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By Yoginder P. Chugh

Effects of moisture absorption on the behavior of the Dykersburg shale overlying Harrisburg coal seam in Southern Illinois are presented in this paper. Cylindrical samples of the shale 2.5 cm in diameter and 5.0 cm long were studied for weight gain and swelling strains at 90, 95, and 100 percent relative humidity and under immersed water conditions. After attaining equilibrium conditions, the samples were studied for changes in the strength and deformation behavior. The depth of moisture penetration was measured using the Vickers microhardness measurements. Using the basic mass transfer concepts, effective diffusion coefficients were computed for the shale.

Jan 1, 1980

By A. L. Mular

Introduction The cost of grinding may be divided into a fixed equipment cost portion and an operating cost portion. The fixed equipment cost is the amount of money required to purchase the grinding mill itself. This amount may be kept separate from the installation cost which, often, is taken as a percentage of the fixed equipment cost. A good average is 43%; therefore a $100,000 ball mill would cost approxi¬mately $143,000 installed. Operating costs are variable and paid per unit time for power, liner replacement, grinding charge replenishment, labor, and mainte¬nance including major parts replacement. Such costs depend on the geographical location and will vary with the cost of labor. Operating costs are difficult to estimate. Most mining companies consider such information proprietary. This section provides a means to obtain an estimate of the equip¬ment cost of a grinding mill with a rated accuracy of±l0%. Compara¬ble methods to determine operating costs are not available. Such costs are presented in table form with suggestions for updating them. Esti¬mating the cost of liners and steel replacement is possible as described in a later section. Capital Costs of Grinding Mills Grinding mill costs have been related,124 by means of equations and corresponding graphs, to the nominal horsepower requirements of mills. All graphs are based on a Marshall and Stevens cost index for Mining and Milling of 300, where a cost index is a ratio used to estimate current prices of equipment from obsolete prices. Where (Cost Index Now) Cost Now = (Cost Then) (Cost Index Then) (I) For example, the cost of an autogenous mill that requires a 3000¬hp motor is estimated as $500,000 (Fig. 95). This was the estimated price when the M&S (Marshall and Stevens, Mining and Milling) cost index had a value of 300. If the index is currently at 400, then the estimated current price of the autogenous mill is: Cost Now = ($500,000) (400) (300) = $667,000 Most professional journals will publish current values of cost indices at least once a month. In subsequent pages, each graph of grinding mill cost vs. mill horsepower has an associated graph of capacity vs. horsepower along with a table that relates mill size to horsepower. The associated graph and the table must be used with caution, since both apply to an average ore under average conditions. Example calculations showing how graphs should be used are shown in the following. It should be stressed that the cost data incorporated herein ex¬cludes the cost of pumps, motors, lines, and classifiers, unless other¬wise stated. Costs are not direct quotes; they are the averages of quotes supplied by manufacturers and analyzed by means of regression analysis. Example 1. Calculate the power required to grind 100 tons of dry solids per hour (in a closed circuit with a classifier) when the

Jan 1, 1985

By T. A. Marcotte, D. J. Bobrow, S. Vaghar

This paper discusses the geotechnical instrumentation program on the Central Artery/Tunnel Project, in Boston, Massachusetts. Particular attention is given to a state of the art Graphic Information System (GIS) application which has been developed for rapid analyses and reporting of data. This application will see extensive use during the ongoing complex downtown tunnel construction. The excavations in soft ground require cuts of up to 120 feet deep and 250 feet wide. Deformation and groundwater monitoring instruments have been installed to monitor the stability of the excavation support systems and the abutting buildings. The use of this application is illustrated with an example from an ongoing contract.

Jan 1, 1997

By Harry C. Verakis

A conveyor belt fire presents a serious safety hazard in an underground coal mine. A review of underground coal mine fires involving conveyor belts which occurred within the past two decades was made by the Mine Safety and Health Administration (MSHA). Belt fires that occurred during this period of time have burned as much as 610 m of belting before the fire was extinguished. In an effort to improve safety, MSHA has undertaken a program to study the flammability of conveyor belts in large and small-scale tests. Large-scale tests were conducted on various types of mine conveyor belts using a range of airflows from 0 to 6.1 m/s. Data were obtained to evaluate the effect of airflow on flame spread and other combustion properties of the conveyor belts. The results from the large-scale tests were utilized to develop a new MSHA laboratory-scale test for evaluating the flame resistance of conveyor belts used in underground coal mines. Agreement of results based on a pass/fail criterion was obtained between the large-scale test and new laboratory-scale test. A summary of test results from the new test is shown for different conveyor belt formulations. MSHA also performed a series of fire tests with wood samples using the new laboratory-scale tunnel apparatus. These tests were conducted at airflows from 0.5 to 5.1 m/s to determine the effect on flame spread and burn damage. The data were also analyzed to determine if wood could be used as a calibration standard. The results indicate the flame spread behavior of the wood follows trends similar to results obtained with conveyor belts. The data did not show the precision needed to use wood as a calibration standard. MSHA intends through rulemaking procedures to replace the "20g" conveyor belt acceptance test with the new laboratory-scale test. MSHA also plans to implement a voluntary acceptance program. Under this program, a. conveyor belt manufacturer could obtain an MSHA acceptance for belts meeting the new test criterion. A belt accepted under the voluntary program could then be used underground. Use of conveyor belts meeting the new test would limit the propagation of fire and reduce the probability of igniting other combustibles in the belt entry.

Jan 1, 1991

By Kimberley MacHardy, Ann Maest, Greg Lawson, James Kuipers

Although water quality predictions are often made in Environmental Impact Statements (EIS's) as part of the National Environmental Policy Act (NEPA), comparisons between predicted and actual water quality conditions are rarely reviewed. A comprehensive review of predicted and actual water quality conditions was conducted for a number of large U.S. hardrock mines with EIS’s. For mines with large differences between predicted and actual conditions, factors responsible for the deviations, such as characterization and modeling errors or engineering failures, were identified. Recommendations are made for how the NEPA process could be modified to better predict water quality at hardrock mine sites.

Jan 1, 2005

By M. C. Fuerstenau

Collophane, cryptocrystalline fluor-apatite, is examined as a substrate for plant water and wastewater remediation. The attenuation of Fe3+, Fe2+, Hg2+, Pb2+, and Cd2+ is presented as a function of metal ion concentration, amount of collophane, initial pH and conditioning time. TCLP results for Pb2+contaminated collophane are presented. Mechanisms of adsorption for divalent and trivalent species are suggested.

Jan 1, 2001

By Matthias Rheinlander

The paper will describe Virtual Design and Construction (VDC) as a tool which will enable us to build a project ?virtually? on a computer before constructing and operating it in the real world. VDC computer models are built, using information from various sources including survey, 3D laser scanning, CAD and GIS to represent the project and its components in a coordinated and consolidated infrastructure model. This model is reviewed and evaluated by, and shared among project teams including owners, con-tractors, designers, engineers, construction managers, stakeholders and operators. It allows teams to fix costly design and construction issues before construction on site. Benefits and lessons learned from global projects will be discussed

By J. H. Calder, J. C. Hower

Three lithotypes of Phalen coal from Cape Breton County, Nova Scotia, Canada, were subjected to a modified Hardgrove grindability index (HGI) test. The testing scheme differed from conventional (ASTM D-409) Hardgrove testing by, in one case, repeating the test three times prior to removing the -200-mesh fine coal from the Hardgrove mill, and, in a second test, removing fines after each of three successive Hardgrove tests. Only the first stage of the latter testing scheme actually conforms to the standard HGI test. All processed samples were analyzed petrographically by a combined maceral/microlithotype analysis. The increase in the percentage of coal ground to -200 mesh increases through the 60- to 180-mesh fractions for each of the lithotypes for the following two apparent reasons. First, with progressive grinding, more of the oversize material is close to 200 mesh (75 µm), creating a "pool" of particles ready to be ground to -200 mesh, and, second, the removal of fines from the system at each step eases the grinding at the successive step. The fines retained system shows the influence of buffering of further grinding by the particles retained in the grinding mill. The Phalen lithotypes and the series of Central Appalachian coals, evaluated previously by Hower and Wild (1994), differ in that an increase in vitrite and a decrease in trimacerite is observed in successive -200-mesh fractions in two of the Phalen lithotypes. This is the opposite of the trend observed in the slightly lower-rank Central Appalachian coals.

Jan 1, 1998

By Austin Whillier

INTRODUCTION Use of air-cooling or refrigeration equipment in underground mines is needed when conventional ventila¬tion techniques do not maintain acceptable environ¬mental temperatures in working areas. Because refrig¬eration can be very expensive, it should be implemented only after all possible and practical steps have been taken to eliminate or reduce heat sources in the mine. As an example, to prevent the main ventilation fans from contributing heat to mine air, they should be located in the air return and not in the air intake. It is particularly important to prevent any direct contact between hot water and ventilation air, especially in mines which encounter large flows of hot fissure water. Any water hotter than the prevailing wet-bulb temperature of the ambient air must be removed by pipes located as close as possible to the water source. This hot water must not be allowed free contact with the incoming ventilation air at any time during the water's passage out of the mine. Although insulation of the pipes carrying the hot water is seldom necessary, direct contact between the air and the water must be prevented so the warm water cannot evaporate. REVIEW OF COOLING PRACTICES Spot Coolers vs. Centralized Refrigeration To eliminate a few specific hot places in an otherwise cool mine, it is possible to use devices known as "spot coolers." A typical spot cooler that uses chilled water is shown in Fig. 1. These devices consist of self-contained refrigeration units that are often mounted on rail cars for haulage to hot spots. The cooling capacities of such spot coolers usually are limited to about 100 kW or 30 "refrigeration tons." A refrigeration ton represents a cooling rate that produces 1.0 st of ice in 24 hr; that is a cooling rate of 3.517 kW (200 Btu per min). Typically, the electric-power consumption to drive the compressor motor of the refrigeration plant in mines is 1.0 kW per refrigeration ton, corresponding to a coefficient of per¬formance of about 3.5. The principal difference between spot coolers and centralized refrigeration plants is the method of re¬jecting heat from the refrigeration system. Centralized refrigeration plants always discharge heat into the reject or return airflow of the mine; often that is the primary influence in selecting the location for the underground refrigeration plant. Heat from spot coolers usually is rejected into drain water or into air that is not flowing to the location requiring the cooling. As a result, spot coolers remove heat from troublesome hot spots in the mine, injecting that heat-plus the electrical energy used by the cooling unit itself-into other working areas where the ambient conditions are cooler. In effect, this is "robbing Peter to pay Paul." In deep, extensive mines, spot coolers usually pro¬vide only temporary and, over the long term, expensive solutions to localized cooling problems. Centralized re¬frigeration plants are preferred for such mines, with cooling distributed throughout the mine as required. Fig. 2 illustrates a typical underground centralized re¬frigeration plant. Centralized plants lend themselves to improved maintenance at reduced costs while offering the economy of size. Refrigeration plants of larger unit sizes have considerably lower initial costs than smaller unit sizes. The remainder of this chapter is devoted to large refrigeration plants, with no further consideration of spot coolers. Cost of Refrigeration Total Cost: The total cost of refrigeration amounts to about $200/kW of cooling per year (1981 US $). This total cost breaks down into approximately three equal parts: 1) Financial charges, which include the interest and amortization on the capital cost of the initial installation, and the cost of necessary underground excavations. 2) Operating and maintenance costs which include the cost of the electric power to drive the refrigeration plant's compressors. 3) Distribution costs which include costs for pump¬ing, insulated piping, and air-to-water heat exchangers. The local cost of electric power, the number of operating months per year, and the method of refrigera¬tion distribution all contribute to the actual costs in¬curred in a given application. However, the variations usually are limited to no more than ±30% of the $200/ kW per year total cost figure. Cost Per Ton: Refrigeration cost per ton of mineral production can be calculated if the annual production tonnage from the refrigerated section of the mine is known. In most cases, this cost will be less than $1 .00/t. However, in deep mines with high rock temperatures, such as those found in South Africa, the total cost of refrigeration can increase to several dollars per ton of broken rock. Distribution In deep extensive mines, distributing refrigeration often accounts for about half the total cooling costs. As a result, careful consideration and planning must be

Jan 1, 1982

By T. S. Golosinski

Introduction Typical dragline rigging includes several sets of ropes: hoist ropes, drag ropes, dump ropes and boom suspension ropes. The drag and hoist ropes contribute most to the dragline operating cost. Both work in very difficult conditions and wear out rapidly. As a result, frequent rope changes are required resulting in substantial cost and productive time loss. Although performance of rigging has a profound impact on dragline efficiency, there are no common, industry accepted standards related to its selection, operation and maintenance. Little information on the subject is available and the practice differs widely throughout the industry. The performance of drag and hoist ropes depends on: •rope construction that should closely match requirements of the specific application, rope operating conditions and its maintenance and • correct definition and application of the rope discard criteria so that full advantage of the rope can be taken without endangering safety of its operation. This paper reviews the present practice related to all these factors. Its findings are based on the wide survey of the present industrial practice. Field survey In total, 33 North American dragline applications were surveyed of which 17 were Bucyrus-Erie machines, 10 were Marion and six were Page (now P&H). Most of the draglines were stripping overburden in coal mines, but the survey also included some unique applications such as phosphate mining in Florida and oil sand mining in Alberta. Hoist drive power of the surveyed machines varied from 300 to 7775 kW (400 to 10,400 hp). Drag drives varied from 300 to 5815 kW (400 to 7800 hp). Boom lengths varied from 37 to 110 m (120 to 360 ft). Bucket sizes varied from 5 to 96 m3 (6 to 115 cu yd). The hoist and drag ropes installed on the surveyed machines were supplied by 10 rope manufacturers and differed widely in type and construction. Most common rope strand constructions included standard strands, Seale, Warrington, Warrington-Seale and Filler-Seale. Results of the survey [ ] Working conditions Working conditions of hoist and drag ropes are quite different. The first cope primarily with bending on rope drums, point sheaves and deflection sheaves, and with tension force imposed by the bucket and its load. The drag ropes are exposed to severe abrasion at the pit crest, and to severe bending and crushing at fairleads, all while coping with tension forces imposed by resistance of the bucket to dragging. Bending of drag ropes at house rollers and at rope drums is of minor importance for their performance. Differences in working conditions result in different wear patterns, which make it feasible to install discarded hoist ropes to further serve as drag ropes. The hoist ropes wear primarily due to fatigue breaks of wires, while the drag ropes are subject to severe abrasion of outside wires, rope deformation and internal wire-strand breaking. The last is often indicated by the reduction of rope diameter. The summary of working conditions for the surveyed draglines is shown in Table 1. For simplicity, Table 1 does not include the data on rope bending at the fairleads, nor at the deflection rollers and house rollers. Similarly, Table 1 does not include the data on drag rope bending (other than that on the rope drum), as its influence on rope life is insignificant in view of severe abrasion and internal wear taking place elsewhere. The following abbreviations are used in Table 1: [D - diameter of rope drum. DP - diameter of boom point sheaves. d - rope diameter x - mean value of the measured parameter. s - standard deviation of x B/d - average bending ratio of ropes B/d =(D+ DP)/2d n - operational safety factor of the rope (rope stength) x (number of ropes) n(bucket load + bucket weight)] In terms of D/d ratio, the working conditions of hoist and drag ropes are similar to that in other related industrial applications. However, the safety factors of the dragline ropes are lower, which requires them to carry relatively large tension forces. The last explains the fact that the life of dragline ropes is low when compared to other industrial rope

Jan 1, 1995

Coal waste materials inhibit direct vegetation establishment due to adverse physical and chemical properties, particularly low water retention and high potential acidity. The Moss # I coal refuse pile is located in Dickenson County, Virginia, and was idled in the late 1980&apos;s with little topsoil resource available for final closure. The refuse was acidic (Total-S = .. 38%; pH = 3.6), black, high (70%) in coarse fragments, and had a low water holding capacity (4.5% in < 2.0 mm fraction). A small plot experiment was established on the refuse pile to evaluate the influence of liming rates (50% and 100% of lime req.) and topsoil thickness (15, 30 and 60 cm) on vegetative growth and leachate quality. Liming and topsoil amendment increased the surface soil pH from < 4.0 to > 6.0 over a two-year period, which resulted in greater vegetative cover and biomass than the control plots. All topsoil treatments resulted in greater vegetative cover and biomass than plots treated with lime only due to improved surface soil physical and chemical properties. A topsoil treatment of 60 cm gave the thickest vegetative cover and biomass yield. Such a treatment, however, would be cost-prohibitive at this location. Application of 27 Mg ha-1 of lime to the refuse surface along with 15 cm of topsoil produced acceptable two-year vegetative cover and biomass, and appeared to be the optimal treatment for this particular situation. Both liming and topsoil had no affect on leachate pH and the electrical conductivity in leachates collected below the plots. This suggests that surface revegetation will have little effect on the quality of water draining through the pile, so long term water treatment requirements may not be reduced by successfully revegetating the pile surface.

Jan 1, 1999

By William G. Pariseau

Few departments of mining, metallurgical or petroleum engineering in the United States have critical mass. Cost per student credit hour (SCH) in departments across the country are usually the highest among engineering departments, which are the highest on campus. Almost all departments are operating at fixed costs and enrollments could easily double without incurring much additional cost. However, the cost per SCH would remain unacceptably high and the probability of doubling enrollment is nil. Declining enrollment is a more likely future. The pressure to reduce SCH from university administration is unrelenting. And it has the potential for doing serious damage to program quality when standards and requirements are lowered in attempts to maintain enrollments and increase retention.

Jan 1, 2002

By John Rushton

Surface mining, from ore extraction to processing, requires many types of equipment to obtain the final product. Mine equipment often includes shovels, drills and haul trucks. Ore processing equipment, including plants, mills and concentrators, are also part of surface mining operations. Surface mining operations emphasize production. Production is where the bills are paid and fortunes are made. However, there must be support to keep the equipment operational. To most production people, equipment maintenance is a necessary evil. Maintenance is not glamorous work. It is often neglected when praise is being passed around. It is easy to recognize higher production, greater capacities, faster cycle times or more loads hauled. But who cares if an engine is replaced in 12 hours instead of 48? The answer is, everyone should - from the hourly miner to the mine manager. Surface mining equipment needs constant maintenance work. It must be lubricated, components replaced, electrical problems repaired and other problems addressed. Repair work for any surface equipment takes time away from production. It costs money. Maintenance costs cannot be avoided. However, they can be minimized. Minewide communication should include operations cooperating with maintenance. Traditionally, operations resist surrendering equipment to maintenance if it can continue to produce. The rationale is, every load that can be hauled today is money in the bank. If the unit is taken down, the money is lost. This thinking is prevalent in many surface mining areas, including the mill. "Not on my shift," the operator thinks. "Let the next crew worry about it. The problem with this thinking is in probability and statistics. The longer equipment goes without maintenance work, the higher the probability of equipment failure. When the failure does occur, it is certain to cause the equipment to remain down longer than if the maintenance department had been allowed to take it down. Many mining operations have closed early in their lives because of poor maintenance practices. This has become the industry standard because substandard levels of performance are acceptable. Management believes it gets five to six hours of work from maintenance people during an eight-hour shift. Actually, the industry average is less than two hours of work per shift. Real productivity levels of less than one hour are not unusual. There are few exceptions. However, a daily work schedule (DWS) will increase productivity at many properties by 50% to 100%. A 200% improvement is not unusual. The DWS will work. Maintenance people are among the most intelligent and motivated people on the property. They deserve the chance to be productive. [ ] In preparing a DWS, management needs to understand two productivity principles - the productivity spread and the real maintenance workday.

Jan 1, 1997

By A. P. Massmann

Design for economical operation, minimum operating labor and minimum maintenance for labor and supplies are virtually inseparable. Each consideration has a vital bearing on the other. Let us assume that we have the design complete for the most economical operation with a minimum of operating personnel. This means, in addition, that we have the fewest number of operating units consistent with the product requirements for a team or metallurgical coal which we are called on to produce. The number of units will vary greatly, being influenced by design, washability characteristics, tonnage and degree of quality control necessary by reason of producing a product for metallurgical, steam or domestic use. The units involved most ordinarily for any type coal preparation plant will number as a minimum the following: 1. Hoppering R.O.M. Coal 2. Feeding

Jan 1, 1959

By V. I. Lakshmanan, Dan A. Mackie

Traditional vat leaching methods have had limited success, primarily due to the high cost of loading and unloading the vats. In general, such batch methods are labor-intensive and costly. Heap leaching has emerged as a viable, low- cost alternative, but low recoveries, normally in the 70% range, offer opportunities for improvement. Also increased regulatory guidlelines for effluent and spent heap leach pads disposal offer potential for alternative methods. The Mackie Vat Leaching Jig (MVLJ) has had its incubation at Ortech International, where bench scale through pilot scale testing continues. Design developments and testwork over the past two years indicate that the time may be right for field trials.

Jan 1, 1990

By S. Finsterle, R. C. Trautz

We performed numerical modeling studies to investigate water seepage into underground openings excavated in unsaturated fractured rock. Water seepage is a key factor affecting the performance of a nuclear waste repository such as the one proposed at Yucca Mountain, Nevada. The amount of water dripping into an underground opening depends on the connectivity and permeability of the fracture network as well as on the capillarity of individual fractures intersecting the opening. We developed a high-resolution numerical model of an unsaturated fracture network and examined the appropriateness of seepage predictions using a fracture-continuum model. Model calibrations to liquid-release tests performed at Yucca Mountain are also discussed.

Jan 1, 2000

By Michael S. Zhdanov

One of the major problems of mineral exploration is the ability to reliably distinguish between uneconomic mineral deposits and economic mineralization. While the mining industry uses many geophysical methods to locate the mineral deposits, until recently, there was no reliable technology for mineral resources identification and characterization. The main goal of this project is to develop a new geophysical technique for subsurface material characterization, mineral exploration and mineral discrimination, based on electromagnetic methods. The proposed new technique is based on examining the spectral induced polarization effects in electromagnetic data using the modern distributed acquisition systems and advanced methods of 3-D modeling and inversion. The recovered parameters of the relaxation model will be used for the discrimination of different rocks, and in this way will provide an ability to distinguish between uneconomic mineral deposits and zones of economic mineralization using geophysical remote sensing technology (Zonge, 1976). The proposed research and development is undertaken using multi-partner collaboration between the University of Utah, four major mining companies, Kennecott Exploration, BHP Billiton World Exploration Inc., Placer Dome Inc., Phelps Dodge Exploration, and Zonge Engineering Inc., which is equipment manufacturer and service provider to the mining industry. The multi-partner collaboration between the academic institution, the University of Utah, and mining industry helps to identify the appropriate mining target and conduct the field experiment for practical demonstration of the developed new technology, which is an important part of the this project.

Jan 1, 2005

By R. W. Fedors, S. T. Green, R. A. Hart, D. B. Walter

Computational fluid dynamic (CFD) modeling of a potential nuclear waste repository drift may be an important tool to estimate moisture redistribution caused by natural convection within the drift. To refine and support CFD techniques, a small laboratory cold-trap experiment was developed. Results from numerical modeling of the laboratory experiment show the sensitivity of the bulk air flow rate and average air temperature within the drift to several key parameters, including various boundary condition assumptions and surrounding porous media properties.

Jan 1, 2004

By John C. Braithwaite

Gold production commenced in New Zealand during the gold rush of 1861, and recovery to date is over 900 tonnes, almost 60% of which was alluvial. New Zealand pioneered development of bucket line dredges during the 1880&apos;s. The advent of the modern hydraulic excavator and increased gold price led to a resurgence in alluvial production with about one hundred plants in operation in 1988. The processing plants are mainly floating, with throughputs of 50-100m3/hr fed by backhoes of 20-40 tonne operating weight. Some larger plants with more sophisticated recovery systems have commenced over the last few years, with the largest using a 180 tonne backhoe.

Jan 1, 1989

"The adsorption of mixed cationic/anionic surfactants with different molar ratios onto quartz was investigated. The cationic surfactant used was dodecylamine bromide (DDA), and the anionic surfactant was sodium oleate (NaOL). The critical micelle concentration values for the individual surfactants and their corresponding mixtures were determined using a platinum plate. The adsorption mechanisms of mixed surfactants on quartz were studied using contact angle measurements, flotation tests and adsorption amount measurements. At alkaline condition, NaOL depresses DDA flotation of quartz without depressing the adsorption of DDA. The molar ratio of mixed surfactants is found to be a key factor in the orientation of alkyl chains and thereby the flotation response of quartz. A mixture of 1/2 or 1/3 DDA/NaOL forms on the surface of quartz when the concentration of NaOL is higher than that of DDA. The adsorption of the mixture decreases the wetting of quartz since the alkyl chains are in chaotically orientation with a conceivable number of head groups directing towards the solution phase. These results provide important insights into the adsorption of mixed surfactants onto quartz surface. INTRODUCTION Froth flotation is known as an effective method of separating mineral particles on the basis of their surface chemistry, specifically their hydrophobicity. [1]. The hydrophobic particles will attach to the bubbles and rise to the top of the containing vessel to form a froth that flows over the lip of the vessel. The hydrophilic particles remain in the liquid [2]. In the process, surfactant is often used as collectors for surface modification in order to change the wetting properties of mineral particles. The primary long chain alkyl ammonium salts are the most widely used as collectors for the flotation of oxide minerals, such as quartz, muscovite, hematite, phosphate and feldspar [3, 4]. Adsorption mechanism of alkyl ammonium salts on oxide minerals has been extensively analyzed [5, 6]. Generally, the ammonium salts are adsorbed through electrostatic interaction of the polar group of the molecules with the surface of these minerals [7]. Despite many obvious advantages of the ammonium salts in low temperature resistance and hard water resisting, however, their unavoidable drawbacks, such as high sensitivity and poor selectivity for slimes, relatively high cost, have limited their applications [8]. In recent years, mixed surfactants have drawn much attention due to their superior physicochemical and surface properties than single surfactants [1, 9-11]. They have been used in many industrial processes such as dispersion/flocculation, flotation, detergent, corrosion inhibition, cosmetics, foods, paints, chemical mechanical polishing, enhanced oil recovery, and nanolithography [12]. Most of the mixed surfactants have synergistic advantage in reduction of the surface tension and critical micelle concentration (CMC); therefore, they exhibit spurious properties, like excellent surface/interfacial activities, and are less expensive than single mixed surfactant systems [13, 14]."

Jan 1, 2016