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In a previous paper the author showed that silver considerably delayed the recrystallisation of lead after cold work. The effect of antimony on the rate of recrystallisation of distorted lead was studied next, and the experiments performed, and the results obtained, are described in this paper. Electrolytic lead was used as a standard of comparison.It was also used to make the lead antimony alloys.This electrolytic lead contained 0.0005% total impurities.It was prepared from commercial lead by fire refining, followed by electrolysis and finally by vacuum casting. The process was described in the previous paper and the description need not be repeated here. PREPARATION OF LEAD - ANTIMONY ALLOYS.The steps taken to procure pure antimony are outlined in the appendix. To make the lead antimony alloys a mother alloy was prepared thus:-A quantity of the antimony (it was received in the powdered form) was placed in a pyrex glass tube, which was closed at one end. The antimony was washed with hot hydrochloric acid (1: 1), then with more dilute acid, then with distilled water, and finally with absolute alcohol; it was dried by warming and evacu,ating the tube. This should leave the antimony more nearly free from surface oxides.The amount of antimony was determined by weighing quickly; the appropriate amount of vacuum-cast electrolytic *Research Fellow, Dept. of Metallurgy, University of Melbourne.

Jan 1, 1936

By W. N. Marchant

The Bureau of Mines performed chemical research to elaborate details of sulfur chemistry pertaining to the citrate flue gas desulfurization process in which sulfur dioxide (SO2), absorbed (as bisulfite ion) in a buffered sodium citrate solution, is reduced by hydrogen sulfide (H2S) according to the overall reaction 2H2S + HSO3 + H+ ? 3S + 3H20. The rate-limiting step in the process was shown to be the reduction by H2S of thiosulfate (S2031-) that is formed as an intermediate. Pseudo first-order kinetic measurements were made at room temperature to retard the reaction sufficiently that adequate analysis was possible. These measurements showed the rates of the following net processes to be -d [S02]/dt (pH 4.5) = (0.409±0.050)min-1 [S02]; d [S2032-]/dt (pH 4.5) = (0.199±0.017)min-1 [S02]; -d [S2032-]/dt (pH 4.0) = (0.0437±0.0265)min-1 [S2032-]~1; -dH2S/dt (pH 3.9) = (21.6±3.0) mmoles/min-atm [H2S]g1.4, where [H2S]g is the partial pressure of H2S in the gas phase, in atmospheres, and the rate describes the consumption of H2S, in millimoles per minute; d[S]/dt (pH 4.0) = (38.0±3.4)min-1 [H2S]1, where [H2S] is the concentration of H2S in solution. Hydrogen ion consumption during the slow phase of regeneration was shown to occur at (2.1±0.5) times he rate of thiosulfate consumption, which is interpreted as evidence for the following as the rate-limiting reaction: 2H2S + S2032- + 2H+ ? 4S + 3H20 A scheme consisting of five sequential general reactions is proposed to account for sulfur formation.

Jan 1, 1981

By J. -G. Swanson

Dust control in longwall coal mining depends primarily on the prevention of dust at the source and secondarily on the suppression of airborne dust. One of the primary methods for the prevention of dust is the wetting of the coal face prior to extraction, which can be achieved with water spray systems installed on the shield supports. The spray droplets must travel across the airway to the longwall face, where high ventilation air speeds can entrain water droplets and prevent them from reaching the face thus minimizing the dust control at the source. Through an evaluation of spray plume characteristics in a wind tunnel, the droplet size and droplet drift behavior was determined. This drift behavior was examined by evaluating the spray plume concentration drop and the droplet size shift near the nozzle exit at varying air speeds. Based on these evaluations, CFD modeling was applied to simulate the spray behavior for varying geometries. These results led to the selection of appropriate spray nozzles capable of carrying water droplets to the face at higher air speeds, resulting in an adequate wetting of the coal face and reduced dust generation.

Jan 1, 2012

By C. R. Kuzell

GEOGRAPHICAILY the industry of reducing and refining of copper continued to migrate from the .United States during 1931. While this country is losing the predominant position of its copper industry, our engineers still enjoy their pre-eminent position as leaders in the design and operation of the new plants being brought into production in the four quarters of the globe. Progress in the art, as indicated by the construction of the recently completed works, has not been revolutionary. Standard methods have been followed, augmented by novel features in flow-sheets demanded by the results of progress in the art of ore-dressing, and by reason of problems incidental to the peculiarities of the new ores being treated. Design and construction progress has likewise been orderly and without radical departure from less recent practice; new features in primary and secondary equipment being logical developments to provide for special conditions to keep abreast of improve¬ments in accessory apparatus. The new plants are being rapidly brought in without hazardous novelties and for the purpose of yielding the principal ingredient, copper, in as large a volume as it will he possible to market; the old United States procedure. In the meantime the domestic plants are shut down or operating at a very low capacity, threatened with untimely demise unless the metallurgist can extract and develop other valuable products from the ores. International agreements .and tariff protection may be a temporary reviving stimulant, but for many mines the days of profit will be numbered unless the ores can be made to yield as values the ingredients now being discarded. Invisible progress in this direction has been made during the past year, and it is to he hoped that time will prove that the outstanding progress of 1931 was in the development of processes for the ultimate recovery of by-products from copper ores, perhaps even relegating the value of the copper yield to secondary importance.

Jan 1, 1932

By Fred C. Schmednecht

For years, the strategy was to contain waste and in many applications, is still the recommended method. Recently, it has become acceptable to create filters below ground for the in-situ treatment of waste. Consequently, in lieu of the waste being contained, it is desirable that the waste flow through a pervious treatment zone. For instance, subterranean pervious walls comprised of iron sand are particularly useful in treating chlorinated solvents in contaminated groundwater. The iron acts as a corrective filter to treat the contaminated groundwater. How are these impervious and pervious walls being constructed? One effective means of constructing an impervious barrier is the vibrated beam method. This method, in brief, involves the use of a fabricated steel beam, a vibrator, a pile driving crane, and a mixing plant. This method results in a 4-6" wide wall with depths up to and exceeding 100'. Over one hundred slurry walls have been installed using this method in the United States. The need for the vibrated beam method continues to escalate due to its reduced health and safety considerations, high level of quality control, lack of excavation, narrow working area, and ability to achieve depths greater than 100'. A recent project was in Marinette, Wisconsin at the ANSUL facility. The project consisted of a 28,000 square foot slurry wall and was completed in January of 1999. Another effective means of creating an impervious barrier, where structural support from the barrier is needed, is the Waterloo® Barrier System. This system involves the use of a pile driving crane, Waterloo® steel sheet piles which are patented steel sheet piles due to their groutable interlocks, and a grout plant. This system results in a water-tight barrier due to the grouted interlocks and provides structural support. A project was completed at Beta Steel in Portage, IN for a retention pond to protect Lake Michigan. Due to the new concept of "filtering", there are few accepted construction techniques available. One effective field proven construction technique is the patent pending mandrel emplacement technique. This technique involves the use of a fabricated hollow steel mandrel with a shoe, a vibrator, a heavy duty crane, and a material feeder. This technique results in a pervious wall which may be comprised of pure iron sand or any other flowable media. Using this technique, the pervious wall may be constructed in varying widths to depths in excess of 100'. A successful installation occurred in October 1997 at Cape Canaveral Air Station in Florida.

Jan 1, 1999

By E. Tabosa, S. Vianna, K. Duffy, W. Valery, B. Andrade, L. Pyle, P. Holtham

Pneumatic flotation cells are high-intensity flotation machines with rapid kinetics, resulting in short residence times. Examples include the Jameson, PneuFlot and ImhoFlot cells. These cells do not use impeller-like mechanical cells. Rather, the air and pulp are mixed in a continuous stream of high fluid velocity through a downcomer or venturi to disperse the air into fine (approximately 0.5 mm) bubbles and maintain particle suspension. Froth wash water is often used to enhance concentrate grade, minimising recovery by entrainment. With the growing application of these types of cells, not only in cleaner stages (their traditional use) but also in pre-flotation and scalping duties, there is a need to predict their performance more accurately for circuit design and optimisation. No widely accepted model for pneumatic flotation cells is available and their metallurgical performance is often scaled up from laboratory and/or pilot plant testing by assuming fixed values for metal recovery, concentrate grade and concentrate percent solids. This paper describes a new approach to simulate pneumatic flotation cell performance and presents case studies demonstrating its applicability. Keywords: Pneumatic flotation, flotation modelling, flotation circuit design, flotation circuit optimisation, Jameson cell

Jan 1, 2020

By M. S. Bickers

Prior to 1970, the environmental liabilities associated with mine construction and management were few. Mostly, they were aimed at health and safety matters. At that time less than two dozen codes were incorporated into the engineering design and they were somewhat static and well known by the engineering designers and managers. These codes issued by governmental agencies and professional associations are listed in Table 1

Jan 1, 1988

By Mohammadreza Khanmohammadi, Kazem Fakharian

"AbstractIn driven piles, the gradual dissipation of excess pore water pressure and consequently the soil setup significantly contribute to increasing the pile capacity with time. The question is risen that how the scale differences affect the stress state variations around the pile shaft and at the pile tip? Would the CPT predictions better correlate with End-Of-Drive (EOD) or Beginning of Restrike (BOR) conditions? On the other hand, are the proposed a and ß parameters for clayey soils mostly predicting the EOD or BOR situation? The main objective of this paper is to back calculate the a and ß parameters from CPT profiles, pile Dynamic Load Tests (DLT) at the EOD and BOR as well as Static Load Tests (SLT) and evaluate how they compare with those suggested in literature. Four precast 400 mm square concrete piles driven in North Azadegan Oilfield in Khuzestan Province were selected for the comparisons. CPT test results were available from the extensive geotechnical investigation at the study site. DLT tests using PDA were carried out at EOD, 1 day and 14-day BOR on all the 4 piles. All the piles were SLT tested as well. The UniCone software was used to estimate the tip and shaft resistances using 4 methods including the Dutch, Eslami & Fellenius, LCPC and Schmertmann. The results indicate that the a-method based on the proposed values in literature has predicted a shaft capacity of 90% of the 14-day BOR. This is while the ß-method has predicted a shaft capacity equivalent to 1-day BOR results only. Therefore, both methods have under-predicted the shaft capacities compared to real capacities measured after the significant soil setup in the study area. Two of the CPT methods have underestimated, one has overestimated and one has matched the estimated 60-day capacity."

Jan 1, 2014

By T. Bhambhani, D. R. Nagaraj

"Ultramafic Ni ores present significant challenges in processing in that both recoveries and grades are not satisfactory, due to the presence of Mg silicates in large amounts. Significant quantities of xanthates and CMCs are generally used for their recovery by flotation. Additionally, given that some of these ores are processed under acidic conditions, CS2 generation due to xanthate decomposition presents some safety hazards. Technology to replace xanthate has been developed for which the safety and health issues are absent. Collector S-10294 shows significant improvements in Ni and overall S recovery at a dosing rate of one half that of xanthate. New polymeric modifiers have also been developed that help reduce the Mg content of the concentrates at very low dosages. The reagents are custom designed for use in normal or hypersaline water, and acid or alkaline conditions.INTRODUCTION The use of xanthates in sulfide mineral flotation is reported to have started as early as 1925 (Nagaraj 2005). The reagent is now ubiquitous in the mining industry, accounting for 40-50% of total collector consumption. In recent years, the development of more selective reagents has rendered xanthates largely obsolete in Cu flotation. However, xanthates continue to be sold in large amounts in the Au, PGM and Ni flotation industry, where bulk sulfide flotation is required. The low cost and easy availability is the largest driver for the use of xanthates in the minerals industry. Of particular interest to this study is the use of xanthates in the flotation of Ni sulfide ores. About 60% of the world’s annual nickel output is obtained from sulfide Ni ores (Hoatson, 2006). A strategically important category of these ores is the ultramafic hosted sulfides, such as those of Western Australia and Canada. These ores are characterized by low Ni grades (<1%), with the Ni minerals being pentlandite, violarite, millerite and mackinawite, and gangue minerals comprising an array of altered Mg silicates, mostly serpentines. The challenges associated with treating these ores are well documented (Eltham et al. 1973; Claridge et al. 1975). One challenge is that Ni recoveries from these ores are lower. This has been attributed to the increased pulp viscosity, as discussed in detail by Vasudevan et al., 2010. Additionally, Mg silicates tend to report to the concentrate in large quantities, which is problematic for two reasons: a) concentrate dilution, and b) the high melting point of Mg silicates tends accelerate wear on the refractory lining of smelters. This has been attributed to network transport to the froth phase via bubble flux and entrainment (Patra et al, 2010)."

Jan 1, 2012

By R. V. Peche

Since 1952, Bond’s equation has been used as a standard to predict the specific energy of grinding for a given duty. The Bond’s equation considers that the exponent is constant for all conditions. Hukki, in 1961, pointed out that the exponent is a function of the particle size, and concluded that the Bond’s equation applies successfully for an intermediate range of sizes; out of this range the Bond’s equation leads to high error as demonstrated in this paper. More recently, in 2004, S. Morrell presented a new formula in which the exponents vary with the feed and product size. This paper is based on Hukki’s approach and presents a new general formula, in which the exponent varies only with the feed particle size. This new equation has been proven in 82 observable cases in Peru and other countries around the world, involving ball mills and vertimills in different configurations over a wide range of feed and product sizes, work indices and mill diameters; in most cases, the predicted value has an error less than 10% with regards to the observed value and no need for any correction factor. Keywords: Grinding, specific energy, particle size, work index, variable exponent

Jan 1, 2020

Kidd Mine in Northern Ontario is now producing from Stage 1 of the Mine D expansion project ? mining between 2100 m and 2700 m below surface. Planning for Stage 2 of the Mine D project is already well advanced and will see the mine extended potentially to 3100 m below the surface. To ensure adequate ventilation is maintained at such depths, two large axial fans have been installed on the 6000 Level (1800 m below surface) as exhaust booster fans to operate in series with the surface exhaust fans. The Kidd Mine 6000 Level Booster Fans are a unique installation with unique and often unexpected challenges. These challenges have been met and attacked through the combined efforts of mine and fan manufacturer personnel with a common focus.

Jan 1, 2009

By Christopher Mark

Multiple seam interactions are a major ground control hazard in many U.S. underground coal mines. The two most common types are: ? Undermining, where stress concentrations caused by previous full extraction in an overlying seam is the primary concern, and; ? Overmining, where previous full extraction in an underlying seam can result in stress concentrations and rock damage from subsidence. The National Institute for Occupational Safety and Health (NIOSH) has completed a major study aimed at helping to identify the location and likely severity of these interactions. In the course of field visits to mines throughout the U.S., more than 300 multiple seam case histories were assembled into the largest data base of multiple seam case histories ever collected. These data were analyzed with the multivariate statistical technique of logistic regression. The study also employed LaM2D to estimate the multiple seam stresses, ALPS and ARMPS to determine pillar stability factors, and the CMRR to measure roof quality. The study resulted in the development of a computer program, called Analysis of Multiple Seam Stability (AMSS), which can help mine planners to evaluate each potential interaction and take steps to reduce the risk of ground control failure. AMSS first evaluates pillar design by calculating the single seam Stability Factor (SFSS) using ALPS or ARMPS. It also automatically generates a LaM2D analysis that provides the additional multiple seam stress so that the final, multiple seam SFMS can be determined. The second part of the AMSS procedure builds upon the statistical findings that overmining is much more difficult than undermining, isolated remnant pillars cause more problems than gob-solid boundaries, and weaker roof significantly increases the risk of multiple seam interactions. AMSS quantifies these effects and predicts the outcome in terms of three levels of risk: GREEN (where a major multiple seam interaction is considered unlikely), YELLOW (where adding a pattern of cable bolts or other equivalent supplemental support could greatly reduce the probability of a major interaction.), or RED (a major interaction should be considered likely, and it may be desirable to avoid the area entirely).

Jan 1, 2007

By Ma Fangyuan, Li Min, Xu Bing, Xian Yushuai, Zhao Tonglin

The study explored a new combination amine collector (XK-28) that can obtain better flotation index and defoaming rate for magnetic iron ore. A pure quartz mineral test indicates that dodecylamine, XK-I, or XK-II has good collection effect on quartz and their best flotation pH is in the range of 6–7. The experimental results of a combination collector (XK-I:XK-II = 2:8.) show that the separation effect for quartz and magnetite is better than that of a single collector. The single factor for actual mineral test shows that when the combined reagent is used as the collector, the best conditions are pH = 7, 90 g/t dosage, and XK-I:XK-II = 2:8, which is consistant with the pure mineral test. The difference for dodecylamine as a collector is that the optimal dodecylamine dosage is 60 g/t. Under the best conditions of actual ore test, dodecylamine and XK-28 are respectively used as the collector to compare the flotation indexes. Study results indicate the magnetite grade using XK-28 as the collector is higher by 3.02% than that of using dodecylamine, recovery is increased by 1.92% over that of using dodecylamine. Based on the measurement results of zeta potential, uses of XK-28 can slightly increase the zeta potential difference of quartz and magnetite to enhance the flotation effect. The mechanism of action between the XK-28 collector and the actual ore from the Gongchangling plant is physical adsorption, and this adsorption also has hydrogen bond formation. Contact angle results shows that XK-28 improves the hydrophobicity of quartz better than dodecylamine. Compared with dodecylamine, XK-28 is more sparsely arranged on the surface of the foam. The foam produced by the XK-28 collector has better air permeability and it is more likely to merge and disappear with bubbles.

Jul 21, 2021

By Wilton Johnson

The Bureau of Mines study indicates that land utilized by the mining industry for mineral extraction and processing from 1930 through 1980 amounted to 5.7 million acres, or 0.25 percent of the land mass in the United States. Land reclaimed by the industry during the same period was 2.7 million acres, or 47 percent of the land utilized. Land use by the mining industry includes surface land used for excavation or mining, for waste from underground mining, and for wastes from milling or processing operations. Data on land use for disposal of overburden waste from surface coal mines and on areas affected by subsidence associated with underground mining are included for the period 1930-71 only. The 10 leading States in total land used for mining over the 51-year period were, in decreasing order, Pennsylvania, Kentucky, Ohio, Illinois, West Virginia, Florida, Indiana, California, Alabama, and Missouri. Seven mineral commodities accounting for 92 percent of land use are, in decreasing order, bituminous coal, sand and gravel, stone, phosphate rock, clay, copper, and iron ore, all predominately surface mined. Bituminous coal production was responsible for nearly half of the total area utilized. Reclamation of mined lands and waste disposal sites over the reporting period was largely on lands utilized in mining bituminous coal; 75 percent of the land used for bituminous coal production was reclaimed.

Jan 1, 1981

"The McArthur River uranium deposit is the single richest uranium ore body developed in the world to date. The ore body’s average grade is 15% U3O8. Proven reserves are in excess of 450 million pounds. The grinding circuit is located underground. The ground ore is pumped to a surface facility where the slurry is thickened and loaded into containers for shipping to the Key Lake mill 78 kilometers to the south.The paper will briefly describe the mining method employed. The underground ore handling and grinding circuit, and surface facilities including slurry load out, wastewater treatment and the freeze plant will be described in detail. The paper will emphasize radiation and safety procedures that are followed to minimize operator exposure to radiation risks.LOCATIONThe McArthur River mine is located 78 kilometres north of the Key Lake mill. It is located at latitude 57’ N in northern Saskatchewan, within the southern boundary of the Athabasca formation. The nearest town is Pinehouse, located 320 kilometres to the south. Access to the mine is by an all weather road and by a private airstrip.HISTORYThe McArthur River Operation is jointly owned by Cameco Corporation (69.805%) and Cogema Resources Inc. (30.195%). Exploration for uranium in the McArthur River area started in 1980 by the Saskatchewan Mining Development Corporation, a predecessor company of Cameco.In 1984, a 12-kilometre long basement conductor was identified at a depth exceeding 500 metres. Systematic exploration drilling of this conductor resulted in the discovery of the P2 North deposit. Subsequent surface drilling programs from 1989 to 1992 delineated a narrow ore zone over a strike length of 1700 meters at a depth of 500 to 600 metres. At this time, geological reserves for the deposit were estimated to be 2.4 million tonnes at a grade of 5.0% U3O8 for a uranium content of 260 million pounds U308"

Jan 1, 2001

By L. B. Davidson

The petroleum literature contains many reports on the relative permeability properties of porous media. However, only recently have studies of relative permeability at elevated temperatures been made.1 -5 Each of these studies presents relative permeability data on water-oil systems in linear sandpacks or sandstone cores under isothermal conditions. The results of these investigations are dissimilar. Edmondsonl found that in Berea sandstone a temperature-dependent effect was present, while Shilobod,2 using the same material, found no temperature dependence after normalizing for initial water saturation. Workers at Texas A&M U.3-5 found that the ratio kw/ko was temperature dependent in an unconsolidated system. The differences in results of these studies are probably a consequence of differences in the material employed, in the experimental procedures followed, and in the properties of the porous systems used. This paper describes an experimental investigation of permeability ratio temperature dependence, conducted to clarify some of the results obtained by previous workers. Each experiment described in this paper involved the isothermal displacement of Chevron No. 15 white oil by either nitrogen, steam, or distilled water and displacement of water by nitrogen. All permeability ratios were calculated using the Welge procedure. Values of the water-oil permeability ratio were obtained over the temperature range 75OF to 540°F. Results indicate the ratio is temperature dependent at low water saturations, probably due to interfacial effects. However, as the saturation is increased, temperature dependence decreases until, above some minimum saturation, the ratio becomes insensitive to temperature. Previously published work indicates that the dependence reappears at very high water saturations, a region of the ratio curve not completely examined in the present study. It appears that water-oil permeability ratios are temperature dependent at the extremes of the ratio curve where changes in interfacial tension with temperature induce variations in residual saturations. Between the extremes, there is a region in which the permeability ratio is insensitive to temperature. The size and location of this region is probably a function of the characteristics of the porous system. Gas-oil permeability ratios calculated from nitrogen-white oil displacement experiments indicate a definite temperature dependence over the range 75OF to 500°F and over the entire gas saturation range studied. Evidence discussed in this report indicates that this dependence is a consequence of molecular slippage in the gas phase. This is an extension of the Klinkenberg effect6 to two-phase systems at high temperatures (the existence of slippage in two-phase systems at room temperature has been demonstrated7). Evidence for a similar source of temperature dependence in superheated steam-white oil experiments

Jan 1, 1970

By R. J. Matetic, David S. Yantek, David K. Ingram

The National Institute for Occupational Safety and Health (NIOSH) has investigated engineering noise controls to reduce sound levels in cabs on air-rotary drill rigs. A recent investigation revealed that some drillers are exposed to A-weighted sound levels exceeding 85 dB even though a cab is used. NIOSH studied the in-cab sound levels of one such rig. First, preliminary tests were conducted in a controlled environment using accelerometers and microphones with spectral analysis to identify the dominant noise sources for in-cab sound levels. The results indicate that vibration transmitted from multiple hydraulic pumps to the control panel produces a dominant spike in the sound level spectrum in the 400 Hz 1/3-octave band. Next, field tests were performed in a production environment to evaluate noise controls to reduce in-cab sound levels. It was found that utilizing hydraulic noise suppressors reduces the structure-borne noise transmitted to the control panel. Further, using hydraulic noise suppressors and enhancing soundproofing reduced the in-cab A-weighted sound levels by as much as 4 dB.

By Peter Broad

"Extracting any metal, especially specialty metals and rare earth elements, requires an in-depth analysis of the costs and potential revenue as well as a due diligence study of other stakeholder interests. Rare earth projects are generally in the billion dollar US range or above, and thus failure to accurately predict any loss of revenue caused by impurities within the product or additional effluent treatment costs can destroy an otherwise viable study. Two prime sources of input impurity are considered: impurities in the material feed and from reagents. The effect of these impurities factor on both the product (in reduced revenue) and on residual “tailings” that may incur additional unforeseen costs.INTRODUCTION“Truth passes through three stages. First, it is ridiculed. Second, it is violently opposed. Third, it is accepted as being self-evident” – Arthur SchopenhauerThe era of unscoped, unbudgeted construction ended about 150 years ago. Today’s engineer must design with confidence; but do so amidst proprietary secrets and unfamiliar pollutants, which present a myriad of interesting challenges. Is it possible to build a viable rare earth plant to a realistic budget, without destroying that which we cherish? Some see this as contradictory but in reality it is the only viable option.WHEN GOOD PROJECTS FAILWithout rare earths the Copenhagen agreement might be null and void. Running a wind turbine with a 30 tonne conventional magnet requires some hefty engineering and prohibitive investment. Despite their ‘Green metal’ applications, production of rare earths have created environmental nightmares, as Folger stated in 2011 when visiting Baotou, or has Jaing Mainheng who has been quoted by Ambrose Evans-Pritchard 6 Jan 2013. This comment is not a criticism of China; the details are from a 2012 white paper published by the Chinese Government that initiated a $6 billion environmental clean-up. Nor is China unique; other plants have had problems as well. There were toxic leaks in the 1990s from Mountain Pass, and the Mitsubishi Bukit Merah Plant (built in Malaysia in 1958) required a $100 million clean-up."

Jan 1, 2016

By Pierre Lacoste-Bouchet, Ian Orford, Mike Cooper

"The Brunswick Mine ore requires fine grinding to achieve liberation of minerals. With 25mm slugs as the grinding media in the ball mill, there was a limitation of a P80 of 25 microns on the CuPb regrind circuit. Results from pilot plant testing showed that the separation efficiency could be increased through fine grinding with a P80 of 15 microns.Options considered were the capital expenditure for new fine grinding equipment such as a vertimill or the use of the present ball mill equipment with smaller media. The Wheelabrator Allevard Enterprise (WAE) millpebs product was chosen because it was a smaller media with a distribution of steel pellets between 5 to 12 mm.With millpebs the grinding product was reduced in size from a P80 of 25 microns to a P80 of less than 19 microns. The power draw to achieve the finer grind was 13% lower while the media consumption was 57% higher.The millpebs grinding media was also tested on the zinc regrind mill. The purpose of the trial was to understand what the media consumption and power draw would be with the charge lowered to achieve an equivalent grind as obtained with the slugs. The media consumption was approximately the same and the power draw was reduced by 30%.INTRODUCTIONThe Brunswick Mine ore is a complex Pb-Zn-Cu-Ag sulphide ore with the grain size of the minerals varying from submicron blebs to particles up to several mm in size. The ore is ground progressively finer in stages. The first step of grinding occurs in the SAG mill circuit. The product sizing 80% passing 95 microns is fed to the CuPb prefloat stage (Cormier et al 2002). The tails from the prefloat is then further ground to a P80 of 35 microns for the CuPb rougher flotation. The final liberation for the CuPb circuit takes place in the CuPb regrind circuit. The regrinding mill further liberates the cleaner scavenger concentrate, second cleaner tails and third cleaner tails. The reasoning for the progressive grinding is to achieve fine grinding without having galena over ground due to its high specific gravity that preferentially sends it to the cyclone underflow."

Jan 1, 2006

By K. M. Phamotse, A. S. Nhleko

"*Paper written on project work carried out in partial fulfilment of BSc (Mining Engineering) degreeSynopsisLiqhobong Mining Development Company (LMDC) has been experiencing problems with boulders after blasting where the fragment sizes exceed the maximum of 800 mm as per mine standard. As a result, the mine has employed various methods to improve the fragmentation. The goal is to produce a run-of-mine (ROM) feed that does not choke the crusher and cause delays in production. In order to achieve this goal, fragmentation distribution within the fines and coarse envelope must be optimized through effective planning of blasting activities and accurate execution.The mine determined the fines-coarse envelope within which the entire crushing system can handle fragments using Split Desktop software. It is expected that both the predicted and actual fragmentation curves lie within that envelope for optimal fragmentation. The Kuz-Ram model with blast design parameters of 2.6 m for burden, 2.8 m for spacing, and 127 mm hole diameter was used to predict the fragmentation. The results show that the blast design parameters may need altering to achieve optimum fragmentation. Furthermore, the execution of the drilling and blasting may be the cause of the fragmentation problems. The mean fragmentation size (X50) differs greatly, unlike the uniformity index (n)s values which are relatively close to each other (0.6 to 2.2). The mean squared error (MSE) values have a large range. A proposed solution is a modified burden, spacing, and hole diameter. It is concluded that blast design parameters need to be reviewed in order to obtain correct predictions. IntroductionFragmentation has an impact on the mine production cycle which comprises, inter alia, drilling, blasting, loading, hauling, and crushing. Correct fragmentation as set by the plant design is important as it reduces the time lost due to secondary blasting and/or loading difficulties. In this paper we analyse blast results at Liqhobong Mining Development Company (LMDC) in Lesotho for a period of six months. This project was undertaken to observe the trends from the blasts and to identify whether any discrepancies exist in the overall blast information. The period under study included 16 blasts."

Jul 1, 2019