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By D. A. Stanley

The Bureau of Mines has developed an equation which describes the dewatering of clay slurries with polyethylene oxide. This dewatering equation is An _ kC + b where n, k. and b are empirical constants and C is the concentration of polymer solution used for dewatering. The minimum amount of polymer required for dewatering. A is determined by titration of the clay slurry with a polymer solution. In this report, the dewatering equation is applied to data obtained for a polyacrylamide dewatering sequence. Fundamental similarities between the dewatering properties of clays and the viscosity of polymer solutions were observed. These similarities were explained by correlating an equation describing polymer configuration in solution with the dewatering equation. This correlation showed that amount of polymer required for dewatering and polymer solution viscosity are both consequences of the frictional interaction of entangled polymer coils.

Jan 1, 1989

By S. K. Nandy, K. F. Unrug

The behavior of the immediate shaly roof within a drift mine in eastern Kentucky, operating in the Coalburg seam, was monitored with roof bolt dynamometers. The dynamometers used were capable of providing readings up to a maximum load of 66.72 kN (15,000 lbs), + 0.022 kN (5 lbs). To observe the influence of moisture on the rate and extent of roof deterioration, 57dynamometers were installed on roof bolts anchored in intake and neutral airways. Monitoring was conducted during the period January 1985 through January 1986, in order to cover a full cycle of dry and wet mine conditions associated with seasonal changes. Analysis of the data showed a trend of roof deterioration induced by condensation and eventual penetration of moisture within the immediate roof bedding planes, leading to rapid disintegration of the mine roof.

Jan 1, 1987

By Kenneth N. Han, Xinghui Meng, A. J. Monhemius, G. W. Warren, E. Peters

The leaching of gold in ammoniacal solutions containing various oxidants in an autoclave has been investigated. Also studied are thermodynamic and kinetic aspects of the ammonia leaching of gold and silver. The variables studied included stirring speed, temperature, concentration of ammonia, ammonium, and various anions. Arrhenius activation energy values were found to be between 47.1 kJ/mol and 86.2 kJ/mol depending upon the kinds of oxidants used. An electrochemical reaction controlling mechanism has been proposed for the dissolution reaction. The leaching behavior of gold and silver from various ores in ammoniacal solutions containing various oxidants has also been investigated. The samples tested included sulfidic as well as carbonaceous refractory ores. The recoveries of gold and silver from these refractory ores were better than 96% within 2-3 hours of leaching time.

Jan 1, 1993

By N. Zouridakis

Radon is an established lung carcinogen and remains the most important environmental radiation exposure. It was estimated that inhalation of radon and its short-lived decay products accounts on average for about one half of the effective dose equivalent from all natural sources of radiation. Radon contained in household water can transferred into room air as a result of agitation or heating. An indoor air concentration of 400 Bq/m3 is obtained if the radon concentration in household water is 4 kBq/L. A typical concentration of radon in ground water is about 1000 times lower than this concentration but sometime elevated concentrations of some kBq/L in ground water are reported. In this work a Liquid Scintillation counting system is used to obtain fast and accurate measurements of radon (222Rn) in water samples from Argolida, Greece.

Jan 1, 2007

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

By Lynton Gormely

Oxygen transfer in a stirred tank bioreactor has been modelled mathematically to determine the trends in transfer efficiency (kg transferred per kWh) with transfer rate, tank depth, superficial velocity, and mixer power.

Jan 1, 1990

By T. P. Weldum, W. Conger

"The objective of the study conducted by the authors is to determine what media size(s) addition will maximize any given plant ball mill’s grinding efficiency (Fig. 1). The functional performance parameters “mill grinding rate through the size of interest,” and “cumulative mill grinding rates” from both plant and smallscale tests are applied to this task. A plant media sizing methodology, and industrial case studies, are provided.BackgroundA previous article (“Ball mill classification system optimization through functional performance modeling,” Nov. 17, McIvor et al., 2017, Mining Engineering) described circuit classification system efficiency (CSE), equal to the percentage of coarse (plus circuit P80 target size) material in the ball mill. It can be measured and then increased through pump and cyclone adjustments. By so transferring the application of ball mill power from fine (minus P80 product size) to coarse material, circuit efficiency is increased and overgrinding is decreased.The other factor that determines overall circuit efficiency is how well the power that is applied to the coarse material grinds it. This is affected by design and operating variables like the media sizing and mill percent solids. From circuit functional performance (McIvor et al., 1990):Circuit production rate of new fine material (t/h) = Power applied to coarse material (kW) x Mill grinding rate of coarse material (t/kWh) (1)The power applied to coarse material is the total mill power multiplied by the CSE. Therefore:Circuit production rate of new fine material (t/h) = Total mill power (kW) x CSE (%) x Mill grinding rate of coarse material (t/kWh) (2) Production rate, mill power and CSE can be measured during a plant circuit survey. The mill grinding rate of coarse material is then calculated. At the circuit target P80 size, this is the mill grinding rate (of coarse material) through the size of interest. The objective is to maximize this rate. The mill grinding efficiency can also be calculated by taking the ratio of the mill grinding rate over the material’s grindability as measured in the lab. Mill percent solids optimization has thusly been achieved (McIvor et al., 2000). For media sizing, maximizing the mill grinding rate on a given ore is equivalent to maximizing the mill grinding efficiency. Ores of different grindability will be part of the discussion."

Jan 11, 2018

By Arthur B. Johnson

Potassium, is one of the three basic plant nutrients along with nitrogen and phosphorus. There is no substitute for potassium compounds in agriculture; they are essential to maintain and expand food production. Potash is found throughout the world in both soluble and insoluble forms. It is only the soluble forms that are economically attractive to process, primarily chlorides and sulfates, with potassium chloride by far the most important potash salt. Sylvinite ore, a mechanical mixture of sylvite (KCl) and halite (NaCl), is a natural mineral of potassium. This ore is found in commercial quantities in the Permian sin region near Carlsbad, New Mexico. Vast tonnages of high-grade de, low-clay-content ores have been mined and concentrated during the past 40 years; however, these deposits are rapidly being depleted leaving high-clay, low-grade ore reserves of over 135 million tons. These ores, containing 13 to 22 percent K20 and 3 to 8 percent clay slimes; cannot be effectively processed by current commercial flotation methods. The present concentration processes for potash minerals in Carlsbad relate mainly to ores containing 1 percent or less of clay-slimes with a K20 content of 20 to 25 percent to yield products to meet the requirements of the agriculture market, namely, 95 percent KC1 equivalent to 60 percent K20. Discoveries of potash ore in Saskatchewan, Canada, have shifted the emphasis on expansion outside the United States. Presently, the American companies are looking to Canada for future growth, expenditures, and expansion. Four of the seven companies operating near Carlsbad are produces of Canadian potash despite the large investment sons required for shaft sinking and milling installations.

Jan 1, 1970

By M. Szatkowski

The studies conducted by Laplante and coworkers (1983,1984) with flotation of galena and quartz, showed that the rate of flotation reaches a maximum value for a certain intermediate value of the air flow rate. A similar, though less pronounced, effect of the air flow rate on the selectivity was also observed. Luttrell and coworkers (1985) showed that the size of air bubbles has a significant effect on both rate and selectivity of coal flotation. These two studies used flotation tests in which no froth layer was formed. This note describes effects of the air flow rate and bubble size on the rate constant and selectivity of coal flotation, using a mechanical laboratory machine with a froth layer allowed to form. Experimental A raw, dry ground Illinois coal was used in flotation tests. The size distribution and ash content of the samples are : Particle size (µm) Weight (%) Ash (%) +600 4.30 41.7 600/200 43.05 19.0 200/74 22.25 16.0 - 74 30.40 24.0 average 20.8 No. 2 fuel oil with a concentration of 200 g/t of feed and M.IBC with concentrations of 11, 22, and 33 mg/L of pulp (128, 256, 384 g/t, respectively) were used as flotation reagents. Tap water with a natural pH of 7.0 to 7.8 was used throughout the tests. No other reagents were used. The starting concentration of coal in the pulp was 86 g/L. A standard flotation machine with the pulp volume of 5 L was used in experiments. Three different air flow rates were employed, resulting in three air escape velocity values of 14, 30, and 70 cm/min. The air escape velocity (AEV) was calculated as the ratio of air flow rate (Qa) to the slurry surface (Ss) in the machine. For a comparison, the upper limit of air escape velocity values used in commercial size flotation machines is between 70 and 100 cm/min (WEMCO, 1983). A machine consisting of a generator of fine bubbles and a separatory column, described elsewhere (Szatkowski and Freyberger, 1985), was used for some tests where bubble size lower than that generated in the standard machines was required. Prior to a flotation test, a batch of coal was mixed with water in the flotation cell for at least 10 minutes to ensure that all coal was thoroughly wet. Then the specified amount of fuel oil was added and the pulp was conditioned for five minutes. The specified amount of MIBC was added 15 seconds before the end of conditioning. After air flow was opened, flotation products were collected for 0 to 30, 30 to 60, and 60 to 120 second intervals. Content of the combustible material in samples was determined by subtracting the ash content from 100%. A photographic technique was used to determine initial distributions of bubble size generated by the flotation machines. Bubbles generated in water containing the prescribed amount of the frother but no solids were photographed against a black background. The average diameter and standard deviation were then calculated for each set of flotation conditions. The method gives results reproducible within k 5% of the average bubble size. Table 1 contains information about the measured bubble size distributions expressed by the average bubble diameter (Db) and the standard deviation ( SD ) . Results and discussion The general formula for calculating the flotation selectivity is: o = recovery of coal/recovery of ash (1)

Jan 1, 1988

By M. Sadegh Safarzadeh, Stanley M. Howard

The two-dimensional predominance area (Kellogg) diagram (PAD) surfaces bounding the three-dimensional Cu-As-S-O system were constructed at 900 K. The computations employed gases As4O6(g), O2(g), and SO2(g) as the independent coordinate species to generate diagrams of As4O6(g) − O2(g),O2(g) − SO2(g), and As4O6(g) − SO2(g), while the third gas (As4O6, O2, or SO2) was fixed. The results indicate that there is no shared boundary between enargite (Cu3AsS4) and copper sulfate (CuSO4); however, at PO2 < 10−6 atm, PAs4O6 < 10−10 atm, and PSO2 > 10−4 atm, CuSO4 is the stable phase at 900 K. The chemical driving force for oxidation of intermediate phases between enargite and copper sulfate is substantial except for basic sulfate (CuO · CuSO4).

By M. Sadegh Safarzadeh, Stanley M. Howard

The two-dimensional predominance area (Kellogg) diagram (PAD) surfaces bounding the three-dimensional Cu-As-S-O system were constructed at 900 K. The computations employed gases As4O6(g), O2(g), and SO2(g) as the independent coordinate species to generate diagrams of As4O6(g) − O2(g),O2(g) − SO2(g), and As4O6(g) − SO2(g), while the third gas (As4O6, O2, or SO2) was fixed. The results indicate that there is no shared boundary between enargite (Cu3AsS4) and copper sulfate (CuSO4); however, at PO2 < 10−6 atm, PAs4O6 < 10−10 atm, and PSO2 > 10−4 atm, CuSO4 is the stable phase at 900 K. The chemical driving force for oxidation of intermediate phases between enargite and copper sulfate is substantial except for basic sulfate (CuO · CuSO4).

By J. Hanna, H. R. Gonzalez

Discussion by N.L. Hargett Dear Dr. Hanna: As discussed in our phone conversation, we are enclosing a suggested revision to the paper, "Application of Bentonitic Clays in Suspension Fertilizer Formulations," published in MINERAL & METALLURGICAL PROCESSING, May 1990. The revision will replace the second paragraph in the article. The reason for the revision is that suspensions have decreased in market share since 1984 and at present comprise an estimated 25% to 30% of fluid mixtures. This represents 6% to 7% of all fluids and only 2% to 3% percent of all fertilizers. Based on national surveys conducted by the Association of American Plant Food Control Officials (AAPFCO), suspensions comprised an estimated 25% of all fluid mixtures in 1974, 33% in 1976, 35% in 1980, 40% in 1984, and 36% in 1988. Revision to "Application of Bentonitic Clays in Suspension Fertilizer Formulations." Suspension fertilizers are an important segment of the fertilizer industry in the United States. Suspensions allow higher concentrations of nutrients (N, P2O5, K2O) than clear liquid solution fertilizers. K20 content of solutions is limited to about 10%, while suspensions can contain more than 30% K2O. Suspensions can tolerate greater variations in temperature than solutions. In suspensions, if crystallization occurs, the crystals can remain suspended in a clay jell instead of settling as they would in solutions. Solubility is not a limiting factor in suspensions, since complete solubility of phosphate is not required. This allows a wider range of phosphate materials to be used. Also, materials commonly used in formulating suspensions are less expensive and more widely available than those for clear liquids (Getsiner et al., 1984).

Jan 1, 1992

By Thomas F. Martin, Daniel J. Sieve

In April 1988, the joint venture of Morrison-Knudsen and Paschen Contractors (M-K/P) was the successful low bidder for the Metropolitan Water Reclamation District of Greater Chicago&apos;s (MWRD) Contract Number 73-164-2H. This project is yet another in the MWRD Tunnel and Reservoir Plan (TARP) devised to control and treat overflows from combined sanitary and stormwater flows. M-K/P&apos;s project, located along the Des Plaines River, consists of 21,569&apos; of tunnel mined to a 35&apos; diameter and two tunnels totaling 14,529&apos; mined to a 12&apos; diameter. All three tunnels will have 12" thick reinforced concrete lining. The path for combined sewer/storm waters to reach the tunnels begins at a connection structure, continues through an approach sewer, a sluice gate structure and down a vertical shaft which is topped by a backflow structure. The purpose of the backflow structure is to prevent water surging to the surface as the system fills rapidly. Backflow preventors represent a change in design after problems developed on earlier constructed portions of the T.A.R.P.. On this project, 14 drop shafts channel flows to the tunnel system. In addition to the drop shafts, contract specifications require an emergency or escape shaft be located not more than 4000 feet from the tunnel headings. Due to long reaches without drop shafts, four access shafts had to be constructed. Thus, the total shafts constructed under this contract will be eighteen plus one 30&apos; finished diameter construction shaft located within Miller&apos;s Meadows. Figure 1 shows the overall Des Plaines River System and the M-K/P contract connection to previous and future MWRD projects.

Jan 1, 1991

By Georgene Renner, William R. Yernberg, Steve Kral

Allmand exhibited its LSC 100Light Sequence Commander, the mining industry’s first automatic lighting and command control system. The system is available with all of Allmand’s portable light towers. The LSC100 uses a preset photocell sensor that allows the light tower to be automatically turned on or off. At dusk, the sensor engages the power unit and then activates the light. At dawn, the sensor automatically shuts off the light. The unit also features an eight-event digital timer that allows eight separate sequences daily. The LSC100’s sequenced light control ensures that the engine is started before the lights are turned on or off. This prevents possible generator damage that can occur if the unit is powered down under load. Also, a low-fuel shutdown sequence helps protect the generator from damage by automatically shutting down the unit when the fuel level reaches 10 percent. And the LSC100’s smart circuitry senses engine water temperature during startup, determines the glow-plug time requirement and schedules repeat intervals, if necessary, to start the diesel engine.

Jan 1, 2004

By Liming Yuan, John C. Edwards, Gene F. Friel, Robert A. Franks

Diesel fuel fire experiments were conducted in NIOSH’s Pittsburgh Research Laboratory’s (PRL) Safety Research Coal Mine (SRCM) to determine the critical air velocity for preventing smoke rollback. Such information is necessary for the preplanning and implementation of ventilation changes during mine fire fighting and rescue operations. The fire intensity varied from 50 kW to 300 kW depending upon the fuel tray area. Airflow in the 2 m high and 2.9 m wide coal mine entry was regulated during the course of each experiment; measured upwind from the fire as an average over the entry cross-section with an ultrasonic airflow sensor; and recorded dynamically with a mine monitoring system. The extent of smoke reversal was monitored with light obscuration monitors, ionization smoke sensors, and visual observations. Experimental results for the critical air velocity for smoke reversal as a function of fire intensity compared very well with model predictions based upon a computational fluid dynamics (CFD) fire dynamics simulator.

Jan 1, 2005

By M. Okido, W. Wei, L. Chai

The kinetics on the growth of Thiobacillus ferrooxidans was studied by use of series piezoelectric quartz crystal ( SPQC ). Arsenic ioninhibits the growth of Thiobacillus ferrooxidans, which is consistent with the previous results by other measuring method. This further demonstrates that the SPQC can monitor the chemical activity of the growth of Thiobacillus ferrooxidans. Cupric ion accelerates the growth of Thiobacillus ferrooxidans. The mechanism was discussed, it was considered that copper promotes the protein enzyme of Thiobacillus ferrooxidans, rustic yanin, to form over the range of cupric ion concentration studied. There action order of cupric ion in accelerating the bacterial growth is 0.067. The growth of Thiobacillus ferrooxidans is dependent on temperature, the apparent reaction activation energy decreases from25.56 kJ / mol to 18.32 kJ / mol because of the addition of 10 mg / L Cu(II) to the bacterial growth solution of pH 2.0 at initial inoculum of 10%.

Jan 1, 2000

By N. K. Bohidar

Evaluation of the stability of a structure under given boundary conditions requires accurate knowledge of the state of stress as well as the mechanism of failure at the critical state of stress. This paper represents a part of the research done by N. K. Bohidar during 1965-1968 for his Ph. D. research in Mining Engineering at the University of Utah. It summarizes the determination of internal stress and failure phenomena around cylindrical openings in viscoelastic media subjected to lateral restraint and vertical uniaxial load. The conditions of lateral restraint and viscoelastic nature of the media represent conditions associated with subsurface openings in rocks. The financial assistance provided to the project from the University of Utah Research (Uniform School Fund), W. C. Browning Scholarship Fund, Earth Science Research Fund, and D. C. Jackling Loan and Scholarship Fund is gratefully acknowledged. The authors also acknowledge the assistance of S. Siugh, W. Jones, and K. McCarter during the project.

Jan 1, 1970

By Jorge Rubio, Luis Valderrama, Carlos Pérez

This work describes the effect of the degree of energy transferred to the pulp, during the conditioning stage, on gold flotation fines recovery, grade and kinetics, with three types of frother (Dowfroth 250, pine oil and MIBC) and amyl xanthate as collector. Best results show an increase in 24% recovery and 50% in concentrate grade, depending on the intensity of conditioning and type of frother (better with Dowfroth 250). For low or intermediate energies (0.5-1.5 kWhlm3 pulp), the fine particles adhere to larger ones, increasing the recovery due to a mixture of carrier (pyrite, in this case) and autogeneous carrier flotation and at higher shear (2-3 kWhlm3 pulp), gold fine particles aggregate themselves. These phenomena are demonstrated by measurements of flotation rate, true flotation (flotation by actual particle-bubble adhesion), and by the amount of entrained particles. Alternatives for this type of pulp conditioning and mechanisms involved are discussed.

Jan 1, 1995

By Rick Sarkisian, Albert D. Bonneau, David Keech

The Drummond Co. Inc.&apos;s Shoal Creek Coal Mine Slope Belt Conveyor transports 5000 mtph (5500 tph) of material over a length of 1400 m (4600 ft.) and a lift of 3609 m (1180 ft.). Four 1500 kW (2000 hp) motors power the 1830 mm (72 in.) Goodyear Flexsteel 5800 steel cable belt. In May 1993, Dodge Manufacturing was selected to design and build the conveyor pulley assemblies for this conveyor. The conveyor includes five pulley assemblies. The 2135 mm (84 in.) diameter by 2135 mm (84 in.) face primary and secondary drive pulleys were designed for a resultant belt load of 3060 kN (688,000 Ib.) This paper describes the design and manufacture of the two drive pulley assemblies. Design details included are component evaluation, material selection, locking assembly selection, bearing selection and lagging selection. Manufacturing details include material procurement, welding process, lagging installation and component assembly. Quality control procedures are also discussed.

Jan 1, 1996

By S. Saklar, i. Bayraktar

The effect of the feed size distribution on the kinetic parameters in pyrite flotation was studied. The kinetic parameters of the particular sizes were determined as a function of whether other particle sizes were present. From the data, a relation between the pyrite flotation rate and the particle size distributions was obtained. It was shown that the flotation rate constant (k) is a function of the size modulus (dt) of the Gates-Gaudin-Schuhmann distribution, and its variation may be expressed as k = a • exp(b . d,) for either individual size fractions or composite feed. Parameters a and b are dependent on the particle size distribution of the feed. Specifically, parameter a increases linearly as the distribution becomes coarser. The above relationship was also tested under conditions in which the collector (KAX) dosage was varied from 130 to 278 g/t. It was shown that the established relation was not dependent on the collector dosage.

Jan 1, 2002