska. Bench scale and pilot plant carbonization studies with several patterns of coal blending were conducted. The data from these tests were compared to data from similar tests on a coking blend used by the Kaiser Steel Corporation at Fontana, Calif. The majority of the samples representing the 20-foot seam was found to have significant coking properties, and produced coke approaching metallurgical quality when blended with selected coking coals. Although oxidation effects were evident in the samples representing upper portions of the seam, it appears that the whole seam could be used.

Jan 1, 1966