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By Jurgen H. Kleinau

This paper details Keeler/ Dorr-Oliver's experience with the application of fluid bed combustion technology to the burning of coal-mining waste. The designs of two stage fluid bed combustors/dryers are demonstrated as useful in the drying of coal, slag and coke, using coal and coal refuse (gob) as fuel. Anthracite mining refuse (culm) is more than abundant in Northeastern Pennsylvania. After demonstration at Shamokin, Pennsylvania, a full commercial-sized fluid bed boiler using culm is used for district heating in Wilkes-Barre, Pennsylvania. Limited research work has shown the utility of using fine coal as filter aid in sludge in¬cineration. With the rising avenues of the suitability of coal as auxiliary fuel in fluid bed sludge incineration, an expansion of these concepts combines the use of coal or coal refuse as filter aid and auxiliary fuel. Lime¬stone addition controls SO2 emission.

Jan 1, 1985

By Matthew Holden, Jason Emery, Matthew Benjamin, Peter Craig

Over the last 25 years, through adversity and a growing level of risk mitigation, Australian Longwall mines have been using higher and higher capacity cable bolts for secondary support. A recent survey revealed 85% of Australian longwall cable bolts are a 70 tonne capacity pre-tensioned hollow post groutable system. Mines avoiding pumping cementitious grouts are sometimes using pre-tensioned resin encapsulated solid cable bolts of over 90 tonne capacity. Results from various industry and University research has led to manufacturer's fine tuning the construction of their cable bolts. A review of this research explains how smooth wire construction is now dominant. Aggressive shear has also led to some mines using 8 tonnes pre-tension instead of the previous common 20 tonnes pre-tension. Lower pre-tension acceptance is now leading to a greater use of threaded cable bolt end fittings, these are tensioned to 8 tonne using the drill motor torque. A recent development is the TEX cable bolt, a slim line threaded end fitting that fits with the drill hole like a standard rock bolt. This cable is being installed at the coal face as primary support off bolter-miners. The laboratory and in-situ testing of the TEX cable is presented. In a bid to improve gateroad development rates and ground support in the last several years, a 100 tonne tensile capacity Goliath cable system was developed. The latest development in resin anchored solid cables is a 110 tonne capacity cable bolt, with laboratory and in-situ testing results presented.

Jun 25, 2024

By M. Churchill, V. R. Degner

Over the past two decades, declining ore body grades and increased operating and construction costs have provided an incentive for the development of progressively larger flotation machines. As a result, the 1.7-2.8 m3 (60-100 cu ft) flotation machine of the 1950's and 1960's is being rapidly replaced by large volume machines. The result has been improved flotation machine specific capacity (tph processed per sq ft of floor space); specific power (hp per tph processed); and process control simplicity. The recent development and satisfactory plant evaluation of the 28 m3 (1000 cu ft) flotation machine allows economic and operating advantages in high tonnage mill applications. Some principles of mechanically induced air flotation and the development of the large 28.3 m3 (1000 cu ft) flotation machine are reviewed here. Performance evaluation in beneficiating metalliferous ores-copper, zinc, and iron-is also given and recently completed plant results provided. Mining lower grade ore bodies results in the need to process increased tonnage in the concentrator. This greater tonnage requires more individual flotation machines, unless individual machine capacity, or cell volume, is increased. As the ore processing rate of the concentrator increases, the benefits of progressively larger individual cells in the flotation section become more significant. The accompanying figure illustrates the advantages of a larger flotation machine. In the figure, the number of machines in the rougher-scavenger section is related to the plant tonnage for three sizes of flotation machine: 8.5 m3 (300 cu ft), 14.2 m3 (500 cu ft), 28.3 m3 (1000 cu ft). The 28.3 m3 (1000 cu ft) flotation machine results in 95 fewer machines, compared to the 8.5 m3 (300 cu ft) machine at 45.4 kt/d (50,000 stpd) feed. The flotation cell reduction is, of course, doubled at a plant production rate of 90.7 kt/d (100,000 stpd). Using fewer flotation machines saves floor space. The 14.2 m3 (500 cu ft) and the 28.3 m3 (1000 cu ft) machines result in a floor area reduction, relative to the 8.5 (300 cu ft) machine requirements, of 134.7 and 418.1 m2 (1450 and 4500 sq ft), respectively, for the 45.4 kt/d (50,000 stpd) rate and 269.4 and 836.1 m2 (2900 and 9000 sq ft), respectively, for the 90.7 kt/d (100,000 stpd) rate. In addition, flotation practice has shown that the "specific power intensity"-power per unit cell volume-is less for increased cell size. This results in a power draw reduction, for the rougher-scavenger circuit, of between 552 and 686 kW (740 and 920 hp) for the 14.2 and 28.3 m3 (500 and 1000 cu ft) machines, respectively, when processing 90.7 kt/d (100,000 stpd). Fewer cells improve machine control and simplify maintenance needs while reducing floor space and power requirements. These are tangible incentives for the development of large flotation machines for high tonnage applications. Flotation Principles The accompanying figure shows how key hydrodynamic regions influence the performance of a mechanically induced air flotation machine. Air is induced into the fluid from the top of the cell, due to the mechanical action of a rotating impeller. The impeller circulates the liquid, or pulp, from the bottom of the cell. The result of these dual flow paths is the mixing of the two fluids and one solid phase in the highly agitated region where physical contact between the air bubbles, and the solid particles to be floated, is accomplished. The particle to be floated is assumed to have a properly prepared hydrophobic surface prior to flotation. Upon leaving the three-phase mixing region, the "floatable particle-air bubble matrix" and

Jan 7, 1982

By J. Weinstein

New erosion, corrosion resistant metal, ceramic, and polymer composite technologies are being developed and marketed by Alanx Products L.P. for use in hydrocyclones, pumps, valves, and other applications in mining and minerals processing. This paper will present an overview of the ceramic/metal composite material system and properties developed, which include excellent erosion/ corrosion resistance. Wear data will be shown from a variety of laboratory tests in dry and wet conditions, in aqueous neutral and corrosive environments. Comparisons will be made with a variety of competitive materials currently used in the industries sited. Data will also be presented on field tests with pumps and hydrocyclones in hard rock mining and other abrasive/corrosive environments. Field data from North and South America, as well as Australian mining applications will be highlighted.

Jan 1, 1992

By Nathaniel Arbiter

The increasing size of flotation cells as the use of flotation has increased is traced from the first United States application of the process in Montana in 1911 through the following eight decades, using data from the Taggart handbooks (1927) and (1945), from the SME Handbook (1984), and from recent manufacturers' catalogues and other sources for the 1980s and 1990s. With the earliest mechanical cells less than 1 m3 (359) in volume and the largest now available at 200 m3 (7,000ft3), three requirements are necessary: 1) demonstration of advantages in replacing smaller volume by larger volume cells; 2) estimating power requirements, impeller dimensions, and operating rpm by correlation equations; 3) considering effects of using larger cells, in necessarily much shorter rows, on process kinetics. The resulting seven tables that address these problems are presented, and their implications discussed.

Jan 1, 1999

By Paul Wellman

The energy problem of the U. S; was emphasized by the oil embargo of late 1973 and early 1974 and it seemed for a time that the Federal Government, under pressure from the general public, might take some corrective action. However, this was not the case. With the gasoline lines gone and an apparent plentiful supply of petroleum products, public concern quickly dissipated., This country is more vulnerable to an oil embargo today than we were in 1973-1974 and this fact should be of concern to all our people. Since 1968, we have used oil and natural gas at a much more rapid rate than we have been able to find new domestic sources; thus, there is an increasing gap between domestic production and consumption and we have filled the gap with ever increasing oil' imports. These imports have increased at an alarming rate in the recent past, some 58 percent in 1976 compared with 1975. Imports are nearing 8 million barrels per day. Then, by applying some simple arithmetic, at $12 per barrel, the resulting trade deficit is $96 million per day or $35 billion per year. Obviously, this cannot be tolerated by our economy over a long period of time. Historically, we have shifted from one fuel to another as is shown in this slide. In 1850 about 90 percent of our fuel requirements came from direct combustion of wood, by 1900 we shifted to coal for 70 percent of our fuel, and now about 78 percent of fuel requirements comes from oil and natural gas. The difference is that the previous shifts were voluntary and the one facing us now is of necessity.

Jan 1, 1977

By Nagashi Ogushi

Geochemical prospecting method using stream sediments, soils and rock chips has been tested to see the applicability for the typical Kuroko field in Japan. Three independent anomalies have been detected in maps being shown by grav densities of cells which are represented by the results of factor analysis as well as distribution of minor elements scaled in logarithm. A small vein type mineralization which has outcrops could be attributed to the origin of them whereas the one near the surface of known Kuroko deposit to a coincidence. The addition of drill core geochemistry is required to complete the interpretation.

Jan 1, 1982

By P. Somasundaran, S. Venkatachalam, N. K. Khosla, K. Khosla

Froth flotation of alumina particles was studied by employing and hydrogen gas bubbles. The effect of current density was studied over a range of 0.1 to1.6 kA/m2. Compared to hydrogen, the use of oxygen bubbles at any given gas evolution rate resulted in a higher flotation recovery. The existence of an optimum current density, for each given particle size was indicated. Flotation at higher current densities resulted in better flotation kinetics, while flotation at lower current densities resulted in better Spas utilization.

Jan 1, 1996

By Kenneth W. Torpey, James G. Ball, Richard F. Flanagan, David S. McAllister

INTRODUCTION Under a Federal grant by the Urban Mass Transit Administration (UMTA) (80%), and sponsored by the State of New York (200), Niagara Frontier Transportation Authority is constructing a 10.3 km (6.4 mile) light rail rapid transit system in Buffalo, New York. The $450 million project, which began construction in the Spring of 1979, is unique in rail transit alignment designs in the U.S.A. and the first predominantly underground light rail system in North America. The Buffalo project is unique in that the first phase of the expandable system follows the alignment of Main Street and is on the surface in a proposed 1.6 km (1.0 mile) pedestrian mall in the downtown business district. The alignment then progresses northward into a cut-and-cover portal and standard double-box configuration for 2.74 km (1.7 miles) at a depth of 9.1-10.7 m (30-35 ft) and subsequently drops into bedrock at depths of 18.3-27.4 m (60 to 90 ft). The 5.63 km (3.5 mile) rock tunnel section consists of two single tunnels 4.88 m (16.0 ft) nominal finished diam. and includes five stations. The rock tunnel line section was divided into a north and south contract of 2 316 m (7,600 ft) and 3 243 m (10,640 ft), respectively. Due to the existing surface facilities and the future station locations, the line tunnel contracts differed considerably in their shaft access

Jan 1, 1981

By L. A. Pursell

The case history, which is the subject of this paper, involves the expansion of a lime plant in Texas. The major new production unit is to be a 545.45 tonne per day coal fired rotary kiln. Auxiliary facilities include additional quarry haulage equipment, modifications to the crushing plant, a hydrator, storage facilities, and a loading system. The existing plant consists of a high calcium limestone quarry, two coal fired rotary and 10 gas fired shaft kilns, nth crushing, screening, stockpiling, hydrating, storing and loading facilities. The proposed modification would increase quarry volume from 545,000 tonnes per year to 805,000 tonnes per year. The plan also includes the addition of dolomitic lime to the product line which requires the purchase and handling of 186,000 tonnes of screened, washed dolomite yearly. A decision to undertake such an expansion was previously a simple and straight forward process. Such a decision now faces a major and complicating obstacle which can delay the project, raise the expected cost, or prevent it entirely.

Jan 1, 1980

By Egons R. Podnieks, Wallace W. Roepke

Mining technology to utilize vital lunar resources will be required for any commitment to permanent lunar basing. Lunar mining must be viewed both as a major supporting element for space resource commercialization and for creation of human settlements on lunar and planetary surfaces to support long-range space exploration. The Apollo missions to the lunar surface and results from fundamental research on analogue lunar rocks by the Bureau of Mines, and others, have provided the initial data for developing lunar mining technology. The multi-disciplinary research projected in this paper outlines the scientific and engineering needs to realize the lunar mining goals.

Jan 1, 1990

By David Court, Rab Brown

The paper discusses the work in progress on the construction of the extension of the High Speed Rail line from the Channel Tunnel into the centre of London UK known as the “Channel Tunnel Rail Link” (CTRL), and particularly on contract 250 which involves the construction of two 5.3Km tunnels 7.15m diameter constructed with two EPB convertible tunnelling machines underneath the densely populated area of East London. In addition the paper will look at an innovative Alliance that has been formed between the Contractors and the Client’s Engineer involved in the construction of the London Tunnel Contracts

Jan 1, 2003

By R. Gong, R. Bhaskar

[Full-scale laboratory experiments were performed to evaluate the effect of ventilation and water spray pressure on the potential dust capture efficiency of a ventilation system comprised of blowing ventilation and an underboon scrubber. The tests were performed for a 1.52-m (5-ft) roof height and 2.36 m-3/s (5000 cfm) blowing ventilation volume, and a 2.44-m (8-ft) roof height and 4.72 m3/s (10.000 cfm) blowing ventilation volume at three face positions of the miner. The scrubber suction rate was varied for the tests to study the effect of scrubber flow rate on capture efficiency. The airflow patterns and scrubber efficiency were determined by performing tracer gas tests. Tracer gas was released through a manifold placed in front of the rotating miner drum. The gas was sampled in a grid fashion using a real-time multi-point gas sampling system. The capture efficiency of the scrubber was evaluated based an the peak gas concentration of the gas laden air drawn through the scrubber and the amount of gas passing, through the scrubber as a fraction of the gas released at the face. The data were analyzed to yield scrubber peak concentrations, potential exposures at various operator locations and the gas removal percentage (GRP). The underboom scrubber appears to work best in the end sump position, relative to the start sump and start slab positions. The capture efficiency scrubber was higher for 2.44-m (8-ft) roof heights compared to 1.52-m (5-ft) roof heights. The data also showed that the scrubber air discharged behind the mining machine may be drawn into the face area by the action of the blowing ventilation system, resulting in recirculation.]

Jan 1, 1994

By Thomas Jancic

Two of the open-pit, copper-molybdenum mining operations of Duval Corporation are located in the foothills of the Sierrita Mountains approximately 30 miles southwest of Tucson, Arizona. In March 1959, mining from a moderate sized porphyry deposit known as the "Esperanza orebody" was begun at a rate of 10,000 tons of ore per day. Some five years later, a sizeable group of mineral claims west of Esperanza became available to Duval. Earlier exploration and development drilling in this area by others indicated an orebody of greater size but lower tenor than that of Esperanza. A drilling program conducted by Duval not only verified the results of the prior drilling but in addition, revealed a continuation of the mineralization to the south and east. Thus was developed the Sierrita orebody. Open pit design delineated 414 million tons of ore containing an average copper content of 0.35 percent and a molybdenum content of 0.036 percent. In April 1966, the office of Emergency Planning authorized General Services Administration to undertake a program under the Defense Production Act to en¬courage additional domestic production of copper in the interest of national defense. Duval's Sierrita Project was brought to GSA's attention in May 1966 on November 28, 1967, an agreement providing for the development of the Sierrita Property was consummated with GSA. In March 1970, the first copper-molybdenum concentrates were produced in the 72,000 tons-per-day concentrator. Pre-production stripping exceeded the scheduled 105 million tons despite the comparatively short lead time available to equip the mine, build the mine plant facilities, and staff the technical, supervisory, and labor forces. Shortly before the first concentrates were produced, plans were being formulated for a program to increase processing capacity by approximately 15 percent, which entailed increasing overall mining capacity by approximately 50,000 tons per day. This program is scheduled for completion early in 1971.

Jan 1, 1971

By V. Svalbonas, S. M. Jones, A. Boughey

In June 1998, the world's largest grinding mill, the 40 ft (12.2 m) diameter 20 MW Cadia gearless SAG mill, was commissioned. This was a leap of over40% above the largest operating SAG mill. A significant saving in capital cost gave the incentive necessary for a single mill line at Cadia, resulting in the selection of the 40 ft (12.2 m) diameter mill. Now, after 18 months of operation, this paper will review the significant items involved in the development, design, manufacture, installation, and operation of this mill.

Jan 1, 2000

By Thomas B. Dellinger

Since 1950, rotary-drilled, large-diameter holes, "big holes", with depths to over 300 feet and diameters of over 30 inches have become numerous and widespread. Fifty such holes have been drilled for access shafts to LPG mined caverns, Over 70 holes have been drilled for various other commercial activities. Over 150 holes have been drilled by the AEC for atomic testing purposes. These holes have ranged up to 300-inch diameters with 1600-foot depths, 120-inch diameters with 570-foot depths, and 72-inch diameters with 4800-foot depths. Equipment is now available to drill a wide range of ?big holes? for varied requirements. At least a dozen drilling contractors are participating or have participated in the drilling of "big holes". The accompanying tables show pertinent historical, technical and economic data on over 250 holes. The illustrations to be presented concurrently with this paper show the various equipment and accessories now in use. (See Appendix I for captions of illustrations.)

Jan 1, 1966

By H. Walqui, S. K. Kawatra, C. H. N. Jeltema, T. C. Eisele

"The literature contains many cases where the shape of hydrocyclone efficiency curves vary from the ideal “S” shape. These often appear as changes in slope, or inflections. These inflections can be divided into two categories: “coarse” inflections, which are caused by differences in the density of the minerals being separated, and “fine” inflections, for which there are a number of competing hypotheses concerning their cause. The existing literature addresses either the fine inflection or the coarse inflection, but no studies have reported both types of inflection occurring at once in a single hydrocyclone test. This paper presents hydrocyclone results from both in-plant studies and laboratory experiments that show both coarse and fine inflections. The industrial significance of both types of inflections is discussed. Hydrocyclone efficiency curves show the probability of any given particle reporting to the hydrocyclone underflow as a function of its size. Selectivity functions, S(d), are used to represent the fraction of the feed material in size fraction “d” that reports to the hydrocyclone underflow. Plotting this selectivity function against particle size leads to the standard S-shaped efficiency curve shown in Fig. 1. In the “uncorrected” curve, the tail of the curve does not go through 0%, because a portion of the solids are bypassing classification. This happens as a result of their being carried along by the water, and is most important for the finest particles. The curve can be “corrected” based on the fraction of the feed water that reports to the underflow, represented by the bypass fraction, Rf. The corrected curve then normally passes through 0%, as shown in Fig. 1.The act of correcting the curve shifts the value of the d50 size, which is the size that has an equal probability of reporting to either the overflow or to the underflow. Correcting the curve to account for the bypass fraction makes it easier to compare cyclone performance under conditions that cause the quantity of water in the underflow to change, but that do not alter the actual corrected cut size (d50c).The corrected efficiency curve can be expressed closely by equations such as Plitt (1976):"

Jan 1, 2015

By T. W. Camm, J. C. Johnson

"Most engineers are bright, hard-working, reliable, and prefer to avoid conflict. An engineering curriculum tends to self-select these characteristics. By most standards, you would expect workers exhibiting these traits to require minimal supervision. But is this true? Is this how most current engineering managers lead? Looking at some current theories on leadership combined with personal anecdotes, this presentation will look at some common misconceptions about leading engineers. INTRODUCTION Engineers are different. There, I said it. From cartoons to movies to television shows and more, there is a reason stereotypes of engineers resonate. Yes, you can find a lot individual variation among specific engineers that you may know, but they do not detract from some of the overarching commonalities that engineers are known for. CLASH OF CULTURES What distinguishes a professional? Raelin (1985), in his book Clash of Cultures, describes professionals as having superior intellectual training, maintaining their own standards of excellence, and being supported by associations that maintain the quality of the profession. Engineers obviously fit all three of these ideas: an engineering degree is still regarded as one of the most difficult to attain; we maintain our own standards both through peer-review and in academia through ABET accreditation; and through professional associations like SME to promote the discipline and foster technical growth. Six characteristics representing professional status are described by Raelin (1985, p. 9): Expertise—prolonged specialized training in a body of abstract knowledge. Autonomy—the freedom to choose the means to solving a problem. Commitment—Primary interest is in pursuing the practice of one’s own chosen specialty. Identification—identifying with the profession and with fellow professionals, both through formal associations and through peers external to the organization. Ethics—providing service without concern for oneself or without becoming emotionally involved with the client. Standards—committed to help in policing the conduct of fellow professionals. Engineers are smart—in the vast majority of cases they are hired specifically for their expertise. At a typical mine the ventilation engineer knows more about mine ventilation than anyone else on site; the same is true for the rock mechanics engineer, the explosives engineer, the planning engineer who uses sophisticated computer software—each is the expert for their particular area at the mine, that is why they have that responsibility. This can present a dilemma for the manager for each (sometimes all) of these engineers. The manager needs to keep everyone on task and focused on the objectives of the organization. Because they have reached their position by following the norms and expectations of the organization, managers tend to have a certain amount of loyalty to that organization. Engineers, on the other hand, tend to have mixed loyalties, having as much loyalty and affinity for their profession as they do for the particular company they happen to be working for at any given time. This dynamic can often lead to power struggles between the managers, based on their positional authority in the organization, and the engineers, based on their specialized knowledge (Camm, 2013)."

Jan 1, 2017

By A. Toulet

Mining in Mexico has an immensely long tradition. In his first letter reporting to his sovereigns from Mexico in July 1519, Hernan Cortes wrote, "To our mind it is probable that this land contains as many riches as that from which Solomon is said to have obtained the gold for the temple ... " 1 But not only precious metals drew his attention; in 1524, short of tin needed for casting guns, he found it in Taxco. As an added bonus, those who went to find it "came upon a vein of iron which according to experts is of exceeding richness." 2 Since then, the mines of Mexico have yielded up enormous amounts of mineral and continue to do so. Mexico is one of the ten principal producers of minerals world-wide both in production volume and diversity. 3 Tens of thousands of tons of iron, sulphur, and coal come from its earth annually; it is the world's second largest producer of barite; third of antimony, fluorite, and graphite. 4 Copper, zinc, silver, molybdenum, arsenic and cadmium and others are wrested from its mines. The wealth is there; but maximizing its potential requires investment, technology, and markets that the government cannot itself provide. Thus mining, since the Mexican Revolution severely restricted by both law and policy, today looks to private investment to develop its potential. Beginning in 1989, Mexico's government cautiously moved to liberalize private sector involvement in mining; 5 using the experience gained from these initial regulatory changes, the current new Mining Law was adopted in 1992 6 and new Regulations were promulgated in 1993. 7

Jan 1, 1994

By S. Shankar

The phenomenon of re-entrainment is seen in many environments, such as sand dune formation, soil erosion, fugitive dust emission, spread of radio-nuclides from accidents involving nuclear weapons, etc. A thorough review of existing literature reveals that work on re-entrainment of dust in mine atmospheres is scant. The present study was undertaken as a part of on-going research in the Respirable Dust Center on the prediction of ambient dust concentrations in mine atmospheres (Ramani, 1983) to identify the contribution of dust re-entrainment to the spatial and temporal concentration in mine airways and included wind tunnel and in-mine studies. A set of seven experiments was performed in the Lake Lynn Laboratory Mine of the U.S. Bureau of Mines under controlled conditions. Two types of source coal dust, anthracite and bituminous, were subjected to airflow with a range of air velocities and walking by two different walkers. Six sets of instruments were used along the test section for sampling the airborne particles. The average air velocity just above the coal dust bed varied between 2.1 to 3.2 m/s. Due to re-entrainment as a result of the airflow conditions, the average respirable dust concentration in the center of the mine airway increased to 0.09 mg/m3 (standard deviation = 0.03 mg/m3) from an initial value of 0.02 mg/m3. The concentration of all (i.e., total) airborne dust increased to 0.17 mg/m' (standard deviation = 0.10 mg/m') from an initial value of 0.02 mg/m'. The action of personnel walking, however, caused higher amounts of re-entrainment. Average concentration of respirable and total dust at the center of the mine airway increased to values of 0.28 mg/m' (standard deviation = 0.12 mg/m3) and 1.53 mg/m3 (standard deviation = 1.0 mg/m3), from initial values of 0.03 mg/m3 and 0.05 mg/m3, respectively. It was observed that the dust concentrations were much higher near the floor, immediately at the end of the coal dust bed. Depending on the level of activity and the amount of dust deposited in mine airways, re-entrainment may become an important source of airborne dust away from the point of actual mining.

Jan 1, 1994