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Although there was a significant improvement in the world economy, recovery in the iron and steel industry followed its traditional pattern and lagged the general economic recovery in 1983. World crude steel production increased from 645 Mt (711 million st) in 1982 to 664 Mt (732 million st) in 1983, an increase of less than 3%. Domestic iron and steel production showed a modest increase over 1982. Weekly raw steel pro¬duction levels ranged between 1.5 and 1.75 Mt (1.6 and 1.9 million st) for the last 10 months of 1983, compared with less than 1.25 Mt (1.4 million st) for the last half of 1982. The peak was 2.5 Mt (2.7 million per st) per week in 1978-79. This how level of steel operations kept most pelletizing and sintering operations at curtailed production levels. Extended shut-downs of pellet plants were also common. Pressure by management to reduce operating costs has continued with an increasing emphasis on maintaining quality and efficient unit productivity. With the extended decline, there have been significant personnel reductions in both operating and headquarters staffs. Many companies are anticipating a long-term how growth period in the iron and steel sector. Pelletizing Domestic pellet production showed a slight increase last year to about 34 Mt (38 million st), less than 50% of total capacity. What is indicative of the inevitable is the announced intention to chose down facilities, such as US Steel Corp.'s Atlantic City ore operations pellet plant at Lander, WY, and the recent aborted move by US Steel to purchase National Steel with the apparent intent to chose the National pellet plant in Keewatin, MN. There is excess domestic pellet capacity that will not be needed. The question is what form attrition will take. South American pellets can be brought into the Chicago area at a cost competitive with Great Lakes plants. This raises the possibility that imported pellets can, in fact, displace domestic pellets at almost any US steel plant. The need to re duce total production costs at domestic pelletizing plants is crucial. Efforts to reduce fuel costs for pelletizing have focused on re placing natural gas or fuel oil with coal. More recently, petroleum coke has become available at an attractive cost, but systems must cope with its higher sulfur con tent, which limits its potential The direct firing of grate-kiln systems with pulverized coal hay been fully proven. The use of solic fuel in the straight-grate process however, is still being developer and no straight-grate system i running on solid fuel alone. The further development o coal-water slurry systems may provide a suitable fuel for both systems. It would appear, how ever, that the same ash quality restriction would hold for coal water systems as for direct-firin systems. So, the choice is based of the relative cost and convenience of a wet-grinding system with additives, compared with a dry-grind direct-fired unit. The high-temperature (two stage) heat recuperation system installed at US Steel's Minnta plant (Step III) has provided the fuel savings anticipated, any modification of additional in durating lines is planned. Other operators are hooking at similar heat recuperation improvement. Another area of potential cost savings has been in the reduction of bentonite requirements. A CCI's Tihden plant, major improvements have been made in this area. Progressive bentonit reductions were achieved over several years from the 8- to 9-kg (16- to 18-lb per st) range to the 5-k (12-1b) level by progressive moisture reductions and chose attention to operating practice. Further reductions were achieve by raising bentonite specifictions. In this instance, the wate plate absorption test specifia

Jan 5, 1984

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By Kendrick C. Hardcastle

Although the basic techniques employed in hydraulic pumping in open pit mining have been practiced throughout the better part of this century, only relatively recently have studies of' sufficient depth and scope been conducted to improve these techniques. This paper will point out the basic techniques involved and some of the more recently discovered data concerning hydraulic pumping, Inherent in all hydraulic transportation systems apart from the consideration of transportation cast, is the outstanding advantage of slurrying the matrix at the mining pit so as to provide a lengthy and thorough soaking and washing of the material while in the pipe line and thereby reduce the necessity for washing, scrubbing and other treatment in the beneficiation plant. This is quite important in view of the constant increase in demand for a cleaner product.

Jan 1, 1964

By Harry R. Dahlberg

Heap leaching is discussed as a unit process. Five factors are identified as the principal elements that can be measured and/or controlled. These are: 1. Material Size 2. Volume of solution 3. Tonnage of heap 4. Time 5. Solution concentretion. The Rate of Extraction Curves obtained from an actual heap or column tests describes the economic viability of the unit process. By analysis of these curves and the five factors end their inter-relationship it is possible to determine how the extraction rate can be increased. The practice of leach and rest used by many operators is analyzed and shown on a hypothetical rate of extraction curve. The influence of the driving force of the hydrogen ion and ferric Lon as Lixivants and its effect on the rate of extraction curve is discussed.

Jan 1, 1979

By Peter J. Lawson, John A. Rideout, Gregory W. Dicinoski

A considerable body of literature directed towards optimising the efficiency of gold and silver extraction from cyanide leach liquors has appeared in recent years. A few examples from dissolution to solvent extraction to adsorption in carbon-in-pulp (CIP) and resin-in-pulp (RIP) operations follow. Selective leaching of gold from copper-gold ores using ammonia-cyanide mixtures extends the % copper which may be tolerated in ores (Ruane, 1991, LaBrooy et al., 1991, Costello, 1991). The removal of copper and recycling of cyanide from high copper ores is another important development (Sceresini and Staunton, 1991). Solvent extraction of alkaline cyanide liquors (Riveros, 1990, Mooiman & Miller, 1991) and solvent aided elution in CIP (Muir, et al., 1985) look promising. Recent definitive papers on the mechanism of adsorption of aurocyanide by activated carbon should lead to more efficient use of CIP (Adams and Fleming, 1989, Klauber, 1991). Strong and mixed weak-strong base resins have been applied in RIP for aurocyanide extraction in Russian Plants since the 1970's (see for example Borovskii et al., 1985, and references cited in Fleming, 1988a). Fundamental studies on applications of commercial weak and strong base resins-in-pulp (Fleming & Cromberge, 1984a, b, Fleming 1988a, Hosking 1984, Riveros and Cooper 1987) have generated much interest in RIP. Detailed comparisons of CIP vs RIP efficiencies at Mintek in South Africa (Fleming et al., 1988b), and the inefficiency of CIP in contact with clay led to the commissioning of the strong base RIP Golden Jubilee plant in 1988. Its operation and improvements have been reported (Seymore and Fleming, 1989, Fleming & Seymore, 1990). The use of anion exchange membranes for aurocyanide extraction shows some promise (Mauthoor et al., 1991). However, the development of new gold specific resins at Mintek (Green et al., 1984, 1986, 1990, Johns and Green, 1991) and elsewhere (Asker et al., 1987, Wilson et al., 1987, Harris et al., 1992) should improve the viability of RIP processes. The polydiallylamine gold specific resins (Hodgkin & Eibl, 1988, Fawell et al., 1992a, b) have high capacities for auro cyanide recovery from high concentration liquors. The use of guanidine based liquid and solid ion exchange extractants for gold was recently reported by the Henkel Corporation (Kordosky., et. al., 1993). Such specific resins tailor made for gold and/or silver cyanide extraction should lower overall cost by improving metallurgical efficiency. In the case of strong base resins this selectivity largely eliminates the problems of poor base metal elution and/or poisoning. The work reported here outlines the approach taken at the University of Central Queensland to synthesize sterically hindered strong base anion exchange resins which are designed to have high selectivity for auro- and argentocyanide complexes. It summarises selectivity tests on low grade cyanide leach liquors from several mines. It also examines an RIP vs CIP comparison, and investigates the elution properties and cycling of one such promising resin.

Jan 1, 1993

By Y. B. He

The effect of the stability and rheological properties of magnetite dense medium on the separation performance of a 6" dense medium cyclone (DMC) is studied with the use of fine density tracers. The tests in which different size fractions of magnetite are used to prepare the dense media reveal that at low medium densities (<1.5 s.g.), the best separation is achieved when using the magnetite with intermediate particle sizes (du3.2=18 11m). At medium densities above 1.5 s.g., however, it is beneficial to use coarser magnetite. With the fine magnetite dense medium, an increase in medium density results in a decrease in separation efficiency. With the coarse magnetite, an opposite trend is observed; the separation efficiency is improved with increasing medium density. An accurate cutpoint control is critical in gravity separation. The cutpoint shift for coarse feed particles is found to decrease with increasing medium stability (density). It assumes values close to zero in a highly stable fine dense medium. When the size of the treated particles is small (<0.5 mm), the cutpoint shift is mainly determined by medium rheology. With increasing medium density, a continuous increase in cutpoint shift is observed.

Jan 1, 1993

By R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981

By Donald Towse

Portland cement is a basic building material in every country in the world, and, fortunately, the raw materials required in its manufacture are widely distributed. It is a low-priced bulk commodity (price one cent a pound or less, with a density of 100 pounds per cubic foot.) Table 1. Approximate price per pound of selected mineral products. Copper 48 cents per pound Aluminum 27 Lead and zinc 15 Asbestos 3 to 18 Pumice 6 Sulfur 2 Bauxite (abrasive) 1.8 Barite 1 to 1.6 Cement 1 Primitive regions may get along without any Portland cement, but most countries use some, and it is one of the most widespread building materials, even if used in small quantities. Use of cement begins early in a country&apos;s development, and, because of local availability of raw materials, cement manufacture is one of the first heavy industries many countries desire. Use and manufacture of cement usually precedes other industrial materials such as steel and heavy chemicals, and consumption levels off for a long period during development of infrastructure and an economic base. With a mature economy comes a very high consumption. Per capita consumption varies ten times or more between nations (Figure 1).

Jan 1, 1970

By Randolph E. Scheffel

Mill design and project execution are very demanding and interactive processes requiring the dedication and cooperation of many parties. The development philosophy which lead to the successful completion of Ranchers Exploration and Development Corporation&apos;s Escalante silver mill is reviewed. The project spanned 25 months from initial feasibility studies to start-up with the major construction effort requiring only 4 months. Mill performance for the second and third months after start-up, as well as fiscal 1983, is compared against the design parameters. Capital and operating costs are also discussed.

Jan 1, 1983

By G. A. Swanquist

INTRODUCTION This article describes the yellow cake drying, packaging and dust collection circuit of the United Nuclear Church Rock Mill. It also describes the layout and floor plan of this section as well as the operation and performance of the various equipment. In order to place the yellow cake section in perspective of the mill as a whole, the mill circuit is described briefly. CHURCH ROCK MILL GENERAL DESCRIPTION The Church Rock Mill contains six operating sections shown in Figure 1 in block diagram form. The grinding section consists of a semi-autogenous mill in closed circuit with a hydrocyclone. The slurry is then routed to the leaching tanks. Sulfuric acid and sodium chlorate is added to dissolve the uranium. The pulp is transferred from the leach tanks to the thickeners where the uranium bearing liquid is separated from the barren solids. The uranium bearing solution is then filtered and pumped to the solvent extraction section. [ ] In the solvent extraction section, the uranium bearing aqueous solution is brought into contact with an organic solution composed of kerosene, amine and isodecanol. The amine extracts the uranium from the aqueous solution. This organic solution is then routed to the stripping section, where the uranium is removed from the organic solution with an ammonium sulfate aqueous solution. The pregnant stripping solution now concentrated to 30 gpl U308 is routed to the precipitation tanks where ammonia is added to precipitate the uranium. The yellow cake is then washed in a counter current circuit of two thickeners and one centrifuge to remove impurities, mainly sulfates in order to meet specifications. The centrifuge centrate is routed to the second wash thickener for forward wash. The centrifuge cake at approximately 50% solids consistency is transferred to the yellow cake dryer via a screw conveyor. YELLOW CAKE DRYING, PACKAGING AND DUST COLLECTION Figure 2 shows the flowsheet of the integrated circuit of drying, packaging and dust collection. This section of course centers around the yellow cake dryer as the main equipment. The other main units are the dust collectors for the dryer and packaging respectively, a yellow cake storage bin and the yellow cake packaging facilities. Physically, all of this equipment occupies four floors, beginning at the top floor: 1. Both dust collector fans, the centrifuge and the screw conveyor to transport the yellow cake to the dryer. 2. The dryer, fire box and both dust collectors occupy the third floor. Figure 3 shows the floor plan of the third and fourth floors. 3. The yellow cake storage bin and dryer dust collector sump are located on the second floor. 4. The ground floor houses the loading facilities consisting of two load out ports from the bin to barrels, a roller conveyor loop, drum weighing scale, and change room facilities. Figure 4 shows the ground floor plan. Figure 5 is an elevation view of this section showing the overall equipment arrangement.

Jan 1, 1981

By John Ralston, Rossen Sedev, Nicola Richards, Quinn Anthony, Geoffrey Reynolds

Derivatised 2,4-diketopyrimidine groups have been anchored to planar gold surfaces as self¬-assembled monolayers (SAMs) for chemical and photochemical studies. Three derivatives, 5-methyluracil (thymine), 5-trifluoromethyluracil and 5-nitrouracil, were chosen for the specific moiety present at position C-5, which can exert strong or weak inductive effects on the adjacent aromatic ring. The measured acidity constants decrease from 5-methyluracil to 5-nitrouracil as the inductive properties of the group at C-5 position increase. When grafted to the surface the acidity constants increase, albeit still adhering to the expected inductive trend. The functionality also exerts an influence on the surface wetting properties. Surface wettabilities indicate that the orientation of the pyrimidine groups at the surface occurs so that the C-5 position has a direct effect upon the surface chemistry. Upon dimerisation the surface contact angles increase signifi¬cantly. At pH 11 the contact angle changes reflect that the dimer has a higher pKa than the monomer. The monomer surface is deprotonated at this pH while the dimer surface is partially deprotonated, leading to changes in contact angle of up to 22 degrees. Large changes in contact angle were observed for 5-nitrouracil which were independent of surface charge and were appar¬ently due to conformational changes alone. Electrowetting of three fluoropolymer surfaces by electrolyte solutions was studied with the sessile drop method. The electrowetting curve (contact angle/potential) is analogous to the elec¬trocapillary curve (surface tension/potential) and may be described by a combination of the Young and Lippmann equations. We demonstrate that for some specific surfaces a systematic deviation occurs at positive potentials. This departure is attributed to double layer effects, namely the adsorption of hydroxide and halide anions. pH, ionic strength and polymer composi¬tion can all influence electrowetting behaviour.

Jan 1, 2003

By Shinya Ono, Tsuyoshi Katoh, Yuusaku Tominaga

In order to monitor/control airflow rate distribution in a complicated ventilation network installed many regulators, an integrated system composed of expert system, exerting programs for simulation to get macroscopic characteristics and drawing diagram, monitoring system and so on is introduced in this study. This system is made up of airflow velocity sensors and computers linked to an local area network. The main results of this study are as follows; 1) Combinations of two branches to monitor aerodynamic resistance change in a branch p in a network can be selected by comparing factor&apos;s value in a branch relation table. 2) When the combination of sensor branches for a branch p does not exist in the network at present condition, the system changes aero- dynamic resistances of appropriate branch q placed a regulator by referring branch relation table. And then the system asks other computer to get macroscopic characteristics under altered value of aerodynamic resistance of the branch q. 3) Dot pattern diagram shows airflow rate relation between an active branch and a passive branch. Three relations such as a) a straight line, b) a monotonous curve and c) extremums are displayed in different patterns for all the sets between an active branch and a passive branch at a time. 4) The ranges of airflow rate in coal faces by changing aerodynamic resistance of any one branch placed a regulator are shown in the Pareto diagram using values in the branch relation table. 5) Ventilation transfer diagram with two parameter, one shows airflow rate in a coal face and the other shows energy provided by main fans to overcome aerodynamic resistance in a net- work, is utilized to decide countermeasures for decreasing energy consumption in the network on increasing of airflow rate in the specified branch.

Jan 1, 1995

By J. Siege1, R. S. Popielak

A computer model was used to simulate the effect of leakage through confining beds on a hypothetical in situ leach operation in an aquifer with a thin mineralized section. The simulations were used to assess economic and potential environmental implications of leakage on an in situ leach operation. Four scenarios were modeled - three representing cases with different degrees of confinement, and the fourth representing a stratified ore zone under a low degree of confinement. The model simulated the travel path and travel time of lixiviant given the hydraulic conditions prescribed in each scenario. The results show that the travel path and travel times vary by about 10% for confinement ranging from nearly impermeable (essentially no leakage) to an extremely low level (thin confinement with a permeability about equal to that of the aquifer) for isotropic aquifer conditions. The aquifer thickness contacted by the lixiviant varied from 85% to 90%. Horizontally, 55% to 60% of the aquifer was contacted. The flow paths in the stratified aquifer were reduced in the vertical direction and expanded horizontally. Forty- five percent of the aquifer thickness was contacted vertically with 88% contacted horizontally. Break- through time for lixiviant traveling at mid-depth of the aquifer increased two times when the leakage was increased from nearly zero to a value representative of thin and highly permeable confinement. Leakage affects the economics of in situ mining by (1) increasing the time required to leach out the ore, and (2) increasing the thickness of the ore zone aquifer to be restored. The potential for environmental impacts appears to be minor. In all four simulations, the lixiviant did not contact the confining strata. Restoration would be limited to cleanup of the ore zone aquifer within the thickness contacted by lixiviant and within the radius of the lixiviant that traveled farthest away from the production cell.

Jan 1, 1988

By Ronald J. Searle, Howard S. Latham

In April 1964 the Bureau of Reclamation awarded the first contract for rapid machine excavation of a water conveyance tunnel . Since then the Bureau has excavated seven tunnels with a total length of over 190,000 feet utilizing rapid excavation tunneling machines. Tunnel bored diameters have varied from 9&apos; 0" to 20&apos; 10" (See Table #1) and tunnel lengths from 5,000 feet to 67,000 feet. Driver and cutter units varied from 43 rolling teeth and eight rolling discs mounted on a rotating head driven by five 200 HP 23,000 volt electric motors to 20 discs and one tricone unit mounted on a rotating head driven by two 100 HP 440 volt motors. All tunnels were driven with a laser beam guidance system governing both line and grade. Trailing conveyors and rail haulage systems were used to dispose of the excavated material. Approved Bureau of Mines diesel locomotive engines for waste trains were prevalent power sources supplemented by battery units on some projects. Sandstone, conglomerates, agglomerates, siltstone, shale, and granite were the geological formations encountered, with sandstone and shale the preponderant material . Tunnel support systems utilized were structural steel ribs; rock bolts and metal plates; and shotcrete, or combinations of these. Progress varied from 51-1/2 feet to 154 feet per working day for three shifts, depending on rock material and supports required. One contractor established a record rate of advance of over 400 feet in one day, 1,905 feet in one week, and 6,848 feet in one month. Such diversified experience in underground machine excavation methods presented an excellent opportunity to observe, analyze, and evaluate associated environmental health problems. Also, to develop measures for controlling

Jan 1, 1997

By Giovanni Toniatti

Brazil is a country with continental dimensions. It hosts many different varieties of terranes and geological environments in its 8.5 million km2 (3.28 million sq miles). This mineral endowment has been important for the country&apos;s economic and social development process since the early colonial period, leaving Brazil with a legacy of almost 400 years of mining tradition. Mining in Brazil started in the early 17th century with diamond and gold discoveries. The made the country the leading producer of both for more than a century. This phase faded by the late 18th century when the alluvial deposits were depleted. It finally yielded to competing countries where technology and capital were being applied more efficiently. This period of mining dormancy lasted from the late 19th century until the first decades of the 20th century when Brazilian mining picked up. There was growing pressure on mineral resources by the affluence of the industrial society. The fast-growing demand resulting from the two world wars had brought the Brazilian mineral wealth back to the scene. This second Brazilian mineral boom made the country a leading producer of iron and manganese ore, quartz crystals and other strategic minerals. This expansion in Brazil&apos;s mining industry provided the basis for the country&apos;s industrialization process. This was led by the Brazilian steel and petroleum industries. Unfortunately, the thrust provided by this modern industrial basis was not properly used as a lever for further industrial development. The upstream industrialization process was not supported by the proper development of the raw materials supplying basis. The resulting industrial model was heavily dependent on imports of raw materials. It did not support the economic growth and faced fierce industrial competitiveness and sophistication that emerged with the postwar economic boom.

Jan 1, 1997

By L. A. Puchkov

Problems of how to use artificial intelligence (AI) methods and expert systems (ES) to control complex industrial projects have lately become significant in practical terms. Mining ventilation systems undoubtedly belong to complex multi dimensional objects of control with stochastic nonstationary and nonlinear processes. Formal models based on automated ventilation control, do require awkward algorithmic solutions, moreover, a number of tasks become the operator&apos;s headache, and the quality of their solution depends on his experience and intuition. These tasks include: - identifying pre-accidental situations and determining their causes; - making well-grounded decisions with regard to changing ventilation regimes, forecast of mining operations taken account of; - controlling ventilation systems in emergency situations, e.g. when in accident, etc. It is in this context that we have for a number of years been investigating how to practically use AI methods for operative control in dynamic aero-and-gas mining processes, one of our research directions being devoted to working out a consulting programme for the ventilation system operator. Our researching and modeling problem was identification of aero-and-gas situation in different parts of ventilation network and its interpretation in order to determine possible causes for gas regime changing and unexpected consequences. We would single out the following stages in this task: I. Operative information procession coming to the data base from the sensors and analysis of hypothetical versions (using the associative rules base) with regard to factors which have caused significant parameter fluctuations (e.g. C%(t), H(t), Q(t) etc.)., 2. Selection of hypotheses in the knowledge base and forming conclusions in accordance with reliability scale. 3. Clarifying the obtained conclusions in dialogue with the operator. 4. Correcting the associative rules base.

Jan 1, 1992

By A. Fiszer

One of the major problems with deep Polish hard coal mines are deformations of close-to-wall galleries since they are necessary and must he kept open for emergency situations and are largely responsible for work disturbances. In order to ensure safe and effective working operations, it is necessary to often rebuild galleries which, apart from the inevitable stoppage of operations, is greatly disadvantageous for the operation of a mine. The largest problems with maintaining the galleries open occur in Lower Silesian Coal Field mines where sometimes it is necessary to reconstruct the galleries several times [(Fig. 1)]. In light of what has been said above, the aim of this article is to present the reader with methods and ways of determining deformations occurring in close-to-wall galleries. Squeezing or convergence of headings is only a portion of the entire process of ground movement caused by ground disturbance due to workings. The term "squeezing of headings" is to be understood as a shortening of the distance between a point set on the roof of a heading and the corresponding one on the floor. It is very important to know the rules governing the process of squeezing in a particular mining system since it affords possibilities to define changes of the process and its parameters at any moment for a given system. This is of utmost importance because convergence of a heading is one of the most significant manifestations of ground pressure which has already been proved by studies performed by Everling (1). On the basis of those studies it has been observed that there exists an almost linear relation between the rate of convergence and loading [(Fig. 2)]. This, in turn, means that the deformations of the already exploited portion of heading are proportional to the pressure changes.

Jan 1, 1982

By G. E. Bandelis

In this paper, solid bed erosion processes and the mechanisms involved when a two-phase liquid-solid mixture flows over a bed of solids in a horizontal annulus are analyzed. There are few such studies available with some authors arguing that the flow rate for erosion is always less than the flow rate for the suspension of solid particles thus necessitating further study and analysis of such phenomena. Tests were run to determine the mechanism of erosion and the erosion velocities of the solid bed formed in a concentric annulus for three flu-ids, water, a low concentration water solution of Carboxyl-methyl-cellulose (CMC), which behaved as a Newtonian fluid and a higher concentration CMC water solution which behaved as a power law fluid. Glass spheres of 2 mm nominal diameter were used as solids. The flow rate at which erosion occurred together with the mechanism of bed erosion were determined visually, while the whole process was video recorded. Our tests indicated different erosion mechanisms and different erosion velocities for the three tested liquids, with the liquid viscosity playing a major role. The first mechanism, observed for water flow, gives erosion from the front of the bed while the second mechanism, observed for the flow of the two viscous fluids, gives erosion from the top of the bed. The role of each of the parameters to erosion is investigated, whether it is turbulent flow with the turbulent fluctuations being more capable to remove the solids from the formed bed or whether it is liquid viscosity, which results in the hydrodynamic shearing force. The results demonstrate that liquid viscosity plays a role in the erosion mechanism and yields different erosion velocities of the solid bed hence, it should be taken into account when modeling flow pattern transitions for solid-liquid flows. Prior studies have not identified this and it could be the reason of the disagreement between theoretical models and experimental and field data.

Jan 1, 2006

By A. G. Mayton, L. Adler

The paper under discussion provides an excellent and concise study of compartment ergonomics for thin seam equipment. However, in accomplishing this, some very important operational aspects were ignored. Since these are likely to be determinant, the authors&apos; response is solicited. The novel shuttle car with its cross-car, end-mounted compartment is operationally very different from the current one. The latter can be considered a highly mobile extension of the continuous miner or loader, with a pass-through conveyor, dump boom, and extra storage. As such, the vehicle&apos;s kinematics are the simple "I," or shuttle pattern, bent for going around pillars, with an intermediate stop at the "change-out point" (Adler, 1986 ). The action of the cable reel adjusts well to this pattern. The proposed shuttle car&apos;s motion either superimposes a "Y," time is lost (5 seconds) for each cycle, space is needed, line of sight is compromised, and a cable anchor is probably required at the dump. A chute is almost prohibitive in construction time and cost. The lack of a dump boom further curtails flexibility in the proposal. The value of a standard hand and opposite standard hand configuration also needs consideration. Many front-end loaders have a rocker (overshot) capability that is designed to achieve an "I” pattern. And crawler loaders sometimes use a side dumping bucket variant for the same purpose. These examples indicate the pattern&apos;s importance. For underground, thin seam application, the ram car has become popular. This is an articulated truck, battery-powered, and with an ejector dump. Its operational patterns would be similar to the novel shuttle car, except without the loading flexibility provided by a conveyor. Since the cab location of the proposed shuttle car largely obviates some of the major advantages of the present shuttle car, the question arises whether this configuration could be adapted to the current ram car and scoop

Jan 1, 1991

By George K. Martin

This paper deals primarily with roof bolting in Kentucky coal mines from January 1, 1949, through December 31, 1954, a six year period. The Consolidation Coal Co? (Ky.), a division of the Pittsburgh Consolidation Coal Co., began systemmatic roof bolting in November 1948 in Mine #204, Elkhorn #3 Seam, at Jenkins, Kentucky, After the efficiency of roof bolting had been proven by the Consolidation Coal Co., this procedure was adopted by other companies throughout the state. In December 1954, approximately 46.4 percent of Kentucky&apos;s total production of coal from underground mines was obtained from roof-bolted areas. Statistics in this report will show the progress made in roof bolting during the six years, in addition to the roof bolting being done as of December 31, 1955. Significant progress has been made in the coal mines of Kentucky in reducing the roof-fall fatality, rates, during the past several years by adopting systemmatic roof bolting. The ground work by the Consolidation Coal Co, (Ky.) in cooperation with the Federal Bureau of Mines was based upon the previous experience of the St. Joseph Lead. Cc, in its Southeastern Missouri Mines and The Tennessee Coal and Iron Division of the United States Steel Corporation. The author gratefully acknowledges the cooperation and assistance given by officials of the various mining firms using roof bolts, the Kentucky Department of Mines and Minerals, and the Federal Bureau of Mines in furnishing photographs and information.

Jan 1, 1958