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By Dianne C. Marozas

The U.S. Bureau of Mines is conducting research to evaluate the potential for developing an in situ leach mining process that will selectively extract manganese from oxide ores, leaving gangue relatively undisturbed. To accomplish this, a geochemical approach, based on deposit-specific information such as mineralogy and geology, is used to constrain the operating parameters of the system. Among the studies being conducted are (1) leaching experiments to define reaction mechanisms and kinetics, (2) detailed petrologic characterization at the microscopic scale to assess Mn distribution and leaching efficiencies, (3) deposit characterization and evaluation, and (4) development of potential Mn recovery schemes. Data from these studies are being used to develop engineering design information for an in situ mining process. Specific details of the Bureau's research project are discussed, including results obtained to date.

Jan 1, 1991

By William Petruk, J. R. Schnarr

The behaviour of minerals during flotation of a fine-grained pyrite-rich base metal ore was investigated by studying the feeds. concentrate and tailings from every flotation cell in the mill of Brunswick Mining and Smelting Corp. Ltd. Bathurst. New Brunswick. The samples were analyzed for Zn, Pb, Cu. Fe, Ag. Sb, Sn, Bi, In. Hg, Mo and As. Mass balances were computed and recoveries of all metals from each cell. from each circuit and from the mill were calculated. The percentages of sphalerite. galena and chalcopyrite occurring as free grains in each product were determined with an image analyser, as were size analyses for free mineral grains and middling particles, and compared with screen analyses. The results show that the behaviour of minerals is defined much better by the recovery of free mineral grains than by the recovery of elements. The recovery of free sphalerite in the zinc circuit is 96 wt %, whereas Zn recovery from the same circuit is 88 wt %. the difference being due to unliberated sphalerite. The recovery of free sphalerite from the mill feed is 88 wt % and Zn 78 wt %. The recovery of free galenafrom the mill feed is 88 wt % and Pb 64 wt %. The recovery of free chalcopyrite is 75 wt % and Cu 45 wt %. Size analyses show that more sphalerite and chalcopyrite can be liberated by regrinding. but the galena is so fine grained that regrinding would not necessarily improve liberation. Silver occurs mainly in tetrahedrite and galena; the main loss in the tailings is due to large (+ 26 um) and small (-1 um} free and unliberated tetrahedrite grains.

Jan 1, 1981

By J. F. Burr, D. Snyder, W. Calder

This paper provides a determination of the equivalent level of protection of the international standards relative to similar criteria used by the U.S. Mine Safety and Health Administration (MSHA) to approve two-fault intrinsically safe (IS) stand-alone equipment. U.S. mining law requires such a determination for MSHA to use alternatives to existing standards. The primary issue is to demonstrate that the international standards for equipment evaluation will provide at least the same level of protection for miners as the document currently used by MSHA.

Jan 1, 2017

By J. F. Burr, D. Snyder, W. Calder

This paper provides a determination of the equivalent level of protection of the international standards relative to similar criteria used by the U.S. Mine Safety and Health Administration (MSHA) to approve two-fault intrinsically safe (IS) stand-alone equipment. U.S. mining law requires such a determination for MSHA to use alternatives to existing standards. The primary issue is to demonstrate that the international standards for equipment evaluation will provide at least the same level of protection for miners as the document currently used by MSHA.

Jan 1, 2018

By J. F. Burr, D. Snyder, W. Calder

"INTRODUCTION “Explosion protection” refers to techniques used to minimize the potential for electrical and electronic equipment to create an ignition while operating in a hazardous location (HAZLOC). In particular, U.S. coal mines are required to use equipment in certain areas of the mine that has been approved by the Mine Safety and Health Administration (MSHA) for use in a methane and coal dust environment to limit the risk of the equipment creating an ignition-capable spark or a thermal energy ignition. In general, MSHA’s regulations for explosion protection recognize two techniques: the use of explosion-proof enclosures (XP boxes) and 2-fault intrinsic safety for electrical and electronic equipment. For 2-fault intrinsically safe (IS) equipment, the U.S. mining industry requirements are unique in that the equipment is certified by MSHA. The criteria MSHA uses to evaluate equipment for intrinsic safety, as required by Title 30 of the Code of Federal Regulations (30 CFR) [30 CFR, Sections18.20(b), 18.68(a)(1), 19.1(b), 20.1(c)(2), 22.6, 23.6, and 27.20(a)], is published in a document referred to as ACRI2001 (1). Outside of U.S. underground mining, many industries and countries accept equipment that is designed to a consensus-based international standard (ANSI/ISA 60079-11 is the U.S. version). The purpose of this study is to provide an overall assessment of the ANSI/ISA 60079-11 standard on 2-fault intrinsic safety and the MSHA ACRI2001 acceptance criteria, and determine if the ANSI/ISA document can be an alternative to ACRI2001 while maintaining an equivalent or better level of safety for miners. There are many drivers for this study. First, immediately after the 2006 MINER Act (2) was passed, there was a perception in the industry that the unique MSHA requirements were contributing to delays or restricting the availability of equipment to meet the requirements of the Act. Additionally, the MINER Act required that mine operators submit an accident response plan for MSHA approval, to be reviewed every six months. MSHA is required to determine if these plans reflect the most recent credible scientific research and make use of currently commercially available technology. There is a concern, as expressed in a recent National Academy of Science (NAS) study (3), that the unique requirements for U.S. mining may be limiting the technologies available to coal miners and technology options for mine operators. Accordingly, the study recommended that the National Institute for Occupational Safety and Health (NIOSH) and MSHA re-examine their technology approval and certification processes to ensure they are not deterring innovation, and explore opportunities to cooperate with other international approval organizations to harmonize U.S. and international standards without compromising safety. Lastly, the National Technology Transfer Act (NTTA) and the accompanying Office of Management and Budget’s OMB-119 circular mandate the use of consensus-based standards by federal agencies. However, for U.S. mining operations, MSHA cannot accept such standards unless a determination is made that such standards provide an equivalent level of protection for the miner, per the requirements of the 1977 Mine Act, the 2006 MINER Act, and 30 CFR Part 18 regulations."

Jan 1, 2017

By D. P. Schlick

As of June 30, 1970, Federal regulations required the operator of each underground coal mine to maintain the average concentration of respirable dust below legislative standards. As of March 1, 1971, over 152,000 respirable dust samples have been processed by the Bureau of Mines. Approximately 75 percent of underground working sections are being sampled in accordance with Federal regulations. Results indicate that 72 percent of the sections which have completed at least one basic sampling cycle are below 3.0 mg/m3, the standard in effect under the Federal Coal Mine Health and Safety Act of 1969 (P.L. 91-173). Notices of Violation were issued on 1,670 underground mines for failure to begin a sampling program; Withdrawal Orders were issued in 64 of these instances. Notices of Violation were issued on 1,121 sections where applicable dust levels were exceeded; Withdrawal Orders were issued in seven of these instances. There are 381 mining sections currently operating on permit; from the interim Compliance Panel.

Jan 1, 1971

By John Wilson

"The Rietveld numerical method was used to calculate the relative abundances of phases in both synthetic binary mixtures and synthetic and natural ores, using X-ray diffraction data. The results were compared with those of image analysis and a chemical method. The ranges of the relative errors were estimated. The Rietveld method is found to be acceptable for determining the abundances of major and minor phases and the amorphous component of partially crystalline samples. In complex mixtures, where there are several trace phases (~<2 wt%) present, the technique becomes less accurate. It is most useful as an alternative technique to image analysis when sample preparation is problematic or sample particle sizes are sub-micron. It is a manual method which requires user input during the fitting process, but is both simpler and faster than image analysis due to the reduced set up time and sample preparation required. It is now implemented at CANMET as a technique for quantitative analysis.INTRODUCTIONUntil recently, the Applied Mineralogy group at CANMET used only image analysis and chemical techniques to determine phase quantities. An effort to develop quantitative X-ray diffraction (XRD) has been on-going for several years. Initially, the reference intensity ratio (RIR) method with external standards was tried with some success (Szymanski and Petruk, 1989). The main difficulty with this technique was finding adequate standard material for each phase present in the sample. This shortcoming made the method ultimately impractical, given the wide variety of materials which come to the laboratory for analysis.The Rietveld refinement method of quantitative XRD (Hill et al 1993, Mumme et al 1996), in contrast, requires no standards and requires minimal sample preparation and therefore a more rapid turnaround compared to image analysis. The Rietveld method also has advantages over image analysis in that it can identify and quantify polymorphs, ultra fine materials and many types of hydroxide products. As well, it is useful for analyzing materials such as bauxite or talc, which do not polish well enough for making mounts of the quality required for image analysis. To quantify amorphous material and phases which are present in amounts below the detection limit of XRD, the addition of a standard material is still required. However, in such cases only one standard is needed, which is alumina."

Jan 1, 2003

By John F. Gamble

Amphiboles are hydrated mineral silicates five of which occur in asbestiform habits as asbestos grunerite (amosite) asbestos, riebeckite (crocidolite) asbestos, anthophyllite asbestos, tremolite asbestos and actinolite asbestos] and non-asbestiform habits (grunerite, riebeckite, anthophyllite, tremolite and actinolite). The asbestiform varieties are characterized by long, thin fibers while non-asbestiform varieties such as cleavage fragments form short fibers with larger widths. The U.S. regulatory method for counting asbestos fibers (aspect ratio [>]3:1, length [>]5 µm) does not distinguish between asbestos and cleavage fragments. The method biases toward increased counts of non-asbestiform cleavage fragments compared to long, thin asbestos fibers. One consequence of this regulatory approach is that workers can be erroneously classified as exposed to concentrations of asbestos (asbestiform amphiboles) above the U.S. 0.1 f/mL exposure standard when in fact they are not exposed to asbestos at all but non-asbestiform amphibole cleavage fragments. Another consequence is that the known carcinogenic effects of asbestos may be falsely attributed to non-asbestiform amphibole cleavage fragments of the same mineral. The purpose of this review is to assess whether amphibole cleavage fragments pose the same risk of lung cancer and mesothelioma characteristic of amphibole asbestos fibers. We identified three groups of workers exposed to non-asbestiform amphiboles: two groups exposed to grunerite (Homestake gold miners and taconite miners) and one group exposed to industrial talc containing non-asbestiform tremolite and anthophyllite in St. Lawrence County, NY. In addition to assessing strength of association and exposure?response trends in the non-asbestiform amphibole cohorts, comparisons were also made with cohorts exposed to the asbestiform counterpart (positive control) and cohorts exposed to the mineral (e.g. talc) that does not contain amphiboles (negative controls). The cohorts exposed to non-asbestiform amphiboles had no excesses of lung cancer or mesothelioma. Similar results were observed in the negative control groups, in stark contrast to the excess risks of asbestos-related disease found in the asbestos cohorts. The only possible exception is the twofold increased risk of lung cancer where exposure was to industrial talc containing cleavage fragments of tremolite and anthophyllite. However, this risk is not considered attributable to the talc or amphibole cleavage fragments for several reasons. A similar increased risk of lung cancer was found in Vermont talc workers, studied in the same time period. Their exposure was to relatively pure talc. There was no relationship between lung cancer mortality and exposure measured as mg/m3 years and years worked. A case?control study reported that all the lung cancer cases were smokers (or former smokers) and attributed the excess to smoking. There were two mesothelioma cases among the NY State talc workers exposed to cleavage fragments of tremolite and anthophyllite, but talc is not a plausible cause because of too short latency and potential for previous asbestos exposure. The positive controls of tremolite asbestos and anthophyllite asbestos exposed workers showed excess risks of both lung cancer and mesothelioma and positive exposure? response trends. St. Lawrence, NY talc does not produce mesotheliomas in animals while amphibole asbestos does. In sum, the weight of evidence fully supports a conclusion that non-asbestiform amphiboles do not increase the risk of lung cancer or mesothelioma. ©2008 Published by Elsevier Inc. Keywords: Amphiboles; Cleavage fragments; Lung cancer; Mesothelioma; Asbestos; Non-asbestiform amphiboles; Grunerite; Talc

Jan 1, 2007

By C. F. Strydom

Paper written on project work carried out in partial fulfilment of B. Eng (Mining Engineering) degree The Self-directed work team concept is the product of relentless performance pressure on American industries during the late eighties and early nineties from other world industrial leaders like Japan. The concept proved to be a successful vehicle by which the Americans could increase their production and improve their safety statistics. The aim of this study is to evaluate the concept as a means to improve the overall performance on South African goldmines. he study is directed by a historical research section, answering questions like: ?Why do we need to change? ?What is a Self-directed work team? ?What are the principle advantages of the concept? ?What are the essential requirements for the concept to succeed? The study also includes an empirical research section that focuses on the current Self-directed work team situation in the South African gold mining industry. The empirical research section is limited to three case studies within Anglogold due to a study period of only six weeks. The mines chosen for the case studies have been labelled Mine 1, Mine 2 and Mine 3. The latter being the base mine allowing for a more in depth study. This includes a perception analysis towards Self-directed work teams within the mine. The study proves the truth of the Latin proverb ?Tempora mutantur, nos et mutamur in illis? meaning ?The times are changing, and we are changing with them?. The chief change catalysts identified are a dwindling gold price and a projected labour crisis. From the study it is evident that the Self-directed work team concept disposes of the critical parameters required for a dramatic improvement in performance. The study also indicates that for the concept to succeed, certain factors like top-level commitment and management-employee trust have to be in place. Sadly, the concluding part of the study proves that very few of the identified critical success parameters are present in the South African gold mining industry. This can serve to explain why the industry only has success with the concept when monetary gain is the chief motivator. The study only focuses on the reasons why the concept is not performing to its full potential and does not wander into explaining the reasons for the current unsupportive milieu in the industry. Self-directed work teams could well rescue the South Africangold mining industry but currently we are driving a square pin into a round hole! [Terminology SDWT?Self-directed work teams AEL?Adventure Experiential Learning ISW?In-stope worker POLC?Plan, organize, lead and control TEC?Total employee costed Silo?The confinement of departments preventing networking JCI?Johannesburg Consolidated Investment Co. Ltd]

Jan 1, 2002

By S. D. Irons

With the passage of recent national and state legislation aimed at tighter regulation of air pollution and smoke control, attention has been focused on a problem which has long confronted the coal industry--the disposal of coal waste with its inadvertent ignition. Even as recent as 1935, a decision from the Allegheny County Court of Common Pleas confirms the stand taken on control of burning coal refuse disposal areas. "The uncontroverted testimony is that there is no feasible method of operating a coal mine without a gob pile on the surface, as no use has ever been found for this troublesome by-product of mining; that every large coal mine has a large gob pile close to its tipple; that sooner or later these piles all ignite through spontaneous combustion; that practically every large mine in Western Pennsylvania has a burning gob pile, and that there is no known means of averting such a fire." 19* As a result of the past "no regulation policy," there are presently on the order of 500 burning refuse banks, of which 213, 142, and 49 are found in West Virginia, Pennsylvania, and Kentucky, respective¬ly. 18 Many of these refuse banks have burned for 40 years or more; many others may continue burning for a like period. It is estimated that the annual emission from these smoldering piles approaches 183,000 tons of pollutents which may produce many deleterious effects on the populace, property and vegetation in the proximity to the sources.16 To date, little progress has been made in alleviating the pollution from the currently burning refuse banks; however, numerous investigators who are researching the causes of ignition, as well as preventative disposal techniques (refuse compaction, layering and crushing to finer sizes) have met with some success in handling the newly discarded coal refuse.

Jan 1, 1967

Pillars with width to height ratios of less than 1.0 are frequently created in underground hard rock mines. The strength of slender pillars can be estimated using empirically developed equations. However, the equations can provide variable results when the width-to-height ratios approach 0.5. This paper investigates some of the issues affecting pillar strength at low width-to-height ratios in hard brittle rock. The investigation includes an evaluation of empirical pillar strength data presented in the literature and observations of pillar performance in underground limestone mines in the eastern United States, supplemented by numerical modeling in which failure processes and sensitivity of slender pillars to variations in rock mass properties are evaluated. The results showed that the strength of slender pillars is more variable than that of wider pillars. The numerical model results demonstrated the increasing role of brittle rock failure in slender pillar strength. The absence of confinement in slender pillars can result in a fully brittle failure process, while wider pillars fail in a combined brittle and shearing mode. The onset of spalling in slender pillars occurs at or near the ultimate strength, while this is not the case for wider pillars. Slender pillars are shown to be more sensitive to the presence of discontinuities than wider pillars, which can partly explain the increased variability of slender pillar strength. Two examples are presented, which illustrate failure initiation by brittle spalling and the sensitivity of slender pillars to the presence of discontinuities.

Pillars with width to height ratios of less than 1.0 are frequently created in underground hard rock mines. The strength of slender pillars can be estimated using empirically developed equations. However, the equations can provide variable results when the width-to-height ratios approach 0.5. This paper investigates some of the issues affecting pillar strength at low width-to-height ratios in hard brittle rock. The investigation includes an evaluation of empirical pillar strength data presented in the literature and observations of pillar performance in underground limestone mines in the eastern United States, supplemented by numerical modeling in which failure processes and sensitivity of slender pillars to variations in rock mass properties are evaluated. The results showed that the strength of slender pillars is more variable than that of wider pillars. The numerical model results demonstrated the increasing role of brittle rock failure in slender pillar strength. The absence of confinement in slender pillars can result in a fully brittle failure process, while wider pillars fail in a combined brittle and shearing mode. The onset of spalling in slender pillars occurs at or near the ultimate strength, while this is not the case for wider pillars. Slender pillars are shown to be more sensitive to the presence of discontinuities than wider pillars, which can partly explain the increased variability of slender pillar strength. Two examples are presented, which illustrate failure initiation by brittle spalling and the sensitivity of slender pillars to the presence of discontinuities.

Jan 1, 2006

By V. van der Merwe, C. W. Siyasiya

"A petrochemical company experiences premature thermal fatigue failure of the casting moulds used in catalyst production. The aim of the project was to find an alternative alloy that would outperform the current low-alloy cast steel used for the moulds. Based on their thermo-fatigue properties, 3CR12 ferritic stainless steel and H11 tool steel were chosen for testing and comparison with the currently used BS3100 B7 cast steel. Samples of each material were subjected to temperature cycling in a Gleeble 1500TM thermo-mechanical processing simulator, followed by surface analyses. The main parameters derived from the test work were the total true strain, the hot strength of the materials, and the number of cycles to failure. Additionally, the coefficient of thermal expansion for each material was measured using a Bähr dilatometer. H11 tool steel yielded the best performance by way of having the fewest surface cracks, the lowest total true strain per cycle, the most cycles to failure, the highest hot strength, and the lowest coefficient of thermal expansion.IntroductionThe Catalyst Manufacturing Plant (CMP) is a dusty environment with temperatures higher than ambient. Molten mill scale (magnetite, Fe3O4) is tapped from the electric arc furnace into the casting moulds at temperatures close to 1600°C. After 10 seconds the magnetite is quenched with a water spray for 50 seconds to a minimum temperature of 400°C. Sometimes there is residual water in the pans from the cooling fog process (water sprays) in the empty mould. In a typical plant, the factory produces 40–44 batches a day containing between 150–300 moulds, depending on the time required for the specific batch (Buchholz, 2014). The typical batch in the CMP usually weighs about 6 t. The thickness of the tap, speed of the casting belt, and tilting angle all affect the number of moulds per batch.One of the main problems faced during the molten mill scale casting is the premature failure of the casting moulds. Cracks initiate on the surface of the mould, which is in contact with the hot catalyst, and propagate into the interior of the mould. Cracks are generally concentrated in the middle area of the mould where the catalyst makes first contact with the mould, and which experiences the highest temperature fluctuations. This suggests that the cracks are caused by thermal fatigue during the casting process."

Jan 1, 2016

By R. S. C. Rogers

The present study examines the use of two vs. one stage of hydrocyclones in a pilot scale ball mill circuit. The approach used for doing this is based on computer simulation with mathematical models that explicitly account for the grinding mill/classification system interaction. Parameters for the ball mill model are obtained using a short-lived radioactive tracer method in open circuit tests of the study mill. The hydrocyclone model, based on the size selectivity concept, is developed and validated in standalone testing of the study hydrocyclones. Simulation procedures are validated through actual tests of the closed circuit grinding configurations of interest. Results of the evaluation indicate that two stages of hydrocyclones in the study circuit would improve classification, increase productivity and lead to more closely sized circuit products when compared to the conventional one stage hydrocyclone circuit. A computer simulation approach for comparing these circuit configurations in larger scale grinding circuits is also suggested.

Jan 1, 1981

By Christopher J. Bise

It is somewhat reassuring to note that the United States has an abundance of coal reserves which can be used to bridge the widening gap between the nation&apos;s energy production and consumption. The U. S. Bureau of Mines has recently established a demonstrated coal reserve base of 434 billion tons. (31)(Figure 1) As can be observed, the major percentage of the reserves is in the West. Although surface coal mining in the West has been gaining increased attention in recent years, western reserves that must be deep mined are significantly greater. Though deep mining activity in the west is very limited and accounts for less than 2% of the annual production, the primary reason, quite probably, is the greater thickness of the seams as compared to those found in the East. Forty-five billion tons of the western demonstrated coal reserve base for underground mining are in seams greater than 10 ft. in thickness. (37) Thick coal seams do not lend themselves readily to extraction with the underground mining methods practiced in the United States today. Experiences from the practices in other countries can serve only as guidelines. Consequently, a concerted effort is necessary to overcome the technical difficulties.

Jan 1, 1975

By K. B. Venkatesh

In view of the importance of size reduction operations from the energy consumption point of view, during the past one hundred and fifty years a number of attempts have been made to develop a general energy-size relationship. However, to date it has not been possible to clearly define the scope of application of various laws proposed, though Bond law has been widely used with a moderate degree of success. It was, therefore, decided to carry out a thorough analysis of the batch ball mill grinding data available in the literature on several minerals that were ground under normal range of operating conditions. In particular, relationships proposed by Bond, Kapur, Voller and Morrell were examined in detail. It was found that the relationship proposed by Kapur in terms of the mean particle size provided a more accurate basis for estimation of energy. The value of the exponent of mean particle size in this relationship, a, is same as the exponent of the particle size in the expression for the grinding rate parameter: ki = Axia, provided the particle size distribution range corresponds to the self-similar or self-preserving regime. In this regime, mean particle size is proportional to the 80 percent passing size and size modulus. From the practical application point of view, it was possible to describe the variation of µ1 over limited, but significantly broad size range, using a constant value of a. Thus, by carrying out a few batches, grinding experiments over the size distribution range corresponding to the contents of the production mill, the applicable value of a can be determined. Keywords: energy-size relationship, ball mill, mean particle size, self-similar size distribution regime

Sep 1, 2012

The support of stopes by means of yielding timber props, with or without hydraulic props, is common in the mining industry. This paper discusses the ability of these support systems to absorb energy during a rockburst, and further investigates their capability to withstand a range of ground velocities. It is found that the rate of stope closure prior to a rockburst strongly influences the behaviour of these support systems during the rockburst. It is recommended that hydraulic props be used in conjunction with yielding timber props in these conditions, and that yielding timber props with an increased yield range are desirable.

Jan 1, 1991

Written contribution by J.F. Curtist The paper is an interesting exercise in the application of mathematics to a theoretical situation presumed to oc¬cur underground. However, some of the assumptions on which the calculations are based are unlikely to occur in any actual mining situation. Assumptions made in the paper Objections to these assumptions are noted seriatim. (1) `it is assumed that any loading-rate effect will be small and can be ignored&apos; (p. 3). A loading rate of the order of 50 kN-12 per cent of the maximum load-is customary to prevent the props being dislodged by the blast. (2) `it is unlikely that all the props will be installed with their full stroke available, but for the purpose of this analysis, it is assumed that this is the case&apos; (p. 3). A more reasonable assumption would be that, with the most careful selection of prop length and the most rigid control of stope width, the available stroke length would on average be reduced by the order of 70 mm on installation-20 per cent of full stroke. (3) `each support unit supports a block of 3,3 m2 bound¬ed by vertical fractures with zero cohesion that are parallel and at right-angles to the stope face respec¬tively&apos; (p. 5). Joughin and Jager have shown this to be an unwar¬ranted assumption in respect of the positioning, the direction, and the dip of the fractures present in a stope in rockburst conditions. If, as assumed, there was `zero cohesion&apos; between such blocks, the removal of a hydraulic prop in row 3 supporting such a block would ensure that it fell out before a profile prop could be installed in the same position. (4) `the height of each of these blocks is 3,0 m, the volume of each block is 9,9M3, the mass of each block is 27 225 kg&apos; (p. 5). The concept of individual props being affected by dis¬crete blocks of known and uniform dimensions of relatively low ground velocities (less than or equal to 2,16 m/s) acting in line with the axis of the prop is one that lends itself to neat mathematical calcula¬tions. It does not, however, bear much similarity to the actual conditions that are exposed underground in excavations where rockbursts occur. Ortlepp and co-workers concluded that `the impulse of the shockwave can cause the displacement of the

Jan 1, 1991

By Paulo Conceição, Carlos Sampaio, Carlos Petter, Ariane Kuerten

The Sensor Based-Sorting technology is an emerging technology applied in mineral separation and concentration. The technique consists in using a method known as Dual Energy X-ray Transmission sorting (DE-XRT). The particle's position over the conveyor belt was evaluated and, despite de variability of results, shows that there is no a position with greater accuracy on the sensor's reading. Results obtained indicated that DE-XRT sorting technology was able to significantly reduce the mass and the ash content of the ore feed, being an interesting option for reduction of costs and improvement of coal quality in beneficiation plants.

Oct 3, 2019

By Burcu Ertug, Necip Ünlü

"On average, one out of two glass bottles coming on to the market is provided by recycling processes in Europe. Glass recycling has been started by the early of 1970s in Turkey, and the recycling amount has been reached to 70 thousand tons recently, while 4.2 million tons which represents only 25 percent of total glass manufactured, is the amount of glass collected annually for recycling in the world. The glass recycle process basically involves collecting, sorting, then crushing and grinding after which glass cullet is produced as the secondary raw material. Removing of the contaminants from the glass cullet, and then melting are the next steps. In this study, recent developments of 100% recyclable without any deterioration in quality, and never decomposable glass, are discussed detailed in terms of glass production processes.1.IntroductionIn many countries, waste glass constitutes a large proportion of solid wastes, and recycling the glass is a major environmental problem all over the world [1]. Among the diverse types of waste, glass is one of the most valuable recycling materials because it can be revived and recycled completely and hardly decomposes [2]. Between 80% and 85% of the mass output from the worldwide glass industry is either in the form of containers for the food, beverage, and pharmaceutical industries, or flat glass for building construction or for motor vehicle manufacture. Other product segments, whilst only constituting about 15% of the mass output, produce high value technical and consumer products by comparison with container and flat glass [3].In the present study, the recycling of Liquid Crystal Display (LCD), Cathode Ray Tubes (CRT) and Printed Circuit Boards (PCB) were examined. The production of glasses and glass-ceramics from glass cullet, synthesis of different compounds from recycled glass were other subjects of interest. The use of recycled glass for construction industry and using glass as aggregate were emphasized. Finally the recovery of glass fibers were investigated."

Jan 1, 2012