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By William J. Miles, Raymond E. Blair

Introduction There has been much publicity of the boom and bust in oil and gas well drilling in the US, and many comments on oil and gas gluts. Despite this, the drilling industry remains the largest consumer of functional industrial minerals. The boom years of 1981 and 1982, compared with the bust year of 1983, show a sharp reduction in the consumption of industrial minerals-1981, 7.4Mt (8.2 million st); 1982, 6.3 Mt (7 million st); and 1983, 4.4 Mt (4.9 million st). Industrial mineral use rebounded in 1984 to 4.7 Mt (5.2 million st). This was the second highest year in history in terms of wells drilled and third highest in footage drilled. Table 1 shows the history of drilling activity and mineral consumption from 1971 to 1984. The uses of functional industrial minerals in oil and gas well drilling were detailed by Miles and Blair (ME, 1983, October, pp. 1397-1400). As an update, this article addresses only the changes that have occurred in the industry in the past two-and-a-half years. It will be directed to mineral use and new products introduced to the industry. Drilling practices A partial explanation of where wells are being drilled with lower quantities of minerals consumed is: • In the boom year of 1981, there was a proliferation of number of suppliers chasing the same business. This resulted in multiple sourcing, multiple warehousing, and a huge increase in inventory that became surplus in the following years. The bust year, 1983, caused a severe shakeout in the number of suppliers of minerals and associated drilling fluids. • In 1981 and 1982, the wild speculation in the oil and gas industry led to drilling of speculative deep gas wells. Oil wells were drilled on a wildcat basis - in areas of unproven productivity. • Presently, the oil and gas well drilling industry is very active. But drilling is directed toward proven areas of production and working over old wells for increased production. This means a large number of wells are being drilled with lower average foot-

Jan 11, 1985

By Edward F. Divers, J. Drew Potts

Powered dust-filtering helmets (dust helmets) have become increasingly prevalent in the underground mine environment in recent years. Many longwall operations use dust helmets to reduce the respirable dust exposures of miners required to work downwind of crusher, roof support and shearer operations. While continuous mining sections are usually less dusty, use of diesel equipment represents a new potential health problem. The U.S. Bureau of Mines recently conducted a study to determine the efficiency of a dust helmet on diesel-size particulate. Two underground evaluations were conducted on continuous mining sections that were using diesel haulage vehicles. Model No. AH21 Racal airstream helmets equipped with AS22-3 prefilters, AS23-3 main filters and AS60476 tyvek full-face seals were tested. Helmets were placed on mannequin heads and mounted in the ram car cabs. Andersen Model 298 (eight-stage) personal cascade impactors were used to sample air inside and outside of the helmet. Impactor stages 8 and the final filters were analyzed to determine submicrometer (0 - 0.93 µm) particle concentrations. Results from this study showed that the dust helmet was 94% (± 3% at 95% confidence level) efficient on submicrometer particles. Typically, very little coal dust is observed in the submicrometer size range, therefore, these particles were assumed to be from diesel exhaust. Results from this study indicate that proper use of the dust helmet on continuous mining sections that are using diesel haulage vehicles may be beneficial to the health of miners.

Jan 1, 1991

By Geoff C. Knight

Longwall mining using the shearer is acknowledged internationally as the most economically successful method of high volume production from underground coal mines. Since its introduction in the 1950s, a great deal has been achieved in developing the capability and reliability of the shearer - and the associated face conveyors and powered roof supports - to produce the competent systems now widely deployed. Equipment will continue to improve as knowledge and experience are gained and through the influence of market forces on the international equipment suppliers. However, current systems are so successful that further improvements are likely to be incremental rather than radical unless a high level of system automation is introduced. The automation concept is not new to the shearer-based longwall system. The opportunity and the need were recognized soon after the shearer and powered support were introduced in the 1950s. In the early 1960s, the British Coal Corp. (BCC; then the National Coal Board) attempted remotely operate longwall systems. Those imaginative early schemes were seen as failures at the time. But, with hindsight, a great deal was learned from the experience that stimulated automation research in several European countries. One of the most fundamental lessons learned was that coalface system automation is not possible until there is a high level of competence in automatically controlling the operation of the production machine itself within the coal seam. Once that is achieved, the rest of the system will follow with comparative ease.

Jan 1, 1991

By John C. LaScola, Gerald L. Finfinger, Michael A. Trevits

Methane gas is an ever-present problem facing the underground coal mining industry. Mine operators are mandated by law to maintain safe concentrations of methane gas within the mine environment. Ventilation is the method of choice used by all mine operators to reduce gas concentrations. When gas emissions are such that ventilation alone cannot control the gas, methane drainage methods must be applied. In the 1970’s, there was a concern that unprecedented gas emissions would be experienced as mine operators began new developments in the deep and gassy Blue Creek and Pocahontas No. 3 Coalbeds. Such concern was well founded as significant gas volumes are being emitted as these mining operations mature. In 1971, the U.S. Bureau of Mines began to tabulate and track the change in mine emissions. These data were extracted from the underground coal mine quarterly air quality-quantity inspections performed by the Mine Safety and Health Administration. At that time, approximately 6560 Mm3/d (231.7 MMcfd) of gas were emitted from 430 underground coal mines. It was determined however, that the mines emitting less than 2832 m3/d (100 Mcfd) contributed only 2.1 percent of the total daily gas flow. In 1988, underground coal mines emitted 8350 Mm3/d (294.9 MMcfd) from 487 mines. This report evaluates the changing nature of U.S. mine gas emissions over the last 18 years.

Jan 1, 1991

By Stephen E. Kesler

Carbonaceous sedimentary rocks are abundant in the upper part of the caldera or maar that hosts the Puebio Viejo Au-Ag district, and much of the Moore sulfide zone is in these rocks. These rocks are thinly layered and contain scattered plant remains and variable amounts of volcaniclastic debris and pyrite. The carbonaceous matter in the sediments yields no X-ray diffraction pattern and is not graphite. It has C/N ratios of 4 to 10, and &13C values of about -25 per mil, which indicate that it is dominantly of marine organic origin and partly degraded. Atomic S/C ratios in the unmineralized part of the carbonaceous rocks have a median of about 1.2, much higher than most similar unmineralized rocks, reflecting not only the abundance of pyrite in the sediments, but also the fact that the organic material contains several percent sulfur. Extracts from the organic material indicate that it is thermally mature and that it reacted with the hydrothermal solutions. Atomic S/C ratios in the mineralized carbonaceous rocks are generally higher than those in unmineralized carbonaceous rocks, ranging to values of 5 and more. These differences indicate that significant amounts of carbon were consumed by the reduction of sulfate and metal chloride complexes in the mineralizing fluids as they passed through the carbonaceous sediments. Direct reaction between organic material and ferrous chloride-bearing, acid-sulfate hydrothermal solutions probably caused continuous deposition of pyrite during this period. Intermittent deposition of gold from chloride complexes by the same reaction was probably facilitated by boiling or C02 effervescence.

Jan 1, 2013

By Dick Ruby

Most miners know that Latin America's major mineral areas are in the mountains of Mexico and down the backbone of the Andes in South America. These areas comprise some of the more rugged and challenging parts of the world's geography, and the risks of doing business in these areas are just as difficult as the impressive task of extracting and processing ores. However, Latin America does contain mineral deposits, some untouched and many to be discovered, that convince venturesome people that the tremendous effort required to mine them will be worthwhile. The rewards have often been adequate in the past. Hopefully, they will be in the future. Background The safest statement that can be made about Latin America is that the "ballgame is constantly changing." Here are some examples: Chile. Under President Frei in the 1960s, Chile welcomed foreign mining interests, even though the government was acquiring significant ownership interests in the big copper properties. Then, under Allende in the 1970s, the big mines were taken over and compensation was denied on various grounds including the allegation that extraordinary profits had been made that should have been taxed. After Allende, an about-face was made in policy; former owners of the expropriated mines received some compensation and new foreign investment was seriously welcomed. Many major companies announced intentions to return to Chile. Some have. Peru. After a military takeover in 1968, Southern Peru Copper

Jan 7, 1983

By Rolf Lange

Introduction The Harzwasserwerke, a public body of the Federal State of Lower Saxony, supplies drinking water to an area bordered by Braunschweig, Bremen, the Northern and the Western edges of the Harz-Mountains. From the very beginning in 1928, the Harzwasserwerke have been systematically constructing reservoirs in the western Harz which have the function of collecting and storing the frequently occuring flood waters as well as providing drinking water. The Grane-Reservoir with a capacity of 45 Million cu.m., the last of the big reservoirs, was built in 1968-1970. A system of connecting tunnels effectively enlarges the natural catchment area of this reservoir and the surplus waters of the Oker and Radau valleys - where floods occur particularly during the spring - are being conducted to the reservoir. By building the 4.800 m long "Radau-Tunnel", the waters of the Radau valley will now also be used more efficiently and its surplus waters will pass to the Grane Reservoir via this new tunnel and the Oker-Grane-Tunnel which was finished in 1970. The Project The Radaustollen has been planned and put out to tender by the Harzwasserwerke as a conventional "Freispiegelstollen". The horseshoe profile has a cross-section of 8 m2. Lining was required only in disturbed rock areas; lining of the whole tunnel was not necessary. The gradient of the tunnel is 4 % 0 rising = 25 m over the total length of 4.800 m.

Jan 1, 1979

By Usko Anttikoski

This article deals with the extensive geotechnical soil and bedrock mapping system maintained in the Geotechnical Office of the City of Helsinki since 1955.Soil maps have been prepared to scales of 1:500,1:2000 and 1:10 000, and most recently in 1978 a new type of bedrock map was printed to a scale of 1:10 000, which covers the whole city area and particularly serves rock tunnel construction. A map showing the types of foundation building in the central city area, which will also serve underground construction activities, is currently being prepared. Since the geotechnical mapping system has been well received in Helsinki and the drawing of similar maps has also been undertaken in other Finnish towns, this mapping method is believed to be of interest to foundation engineers and tunnel constructors elsewhere too. THE GEOTECHNICAL CONDITIONS IN HELSINKI The geological structure of Helsinki, and of all Finland, presents two types clearly differing in character and age; bedrock and soil. The solid bedrock is Precambrian, while the soil dates from the Pleistocene Age and has mainly been deposited during and after the Ice Age. Thus the history of Helsinki only displays very old and very young formations, but nothing in between. The ground level in Helsinki is +62 at its highest and the sea bottom close to shore is -15 at its deepest. The deepest rockhead has been found to lie at the -70 level.

Jan 1, 1979

By Hu Wenjun, Li Maonan

Mine ventilation systems mostly are ventilation networks with fixed-quantity branches. The solution to the problem of ventilation networks with fixed-quantity branches is obtained by choosing the fixed quantity branches as chords and using Hardy Cross iteration method. The solution usually is feasible but may be not economical. This paper takes minimal annual ventilation electricity cost as an object, considering the power of main fans and boosters, natural ventilation pressures and resistances of regulators, and builds up a nonlinear programming model. A set of initial values of the nonlinear programming can be obtained by using Hardy Cross method, The MAP technique then used to find a set of local optimal solution The running procedure with a simple example is dicussed. Mine ventilation systems mostly are ventilation networks with fixed quantity branches. In the networks, air quantity of branches are generaly regulated by increasing branch resistance or air pressure. A comon approch to the of the networks with fixed quantity branches is to formulate it based on Kirchhoff's current and voltage laws and the solution is obtained by Hardy Cross iterative technique. The solution usualy is feasible but may be not economical. The optimal plan of air quntities regulation of mine ventilation network should be to make the solution satisfy both required quantity of work places and minimal ventilation operating cost. It is the purpose of this paper to introduce a new approch and computational experience of nonlinear programming of the optimal problem.

Jan 1, 1991

By William N. Poundstone

Consolidation Coal Company, through the mining research division of Continental Oil Company, is now undertaking the installation of a full-scale hydraulic transportation system to carry coarse coal from the mine face to the preparation plant. The facility will include all elements of a fully integrated haulage system including face haulage for both longwall and continuous miner sections, multiple slurry feeds, hydraulic hoisting and 2.4 miles of overland slurry transport. This undertaking is a result of more than six years of extensive research and testing, some of which has been previously releasted to the literature. To provide a reference framework for system description, a brief history of recent coal mining technological developments will be discussed. Following World War II, mechanization resulted in mining machines that were bigger, faster, more powerful, and more reliable. These changes were essentially refinements in an established mining system. They made mining more efficient than before, but they did not change the system itself. By about the mid-1960's, it was apparent that further improvements in the extraction machinery would have a declining payoff, simply because .the worst bottlenecks were in other components of the mining system. A modern continuous miner cannot be run continuously because other operations underground cannot keep pace. In fact, most continuous miners today run at less than 20% of capacity. Consequently, in 1969 Consol and its parent company, Continental Oil, undertook a long-range R&D program to try to change the basic haulage system for the better. Most coal at present is removed from the mine by a hand-off system in which the coal is transferred from the continuous miner to a shuttle car, often with the aid of a separate loader; then from the shuttle car to a system of conveyor belts; and finally from the conveyors to rail cars. The system involves a large amount of moving equipment underground, many transfer points, and the everpresent chance of operational delays and safety hazards. By removing the haulage bottleneck, we believe we can speed up underground operations substantially; and perhaps more important, we could make it feasible to develop a whole new generation of mining equipment that is inherently safer. What we wanted was a truly flexible, continuous haulage system with inherent safety advantages. Ideally the system had to meet three major requirements: First, the coal had to enter the transport system at the face, and remain in that same system all the way to the preparation plant. Second, the system had to be flexible and extensible so that it could- follow the mining machine as it advances and retracts. And, third, the system must be inherently safer than current haulage technology.

Jan 1, 1977

By K. Hansson, A. Jedborn

Background In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore, from a system's point of view, slurry transporta¬tion would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alterna¬tive examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasi¬biliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. l . The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then tempo¬rarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situ¬ations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe¬line, and personnel are needed. In a first phase (Fig. 2), a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1989

By Anders Sellgren, K. Hansson, A. Jedborn

In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore. from a system's point of view, slurry transportation would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alternative examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasibiliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. 1. The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then temporarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe- line, and personnel are needed. In a first phase (Fig. 2). a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1990

By K. A. Brunk

This paper describes Newmont's research, operations, history and experience with pre-treatment of preg robbing carbonaceous ores. The chemistry, physics and kinetics of chlorine treatment are described. Present operations, control strategy and projected changes are discussed. Included is a summary of the adsorption of chlorine into pulps and off-gas scrubbing.

Jan 1, 1987

By F. C. Bond

History The breaking and shaping of rock was one of man's earliest occupations. In the Paleolithic Age long before the dawn of history, arrow¬smiths and the makers of stone axes, hammers, knives, scrapers, spears, and borers were highly respected members of society. In early historical times stones for building blocks, roads, and city walls were shaped by slaves and convicts, who also did most of the mining. However, great artists erected beautiful stone sculptures, while gifted architects planned imposing temples and monuments. Until well into the 19th century nearly all rock was broken laboriously by hand. The small rock required by John MacAdam for his macadamized roads in England in the 1820s was produced by women and boys seated alongside the roadside with hand hammers and legs wrapped with rags. Eli Whitney Blake, a nephew of the Eli Whitney who invented the cotton gin, developed the first successful jaw crusher before 1870. The gyratory conical crusher soon followed. Comparative tests established its large capacity advantage over the jaw crusher, as well as its greater cost for a given feed size. Both types have been in use for more than 100 years. Crushing rolls appeared before 1900. Thomas A. Edison made very large diameter rolls which were excessively long; they failed because of shaft deflection. Various types of disk crushers and edge runner rolls appeared about this time. The older methods of reducing rock were adaptations of other processes. The stamp battery of dropping weights effected crushing by simulating heavy hammer blows; the much earlier arrastre, in which heavy stones were dragged in a circular path over the ore by animal power, came from the prehistoric method of grinding grain between two rubbing stones, while the jaw crusher was adapted from simple squeezing devices. But the tumbling grinding mill was not just an adaptation; it was an invention, because it required thinking on a somewhat higher order-there was no prototype. Its nearest antecedents were probably the small closed tumbling drums used in England more than a century ago for cleaning and polishing small iron castings. The date of the first tumbling mill actually used to grind rock is unknown, but it was later than the American Civil War (1861¬1865). It was almost certainly a closed or batch mill in which rock was placed and rotated until it reached the desired particle size. It could have been operated either wet or dry. The first published refer¬ence to such a batch mill was one introduced by Alsing in England (1870) for the grinding of calcined flints for pottery work.21 There are several rather indefinite reports of grinding mills in the early 1890s, including an overflow ball mill in the Helena and Livingstone reduction plant in Montana which may have been the first of its kind. 11 Many of the first mills, which were called tube mills, used hard rock both as grinding media and as mill lining. The rock used was preferably stone from the Normandy and Danish beaches, when it could be imported. This remarkable siliceous stone was already widely used for grist mills throughout America, and its resistance to wear was greatly respected. The decade of the 1890s saw the development of tumbling mills with continuous feed and discharge and their extension into different industrial uses. By 1895 some experience had been accumulated. Iron grinding balls were being tested and the proper speed of rotation was being determined. The Clark Patent tube mill was featured in an E. P. Allis bulletin of 1890, which may have been the first published description of a tube mill. More than 1,000 Gates tube mills had been built by Allis-Chalmers before 1913. Many of these were used in gold mining, espe¬cially in South Africa. The 5 x 22-ft size was particularly favored for grinding portland cement; the use of tumbling mills in the manufac¬ture of cement began about 1900. A great deal of attention was paid to the mill lining. Metal was then relatively expensive, and the general approach was to trap some of the rock grinding media into mill lining pockets. This rock would then absorb the wear and protect the metal lining. In the first ten years of the 20th century there were several different types of pocketed liners, with different manufacturers advancing the superior claims of their patented arrangements. The Osborne liner, developed in South Africa, was probably the most successful. 21 Another item which attracted much attention was size classifica¬tion within the mill and in ancillary equipment attached to the mill and rotating with it. The Krupp type, with interior screens protected inside the mill lining, was developed quite early in Germany, possibly before 1890. The Dorr reciprocating rake classifier (1907) had not yet been invented, and many strange and impractical screening and classifying devices were proposed. In these unsatisfactory machines the two separate processes of size reduction and classification were combined into one operation. It was many years before recognition came that a machine is most efficient when it is designed for one specific purpose. There was much industrial wastage before the opera¬tions of grinding and classification were finally separated. After 1900 the grinding of portland cement raw material and of cement clinker required large numbers of tumbling mills. Most of the raw material was then ground dry. This was also the heyday of gold mining. The old stamp mills that were used in great numbers for grinding gold ore did not grind sufficiently fine to liberate all of the gold, and the new tube mills were installed following the stamps. After 1910 larger diameter tum¬bling mills with larger grinding media were developed. These could receive the finely crushed ore directly, and the inefficient stamp batter¬ies were gradually eliminated. The Rand in South Africa was the greatest gold producer, treating immense quantities of rather low-grade but consistently free milling gold ore. The first tumbling mills, or tube mills, went into operation there in 1904. They were so successful that within a few years no new stamp batteries were installed in the district, even though old ones continued to pound away until after 1950.3 The early tube mills on the Rand all employed Normandy or Danish pebbles, which had to be imported at considerable expense. Their reported wear was as low as 4 lb per ton ground. 4 Many of the mills were lined with the same tough Danish stone cemented into place, while others used the pocketed steel Osborne liner. It was in 1907 on the Rand that an important test was made using hard native ore for grinding media in place of the expensive imported pebbles. 3 This ore, called banket, did not wear as well as the renowned Danish pebbles, but the cost per ton of grinding was definitely reduced. Many of the tube mills on the Rand were soon grinding with native ore. This was the beginning of the development now called autogenous grinding, in which the ore grinds itself. This is treated under a separate heading. See Subsection 3C, Chapter 4. Gold mining was important in America also, and its grinding history follows that of the Rand. Danish pebbles were replaced by native ore in Santa Gertrudis, Mexico, in 1913, 4 and in Consolidated Gold Fields, Nevada, in 1914. 5 Other properties followed suit. However, it gradually became apparent that the capacity of a given mill could be almost doubled if rock grinding media were re¬placed by cast iron or steel grinding balls. In order to increase plant grinding capacity many rock media mills were converted to iron grind¬ing media in the second decade of the present century. In some mills the motor size was doubled; other mills were cut in two and another motor was provided for the second half. Grinding mills began to assume a more modern appearance. Crushing rolls were formerly much used following crushing in jaw or gyratory crushers and preceding grinding in ball mills. How¬ever, the roll surfaces wore rapidly, and skilled maintenance was required to obtain even wear. Rod mills could take feed of the same crusher product size and reduce it finer. The first rod mill was con¬structed by Mine and Smelter Supply Co. and was tested in Canada

Jan 1, 1985

By F. W. Brackebusch

In the paper presented by F.W. Brackebusch on the "Basics of paste backfill systems," paste fill was presented as a backfill process that offers many advantages over other alternatives such as hydraulic fill and rock fill. It was observed, however, that the mechanics of paste-fill transportation in pipelines is not fully understood and that the success of its implementation in practice depends on the precise control of the proportions of fines, aggregates and water in the mixture. The author emphasized the necessity for laboratory and full-scale testing to determine the suitability of a material to be used as paste fill. Although experimental testing is definitely an important operation in the design of a reliable paste fill, advances in the understanding of the fundamental laws that govern paste flow in pipelines are equally relevant. Ideally, the results of such efforts would provide predictive models for the rheology of the paste and the corresponding pumping-energy requirements. In an attempt to achieve this objective, an ongoing research program has been active at McGill University's Department of Mining and Metallurgical Engineering. Preliminary results showed that the underlying mechanics of paste-fill flow in pipelines may be described in terms of what is known as "plug flow," as depicted in Fig. 1. Although this flow mechanism has been known qualitatively for many years, few attempts managed to offer a comprehensive model (Gandhi, 1987; Duckworth, 1986). In our model, it is assumed that coarse aggregates present in the mixture form a porous core, which is filled with paste (fines and water) and is surrounded by an annular non-Newtonian (Bingham-plastic) lubrication layer, which enables the mixture to flow with pressure losses that are easily overcome by common positive-displacement pumps. In summary, our model predicts the thickness of the annular lubricating layer, which controls the magnitude of the pressure losses. This is done either by collecting data on the apparent viscosity of the Theologically-active fine fraction and using an analytical equation (Bouzaiene, 1994), or is done by measuring the void content of the coarse aggregates and deducing the thickness as a function of the fine- to coarse (weight or volume) fraction ratio in the mixture for a given pipe diameter and solids concentration. Our model predictions showed good agreement with data from Duckworth et al. (1986), as shown in Fig. 2.? References Bouzaiene, R., 1994, "Hydraulic transportation in mining," Ph.D. dissertation in progress, McGill University, Dept. of Mining and Metallurgical Engineering, Montreal, P0, Canada Duckworth, A., Addie, G.R., and Maffei, J.R.,1986, "Mine waste disposal by pipeline using a fine slurry carrier," Proceedings of the 11th International Conference on Slurry Transportation, Slurry Transport Association, Washington DC. Gandhi, R.L., 1987, "Effect of rheological properties of fines fraction on slurry hydraulics," Proceedings of the 12th International Conference on Slurry Technology, New Orleans, LA.

Jan 1, 1996

By P. Worley

With Utah Power & Light Mining Division's commitment to operate the safest and most productive mines possible, a decision was made to purchase and install mine-wide monitoring systems for use in the mines. After review of several different systems the Conspec Senturion 200 system was chosen. Three separate systems were ordered to fulfill the monitoring requirements of the six mines in operation: one system for the Deer Creek Mine one for the Wilberg/Cottonwood complex, and one for the Deseret, Beehive and Little Dove complex.

Jan 1, 1989

By Kelly F. Bailey

The issue of regulating non-asbestiform varieties of the amphiboles actinolite, tremo1ite, and anthophyllite to the same degree as asbestos has been studied in detail by the National Stone Association. The ramifications of such a regulation, and the over simplistic regulatory definition of a fiber, are enormous not only for the aggregate industry and its users, but also for the nation. Evaluation of the health, geological, mineralogical, technical, economic and legal issues do not support an "asbestos-like" regulation of non-asbestiform minerals. The adverse consequences of regulating non-asbestiform varieties of actinolite, tremolite and anthophyllite like asbestos would be magnified many times if a similar standard for generic mineral fibers were promulgated.

Jan 1, 1988

By J. Pereyra, M. DeMarco

"INTRODUCTION It can be vital that bolted joints holding subassemblies together remain secure. Bolts should resist the loosening effects of vibrations and dynamic loads yet need to be easily removable during maintenance. When a threaded fastener is subjected to vibration, the rapid movement causes a lowering of friction against the threads and a subsequent loss of preload. The loss of preload allows the fastener to vibrate loose and possibly fall out. As the diameter of fasteners increase desired preload becomes increasingly more difficult to achieve. Input torque requirements for achieving proper bolt preload is exponentially related to any given fasteners diameter. This means that specialized tooling or bolting practices are usually required to attempt the tightening of fasteners to reach their required preload. Traditional methods can cause problems with joint integrity and do not provide consistent bolt load across the bolted surface. BOLTED CONNECTIONS A failed critical bolted joint can lead to high maintenance costs, machine downtime, and risk to worker safety. Control of the clamp load in a bolted joint is vital. However, when faced with a problem joint, it is not surprising that the design engineer will not have an answer if asked about the clamp load. Torque calculations must always be based on the existing conditions that often are vague. Unless all parameters are correct, the calculation will be unreliable. Examples of parameters are: • Thread condition of the fasteners • Hardness of contact surface • Material (steel, aluminium, copper, etc.) • Extra friction from a locking fastener • Extra friction from an adhesive • Lubricant on the thread • Type of bolt head (flanged, regular or serrated) • Surface coating of the bolt • New or reused fastener During tightening, bolts are subjected to both tensile and torsional stress. The total stress in a bolt can be calculated using the formula: Total stress = vsx2 + 3txy2 In order to maximize the desired tensile stress (sx) it is vital to minimize the torsional stress (txy). Tensile stress (clamp load) is achieved when the bolt is axially elongated. Unwanted torsional stress (twisting) in bolts arises during tightening due to thread friction. High thread friction increases torsional stress and causes yielding at lower clamp load levels than normal. A lubricant is necessary to minimize the malign torsional stress. However, many commonly used bolt-locking systems are based on increased thread friction (deformed nuts, adhesives, etc). To minimize thread friction and concurrently safely secure the joint has often incorrectly been considered impossible. The use of locking systems that increase thread friction is the single most common reason why the full capacity of bolted joints is not utilized. The following example illustrates the problem."

Jan 1, 2015

By R. H. Murray

The Kamyr countercurrent tower leach process consists of a true countercurrent leach ?with a flocculated ore and fibre mixture entering the top of the tower and a cyanide solution or alternative lixiviant as well as a barren wash solution entering the bottom of the tower. It offers the possibility of including the entire leaching, solid-liquid separation and clarification processes into one unit or tower as opposed to the multiple leach tanks, filtration and clarification sections required in a conventional leach and precipitation plant. The tower process has been demonstrated in the laboratory and found to be effective for leaching gold, silver and other amenable ores. Based on laboratory results, Gencor and Kamyr have jointly constructed a 200 ton per day pilot plant in order to prove the process on a larger semi-production plant scale.

Jan 1, 1989

By Carrie Smith

The 2025 Women in Mining USA National Conference, held April 15-18 at the Grand Hyatt Riverwalk in San Antonio, TX, brought together hundreds of professionals for a fourday event focused on empowering women and creating space for all individuals in the industry. Under the theme “Empowering the Future of Mining,” the conference served as a dynamic platform for professional development, knowledge-sharing and cross-sector collaboration. Organized by Women in Mining USA (WIM USA), whose mission is to foster equity, sustainability and leadership in mining, the event featured an extensive agenda including workshops, keynote speeches, panel discussions and networking opportunities. Attendees ranged from industry veterans to emerging professionals and students, united by a shared commitment to progress within one of the world’s most essential and traditionally maledominated sectors.

Jun 1, 2025