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By A. Grimod

The European Guideline ETAG 027 - 2008 (?Guideline for European Technical Approval of Falling Rock Protection Kits?) defines the procedures to test rock fall barriers. Nowadays, this guideline is essentially the only test and construction framework utilized by manufacturers and it is also starting to enter in the mentality of designer who can clearly specify the performances of rockfall barriers. ETAG 27 (and the related European Technical Approval and the CE marking) represents a milestone for the rockfall barrier market, because it gives the possibility to compare the performances of different fences and it ensures the quality of the certified product. For these reasons ETAG 27 presently constitutes the base for tenders all around the world. Although rockfall barriers are in common use, designers and consultants do not have a clear overview of the real performances and the technological limits of these structures. The first problem can be solved using ETAG 27 criteria, which allow designing rockfall barriers at ultimate and serviceability limit state. Even though ETAG 27 uses crash tests performed with the Service Energy Level (SEL) and Maximum Energy Level (MEL), design with these concepts is not obvious or in common acceptance. Some preliminary questions must be solved. When can these energy levels be correctly applied? In the case of rockfall simulations, what probabilistic approach must be used for the evaluation of the barrier? Unfortunately ETAG 27 does not give any information or tests relating to the technological limits of the fences. Therefore some important questions arise. How do the site specific aspects of installation affect the barrier performance? What is the behaviour of the barriers under extreme conditions? And finally, the main question lies in the background: what is the functional reliability of ETAG 27? All of the above questions require comprehensive answers, which is consistent with the experience of designers, installers and producers. Officine Maccaferri S.p.A., with its 130 year experience in the field of rockfall protection, collected a large number of case histories with barriers impacted by single or multiple blocks, debris flows, and snow avalanches. Throughout these case histories and in accordance with a recent Italian standard code, it has been possible to implement a reasonable design approach that considers the efficiency of the barrier, the uncertainties of the input data, and the fore

Aug 1, 2013

By A. Shakoor, T. P. Gautam

The long-term slaking behavior of clay-bearing rocks in the field can be quite different than the short-term slaking behavior in the laboratory. To assess this difference, 12 replicate samples of each of 20 clay-bearing rocks, including 5 claystones, 5 mudstones, 5 siltstones, and 5 shales, were exposed to natural climatic conditions for 12 months. After each month of exposure, one replicate sample of each rock was removed from natural exposure and its grain size distribution was determined. Disintegration ratio (DR), defined as the ratio of the area under the grain size distribution curve of the slaked material for a given rock to the total area encompassing all grain size distribution curves of the tested samples, was used to quantify the amount of disintegration under natural climatic conditions. DR values of naturally exposed samples were then compared with the corresponding values of second-cycle slake durability index (Id2) determined in the laboratory. The relationship between Id2 and DR after, 1 month of exposure, exhibits an R2 value of 0.65 but the relationship deteriorates when DR values for 6 and 12 months of exposure are used, with R2 values of 0.63 and 0.25, respectively. However, when DR values for laboratory samples, computed from grain size distributions of slaked material after second cycle, are compared with DR values of naturally exposed samples for the three periods of exposure, the relationship improves with the corresponding R2 values being 0.61, 0.71, and 0.41. These results suggest that DR may be a better parameter for assessing the long-term disintegration behavior of clay-bearing rocks than Id2.

Jan 1, 2015

By C. M. Clapperton, K. E. Darke, A. J. Boyce, C. M. Rice, S. D. Redwood

"Abstract - The Kori Kollo gold deposit is hosted by a c.15 Ma sericitized dacite stock in the La Joya district on the Bolivian Altiplano. The mineralization is of the Bolivian polymetallic vein (BPV) type, with a significant amount of the gold reserves contained in the oxide cap. Field and microscopic studies, K-Ar dating, ??34S values of pyrite, alunite and jarosite, and ??18O values of alunite and jarosite demonstrate that some of this gold is supergene. Two types of supergene gold are present: minute gold grains, 1 to 10 microns across, of high fineness and associated with jarosite and iron oxy-hydroxides; and ‘invisible’ gold, which is probably sorbed onto minerals such as jarosite and iron oxy-hydroxides.Pyrite ??34S values fall in the limited range -3.5 ± 0.6‰, consistent with a magmatic origin. Sulfate ??34S values are homogeneous and overlap closely with those of pyrite (-1.2 ± 2.1‰), which supports petrographic evidence that sulfates replaced pyrite. By contrast, sulfate ??18O values are inhomogeneous with a large range (alunite +2 to 26‰, jarosite -5 to +18‰). These data are interpreted as reflecting kinetically controlled reactions involving various proportions of atmospheric oxygen and local meteoric water. This is typical of supergene alteration and is due to the seasonally dry climate of the Altiplano.The oxide-zone mineralogy is dominated by sulfates and iron oxy-hydroxide minerals with some phosphates and halides, which indicates that the groundwaters were acidic, oxidizing and relatively saline. Under these conditions the most likely complexes responsible for gold mobilization are halides.Geological data and K-Ar ages indicate that oxidation of the BPV deposits began at least around 11 Ma and possibly up to 15 Ma ago, and that supergene alteration lasted until around 4 Ma. During this period, erosion was minimal. Although supergene processes have been active on the Bolivian Altiplano since the Middle Miocene, in contrast to Northern Chile, enrichment of gold has not occurred."

Jan 1, 1997

By C. M. Clapperton, K. E. Darke, A. J. Boyce, C. M. Rice, S. D. Redwood

"The Kori Kollo gold deposit is hosted by a c.15 Ma sericitized dacite stock in the LaJoya district on the Bolivian Altiplano. The mineralization is of the Bolivian polymetallic vein(BPV) type, with a signi?cant amount of the gold reserves contained in the oxide cap.Field and microscopic studies, K-Ar dating, ??34S values of pyrite, alunite and jarosite, and ??18O values of alunite and jarosite demonstrate that some of this gold is supergene. Two types of supergene gold are present: minute gold grains, 1 to 10 microns across, of high fineness and associated with jarosite and iron oxy-hydroxides; and ‘invisible’ gold, which is probably sorbed onto minerals such as jarosite and iron oxy-hydroxides.Pyrite ??34S values fall in the limited range -3.5 ± 0.6‰, consistent with a magmatic origin. Sulfate ??34S values are homogeneous and overlap closely with those of pyrite (-1.2 ± 2.1‰), which supports petrographic evidence that sulfates replaced pyrite. By contrast, sulfate ??18O values are inhomogeneous with a large range (alunite +2 to 26‰, jarosite -5 to +18‰). These data are interpreted as reflecting kinetically controlled reactions involving various proportions of atmospheric oxygen and local meteoric water. This is typical of supergene alteration and is due to the seasonally dry climate of the Altiplano.The oxide-zone mineralogy is dominated by sulfates and iron oxy-hydroxide minerals with some phosphates and halides, which indicates that the groundwaters were acidic, oxidizing and relatively saline. Under these conditions the most likely complexes responsible for gold mobilization are halides.Geological data and K-Ar ages indicate that oxidation of the BPV deposits began at least around 11 Ma and possibly up to 15 Ma ago, and that supergene alteration lasted until around 4 Ma. During this period, erosion was minimal. Although supergene processes have been active on the Bolivian Altiplano since the Middle Miocene, in contrast to Northern Chile, enrichment of gold has not occurred."

Jan 1, 1997

By T. Takayama Filho

Votorantim Metais Zinco is a smelting facility which produces SHG zinc, zinc alloys, sulfuric acid, liquid SO2, copper sulfate, cadmium metal and silver concentrate. It is located at Juiz de Fora, Minas Gerais State, Brazil. This facility has been operating since 1980 and its current capacity is 95,000 tonnes of zinc per year, using mostly imported sulfide concentrates.The plant uses the conventional roast-leach electrowinning (RLE) process, in which iron is removed from the circuit as jarosite. The residues generated by the RLE process, which amount to about 200 t/day (dry weight), were until 10 years ago, deposited in tailing ponds having no impervious lining. More recently, jarosite and other wastes have been deposited in a properly built reservoir, but this is reaching its maximum capacity. In order to decommission these disposal operations and to reclaim the sites in an environmentally acceptable manner, Votorantim Metais concluded in 2004 a rehabilitation project for the old tailings pond and prepared detailed plans for managing the current and future disposal of jarosite-bearing residues. This paper describes the procedure for removing iron from the hydrometallurgical circuit and discusses the past and current status of the residue disposal sites, as well as the future challenges related to the environmentally acceptable handling of iron residues.

Jan 1, 2006

By W. G. McBride

INTRODUCTION Preparation of this article, at the request of the Committee on papers, has been undertaken in the hope that it may stir up helpful discussion and constructive criticism. The Canadian mining industry is in the dumps and there is no use trying to deny the fact; nor does it help to say that other industries are in as bad or a worse condition. Let us face the situation honestly and see just how serious it is. The entire financial and industrial world is in a state of depression. The gravity of the case is no doubt exaggerated by contrast with the wave of optimism and false prosperity which pre-ceded it. The prevailing gloom and pessimism are just as irrational as was the former super-optimism. As far as we mining engineers are concerned, the causes are quite beyond our control, and it is the effect upon our own industry which interests us most. This effect is an accumulation of mineral products with consequent depression in price to a point where profits have almost, if not quite, disappeared. Whether this accumulation of stocks is the result of over-production or under-consumption is difficult to say, but it would seem likely that there is a combination of the two. The introduction of selective flotation has brought about a large increase in the production of zinc, a moderate increase in the production of lead, and a comparatively small increase in copper. It has turned relatively small mines into great producers, as in the case of the Sullivan, revived expiring mines like La Cananea, and brought-in new ones like Flin Flon

Jan 1, 1931

By E. Watkins, J. C. LeRoy, A. Vivyurka

"TAILINGS REVEGETATIONWelcome to the Revegetation Session. My instructions are to introduce the Speakers and to provide a general introduction.I do the latter with humility not only because of a lack of personal experience or because tailings revegation is not well established technology, but because any general statement on a country or industry as diverse as the Canadian Mining Industry is bound to be wrong somewhere.Laying aside normal standards of accuracy I proceed as follows:In considering the question of tailings revegetation, one asks two basic questions:1. What ought to be done?2. What can be done?The first question can be broken down into three kinds of criteria or objectives:a) Legislativeb) Technicalc) AestheticThe first I leave to the politicians and the last I will avoid. Aesthetics is the philosphy of taste. It is concerned with the appreciation of things rather than what they really are. It is much easier to get into an argument on aesthetics than to get out of it feeling you have said something of consequence.Technical criteria can be divided into the following sub-objectives: 1. To isolate the tailings basin so surrounding land is not degraded. The first step here is stabilization of the surface to prevent wind and water erosion."

Jan 1, 1975

By P. H. Jennings

Commercial tellurium commonly contains heavy basemetal impurities such as lead, copper, iron, etc., as well as selenium and oxygen. The base metals may be removed by vacuum distillation or by zone melting. Vacuum distillation is a convenient method for treating moderate quantities (~ 100 kg.) for the production of a material having a total content of heavy metals in the 10-30 ppm range. By zone melting, it is possible to lower all the heavy metal impurities tm below the limits of spectrographie detection. Neither vacuum distillation nor zone melting is effective for the removal of selenium, but treatment of molten tellurium with hydrogen converts selenium to hydrogen selenide. This procedure is too s low to be considered except as an adjunct to zone melting. In the course of air-melting, many of the more easily oxidized impurities form a dross and may thus be readily removed. Substantial refinement with respect to iron, tin, chromium, aluminum, sodium, silicon, magnesium, calcium and manganese can be achieved in this way. Although tellurium can be electrodeposited from either acid or alkaline solutions of the dioxide, various factors militate against successful electro-refining, amongst which may be mentioned the brittleness of cast tellurium anodes and the tendency for the metal to be deposited in dendritic form, even at low current densities.

Feb 1, 1966

By H. A. Pearse

The history of the flotation process, as applied to the practical field of ore dressing, dates back somewhat less than twenty years; yet this fact may be overlooked in view of the present universal dependence on this process in the concentration of almost all classes of ore. It is therefore interesting to recall that Britannia was the scene of some of the earliest experiments with flotation on a commercial scale. In 1911, sixteen years ago, an Elmore vacuum machine was operated for several months with encouraging results. In 1913, some time before flotation was very generally adopted elsewhere, a Minerals Separation plant, capable of treating 600 tons daily, was operating successfully at Britannia. Considering how little was known of the theory and practice of flotation at that time, the recovery was remarkably good, but it was not until the present mill was completed in 1923 that any attempt was made to raise the concentrate grade by differential flotation, or, in other words, to float the chalcopyrite while rejecting the then worthless pyrite with the tailings. The first ore passed through the present concentrator in March, 1923. As originally designed, the rated capacity was for 2,000-2,500 tons, but this has been increased year by year, until at the present time 4,000 tons daily can be treated economically. During the first few months of operation the concentrate grade averaged only 10-11 per cent Cu, but by the end of 1923 this had been raised to over 20 per cent, due to the elimination of the pyrite by differential flotation. In 1923 the tailings averaged approximately 0.25 per cent Cu, but also in the matter of recovery a considerable improvement has been effected, for during the year 1926 an average tailing of 0.15 per cent Cu was maintained.

Jan 1, 1927

By T. G. Ewart

INTRODUCTION FOR THE purpose of this paper, western bituminous coal is defined as steam coal produced in Alberta and southeastern British Columbia for sale primarily to railways and industries. Before examining present and future markets of this industry, it is proper to review briefly financial and operating conditions within the industry. The report of the Royal Commission on Coal, 1946, supplies financial data from which some conclusions may be drawn. In 1944, the industry as defined above had share capital of $24,886,696. In the fifteen years 1930 to 1944, inclusive, the industry produced: 44,364,211 tons of coal. The profit on this tonnage amounted to $4,678,382.58, which represented 10.6 cents per ton of production. Miscellaneous earnings increased the amount available for dividends to shareholders to a total of $8,487,952.41, which is equal to, 19.1 cents per ton of production. Dividends actually paid in this fifteen-year period amounted to $12,332,040.64, which represented an annual average rate of 3.40 per cent on share capital. It is worthy of note that, in 1944, the net current position of the industry, including investments, was $8,753,874.51. The writer believes that the financial position of the industry with respect to capital, current financial position, and net earnings, has not significantly changed from the fifteen-year period out-lined above. The figures quoted indicate the following situation: (1) The western bituminous coal industry is adequately financed. (2) The industry has followed a conservative dividend policy and has accrued funds for contingencies. (3) The industry has sold its production at very reasonable prices as evidenced by its moderate earnings per ton of production.

Jan 1, 1951

By L. Tafaghodi, S. Sarbishei

"Following the depletion of nickel sulfide ores, laterites are considered as one of the main sources for producing nickel and cobalt. One of the primary steps in pyrometallurgical extraction of nickel from laterites involves partial reduction of the ore at temperatures below 1000°C. Studying the kinetics of laterite reduction provides valuable information on reducibility of nickel laterites using coal as a reductant. The present study investigates the reduction behavior of a nickel laterite by thermal gravimetric analysis and differential scanning calorimetry. Non-isothermal reduction tests and X-ray diffraction analysis were performed to find the main reactions that occur during the process. The kinetic parameters were calculated by Coats-Redfern method. The general model of simple reaction f(a)=(1-a)n was used for calculating kinetic parameters.INTRODUCTION Nickel (Ni) naturally occurs in the form of sulfide and laterite ores. Laterite ores are mainly available in tropical regions. Weathering phenomena during millions of years has led to the formation of several layers with specific mineralogy (Butt & Cluzel, 2013). The three main layers downward are limonite, smectite, and saprolite (Crundwell et al., 2011). Although laterites are the most available nickel ore resource (70%), their share in the overall Ni production is limited to 40% (Crundwell et al., 2011). The main reasons for not materializing the full potential of laterites are as follows (Ray et al., 1985): •,Laterite ores contain high water content (free and combined) which in turn results in high energy consumption of the extraction process. •,As Ni is present in the form of solid solution in laterites, physical upgrading methods such as flotation are not effective in increasing the Ni content before the extraction process. •,Extracting Ni from sulfide ores involve exothermic reactions which are beneficial due to decreasing the overall energy requirement of the process, however there is not efficient release of energy in the process of extracting Ni from laterite ores."

Jan 1, 2017

IDS: The Company IDS Ingegneria Dei Sistemi, founded in 1980, is a world leader as a high technology solution provider in selected niche defense and civil market sectors ? Private and Independent ? IDS?s Quality System is certified ISO 9001 Ed. 2000 ? IDS has the highest Military Clearance ? Continuously measure (24x7) mine wall movements with sub-millimetric accuracy at long distance (up to 4 Km) ? Provide reliable early warning for progressive slope movements potentially leading to mining wall failures ?Map the long term evolution of slow moving slopes to get a better knowledge of the rock mass strength

Aug 1, 2013

By G. A. Vissac

In this paper we intend to deal particularly with the economic side of coal washing as it applies to our mines. We will then discuss why the dry-cleaning process has been selected, and give a brief description of our plant, dwelling only on its specific features and referring the reader to the numerous papers published on the subject for any details on the principles and machinery used in dry cleaning. ECONOMIC DISCUSSION The washery introduces an extra cost for depreciation, operation, and loss in treatment. A certain proportion of the coal, now being sold, and which the miners have been paid for. will have to be wasted. As a whole, the total increase in cost would be approximately 60 cents per ton. On the possibility of recovering this amount depends the success of the washery. There are two advantages to be expected from the washery, the first, which occurs at the mine and which will be certain; the second, more difficult to estimate, because it depends upon convincing our customers that they should pay more for a better product.

Jan 1, 1924

By Terry Gerritsen, Sylvain Picard, Bruce MacKay

Virtually all new technologies require a control system. Any intense process has many process variables that react quickly making manual control difficult. Control system design has a significant impact on the success of a plant especially during ramp-up. New technology projects are particularly sensitive to control system design as the process response is unknown. This paper discusses the important factors encountered in the design of a control system for a new technology. Examples are given of successful new technology plant start-ups where tight coordination of the process control system implementation, process design and selection major equipment led to rapid ramp-ups.

Jan 1, 2011

By P. H. Jennings, J. C. T. Farge

Commercial tellurium commonly contains heavy base-metal impurities such as lead, copper, iron, etc., as well as selenium and oxygen. The base metals may be removed by vacuum distillation or by zone melting. Vacuum distillation is a convenient method for treating moderate quantities ( 100 kg.) for the production of a material having a total content of heavy metals in the 10-30 ppm range. By zone melting, it is possible to lower all the heavy metal impurities tm below the limits of spectrographie detection. Neither vacuum distillation nor zone melting is effective for the removal of selenium, but treatment of molten tellurium with hydrogen converts selenium to hydrogen selenide. This procedure is too slow to be considered except as an adjunct to zone melting. In the course of air-melting, many of the more easily oxidized impurities form a dross and may thus be readily removed. Substantial refinement with respect to iron, tin, chromium, aluminum, sodium, silicon, magnesium, calcium and manganese can be achieved in this way. Although tellurium can be electrodeposited from either acid or alkaline solutions of the dioxide, various factors militate against successful electro-refining, amongst which may be mentioned the brittleness of cast tellurium anodes and the tendency for the metal to be deposited in dendritic form, even at low current densities.

Jan 1, 1966

By J. C. Nicholson

The "V" system was first introduced by the West Virginia Coal & Coke Company at their No. 9 mine, Norton. Norton is a small mining village on the B. & 0. railway, about seven miles from the town of Elkins. This year, accompanied by G. D. MacDougall, I visited this mine to obtain first-hand information regarding this modem system and to determine its adaptability to conditions in Nova Scotia. The following is a brief account of my observations. The seam that is being worked is known as the Lower Kittaning, which, at this point, has an average thickness of 7 feet. This includes a band of hard clay varying in thickness from 8 to 14 inches, which occurs about 2 feet from the pavement, also 8 inches of inferior coal at the roof. The latter is not recovered, which leaves to be mined about 5 1/2 feet of clean coal. The roof is slate and shale for the first 30 feet, followed by 40 to 60 feet of sandstone; the total cover is from 80 to 250 feet thick. The seam lies very flat, with a dip of only 2 per cent., and, as the dip is toward the main entrance, it' is in favor of the load, also providing natural drainage for any water the mine makes. As the mine is not gassy, and there is no water to be pumped out, one can understand that mining conditions are practically ideal.

Jan 1, 1924

By J. R. Welsh, D. J. Palmer, O. M. G. Newman

"The use of calcine for selective zinc precipitation requires specialist knowledge to achieve satisfactory flowsheet design. Calcine from the bed of a fluid roaster is preferred, and factors such as the size distribution of ground calcine, the stoichiometric ratio of the calcine precipitant to the initial dissolved zinc concentration and the moisture content of the resultant filter cake are parameters which require critical attention. Extensive testing in a fully-integrated electrolytic zinc plant pilot plant has demonstrated that properly designed selective zinc precipitation plants using calcine as the precipitant are reliable operating units.The advantages of using calcine instead of limestone for selective zinc precipitation include a decreased cost of reagents, as well as avoiding the formation of by-product gypsum.Process simulation modelling has demonstrated that the use of calcine in this context has widespread application in the design of zinc plants with superior impurity and water balance control characteristics. A particularly powerful approach involves closed circuit washing. In this mode, filtrate from the selective zinc precipitation step is used to wash a zinc plant residue such as jarosite, and the resulting filtrate is returned to the selective zinc precipitation section. IntroductionThe Selective Zinc Precipitation (SZP) Process, developed and patented by the Electrolytic Zinc Company of Australasia Limited (EZ) may be used to achieve simultaneously, water balance and impurity control in an electrolytic zinc plantO,2). Impurities which may be controlled by the SZP process include chlorine, magnesium, manganese, potassium and sodium.In the SZP process, basic zinc sulphate is precipitated from a zinc sulphate solution by either limestone, lime or a compound containing zinc oxide, usually calcined zinc sulphide concentrate (calcine) or alternatively zinc oxide fume .The solids containing basic zinc sulphate are separated, typically, by filtration and returned to the zinc plant circuit as a neutralizing agent. The zinc depleted solution can either be discarded, or used to wash residues, thereby achieving water balance and impurity control (Fig. I)."

Jan 1, 1988

By P. F. Venus

In December 1977, a large pillar was removed by blasthole mining methods at lnco's Creighton No. 3 mine. Although the 2,363.000-ton blast was not a size record for a single underground blast at Inco, it did represent a significant departure from conventional blasthole mining. In terms of the application of nominal 6-inch-diameter boreholes underground, coupled with bulk slurry explosives, the magnitude of the blast is considered a first and very important development. The application of large-diameter boreholes in underground operations was first tried successfully in 1973. Since that time over 13/4 million feet have been drilled in this adaptation of an open-pit technique to underground operations. The blast, which required 708 holes, totalling 131,500 feet of drilling, was loaded in a six-week period, using special slurry handling techniques developed jointly by Inco and CIL. Inco operating and engineering personnel consider this application of large boreholes and slurry explosives for blasthole mining a very important milestone in modern mining technology.

Jan 1, 1978

By I. V. Kojo

Continuous development of technologies has been self-explanatory for Outokumpu Technology's engineers for decades. The aim in the development has always been the improvement of profitability of the copper production but also the sustainability of the processing i.e. high copper recovery, low external fossil based energy utilization, high emission captures and of course low operation and investment costs. Outokumpu Technology's long-term view of the development is a change towards continuous processes and elimination of number of unit operations in smelting i.e. elimination of possible sources of emissions. Outokumpu blister smelting processes i.e. Outokumpu Direct-to-Blister Flash Smelting and Kennecott-Outokumpu Flash Converting have shown excellent operational results during the years of operation and it is expected that soon more than 10 % of the world primary copper is produced by these technologies. The paper will give some background of this development trend.

Jan 1, 2007

By Deepak Bhadouriya, Saurabh Singh, Aditya Raj Suman

"An experimental study of chemically enhanced filtration and dewatering of fine coal slurry with and without additives was carried out. The experiment involved determination of effect of slurry concentration, additives (cationic, anionic and nonionic), filter area, filtration velocity, particle size and cake resistance on dewatering and filtration using vacuum filtration cell. Finally, the rate of filtration and dewatering and optimum amount of additives used were determined under suitable filtration condition. Overall, it was found that the filtration and dewatering characteristics were enhanced by increased slurry concentration, larger filter area and larger particle size. In addition of additives, the kerosene was found to be best for enhancing the filtration which yielded cake moisture reductions of approximately 17% by weight. At optimum dosages, the final cake moisture contents were reduced by 3.6% and 5.8% by weight using the appropriate surfactant and flocculent, respectively.INTRODUCTIONFor India, it is estimated that coal constitute a principal source of energy for the next several decades. It is majorly used at thermal power plants to generate electricity and also to a great extent in steel and cement industry. It has been observed that its consumption is increasing so as to meet the energy requirements of the country. It has been found that in a coal preparation plant; approximately 20-25 percent coal is of the mined coal of the range below minus 0.5 mm (28 meshes). Generally, this fine fraction is discarded due to its high cost of processing and further utilisatisation is very minimal. But with ongoing research and development for advanced clean coal technology, various new techniques have been evolved to maximize the utilization of available coal. In recent time technologies such as column flotation, cleaning of fine size coal to lower the ash content and low pyretic sulfur is feasible at high recovery value. But in various researches it is found that one of the biggest hurdles in utilization of fine coal cleaning technology by the coal industry is the economic dewatering of the fine clean coal product. Thus, until an economical and practical solution to dewatering of fine clean coal is achieved, the efforts devoted to developing fine clean coal technology will be wasted. In this paper an effort toward enhancing and developing the optimum technique of dewatering and filtration of fine coal slurry has been made."

Jan 1, 2016