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By B. Daniels, A. Tieu, W. Yan, B. Ren, M. Vijay

"Plasma or thermal sprayed coatings are becoming more widely used in a wide range of industrial applications. The preference to use thermal sprayed coatings is mainly attributed to its superior wear, and corrosion resistant properties akin to hard chrome but without the harmful health implications of chrome in the coating process. Specifically, High Velocity Oxygen Fuel (HVOF) plasma sprayed Tungsten Carbide & Cobalt (WC/Co) coatings are presently being tested on aircraft landing gear. This coating can be efficiently applied and is highly resistant to corrosion and wear. However, HVOF coating's durability also makes it difficult to be stripped for inspection and overhaul. Current coating removal methods usually involve a multi-step process whereby the coating must undergo diamond wheel grinding, chemical dipping, grit blasting, ultrahigh pressure (UHP) 380 MPa (55 Kpsi) water blasting, and subsequent chemical dipping, depending on the type and thickness of the coating. As explained by Yan, et al [l]"" a unique method has been developed for removing HVOF coatings with the use of only water. The method consists of modulating a stream of water at a pressure of only 69 MPa (10 Kpsi) with ultrasonic waves to generate a forced pulsed waterjet. The resulting cyclic loading of the hard coating due to the impact of the water pulses easily causes it to flake off the substrate. This paper will briefly describe the fundamentals of the forced pulsed waterjet technique with an emphasis on insitu applications of the removal of various plasma and HVOF coatings."

Jan 1, 2006

By Patrick Fantin

Mining ? The Past Safety Not a priority; Health Not a priority; Energy Manual Labour; Horsepower? Literally. Financially Robust For the Owner?s; Mining ? The Future Safety Zero Harm Health Understanding and elimination of Industrial Disease Energy Efficient use of energy through technology. Full value of renewable energy. Financially Robust Continually build shareholder value through optimum CAPEX/OPEX.

May 1, 2013

By M. Melashvili

"The SART Process was introduced to the gold mining industry in the late 1990s and several commercial plants were built over the following 10 to 15 years. Most of these plants have operated well and enjoyed plant performance in line with expectations based on the underlying chemistry of the process. However, two of the plants have consistently suffered from high reagent consumption and poor product quality, which has motivated a re-examination of the process fundamentals in the hope of understanding the reasons for poor performance when it occurs, and developing mitigating strategies.This testwork along with a review of the relevant literature has demonstrated that the amount of sulphide added to the primary reactor, relative to the copper concentration, is the most critical parameter dictating plant performance, with pH and residence time important secondary factors. The primary product in the SART Process is synthetic chalcocite, Cu2S, and the stoichiometric requirement for sulphide addition is therefore 0.5 moles per mole of copper in solution. Testwork shows that when sulphide is added in excess of this amount (generally with the express purpose of lowering residual copper in solution) it not only fails in its intended purpose, but can have a negative impact on the settling and filtration characteristics of the copper sulphide product, as well as the final product quality. Moreover, these problems are exacerbated the higher the selected operating pH in the pH 3 to 5 range, and the longer the residence time of slurry in this pH range.Freshly precipitated copper sulphide compounds are known to be very fine and slimy, with poor settling and filtration characteristics. It is therefore important to engineer the SART Process in such a way as to encourage growth in the average particle size of the precipitate crystals. This was the rationale for recycling a portion of the thickener underflow to the primary reactor when the SART Process was developed. The theory of crystallization predicts that compounds of extremely low solubility, such as chalcocite (Cu2S: KSP = 10-48), are likely to experience growth in crystal particle size through aggregation rather than heterogeneous nucleation. On this basis, the SART practice of recycling a portion of thickener underflow to the primary reactor in the commercial plants is unlikely to lead to crystal growth via nucleation, and could in fact promote breakage of the aggregates that formed in the thickener. Recycling thickener underflow could therefore have the opposite of the desired outcome. The result of testwork that is reported here lends credence to this theory and questions the wisdom of recycling in the SART process.There is evidence to suggest that operating the copper sulphide thickener in the SART Process without flocculants may be beneficial. Thickener underflow densities could improve from the current 10 - 15% range to 40 - 50%. This will not only reduce the flowrate of slurry to the filter, but the filtration properties of the slurry (filtration rate, filter cake moisture content) will also improve. This is one of the most critical engineering and design aspects of the SART Process, and the achievement of high solids density in the thickener underflow will not only improve the health and safety aspects of the operation, but also has the potential to lower capital costs and improve the quality of the final product."

Jan 1, 2016

By J. Koponen, M. van de Vijfeijken, L. Bos

"The availability and reliability of large grinding mills and their Gearless Mill Drives (GMDs) are of paramount importance for the grinding process. Although GMDs have been in service for decades, and are known for their operational efficiency (reliability and robustness with reduced operating cost per ton), compared to conventional drives, in the last decade due to certain reasons some failures led to serious unscheduled shutdowns and significant insurance events.The majority of these notable ""insured"" failures have happened on GMDs installed on relatively large mills, as the financial impact of a failure is proportional to the through-put of the mill. Despite the occurrences of these events, as mining operations are driven by lower ore grades and the ability to continue reducing operating costs per ton, the trend for having larger mills is still very present and there is no reason why this trend should not continue.It is obvious that if not handled correctly, the possibility of large value GMD loss events, coupled with the continuing increase in mill size or required availability and reliability targets, could potentially lead to major conflicts as it becomes difficult for the insurance companies to judge the risk and hence may result in hesitation to insure these grinding circuits. This paper will not only describe the conflicting situation from both an insurer's perspective and from an equipment supplier's perspective, but also how these conflicting challenges can be successfully mastered. It has to be noted that not only the mining companies themselves, but also their insurance companies and the involved equipment suppliers have a common interest: to reach the highest availability and reliability in order to avoid or at least minimize unscheduled shut down times, and most importantly prevent unplanned events that result in extended production stoppages or reductions in through-put of GMD driven mills.From the very beginning onwards, the complete service concept together with emergency plans must be integrated on every level in the design of the grinding circuit. This approach must be implemented continuously and without compromises by all involved parties during all following stages, such as basic design, detail design, manufacturing, quality inspection and testing, packing, transportation, storage, installation, commissioning, operation and maintenance. The concept of formal risk analysis related to mill operations, maintenance and spare parts strategies must be an integral part of mining companies’ decisions related to the selection for installation and continued use of large mills powered by GMDs. The service concept in place for operating mills shall also have a clear focus on very early detection of any abnormalities, so that the risk can be mitigated and unscheduled shutdown times avoided. However, if a major failure occurs, then procedures and tools shall already be available in order to ensure that production impact is minimized."

Jan 1, 2012

By L. T. Munn

"IT IS NOT necessary here to dwell on the fact that sulphur is currently in short supply. It may be well, however, to •point out that this •storage is not merely a temporary condition, but that, judging by trends in world consumption of sulphur, it may •be anticipated that requirements will double in the next ten years.Even with improved techniques for mining elemental sulphur in the United States and the new finds in Louisiana, there is much concern as to how these future sulphur requirements can be met. It appears necessary to .turn more and more to other sources than elemental sulphur for •additional production. The three major alternative sources were:(1) By-product sulphuric acid •from smelter gases.(2) Sulphur and sulphuric acid from iron sulphides.(3) Sulphur from hydrogen sulphide contained in natural gas. The Jumping Pound sulphur plant of the Shell Oil Company of Canada, Limited, is designed to produce sulphur by the •third of these methods.Jumping Pound is a natural gas field about 20 miles west of Calgary where Shell has a number of producing wells."

Jan 1, 1952

By G. S. Hume

FOR MANY YEARS Canada lived in the hope of attaining a substantial production of petroleum and natural gas that, to geologists familiar with the sedimentary basins of this country, seemed an entirely reasonable expectation. There have been surges of exploration activity stimulated by oil and gas discoveries and development such as occurred in Turner Valley. However, it was not until the Leduc field in Alberta was found in 1947, and the significance of that discovery was realized, that the search became really intensive and continuons. Expansion now has become a matter-of-fact attitude based on results which would have been considered impossible of achievement a few years ago.

Jan 1, 1957

By Richard E. J. Putman

THE MINERALS INDUSTRY has always been a heavy user of energy in most of its unit operations: power in the mining, comminution, flotation and electro-winning processes; fuel in its smelting and other metallurgical processes; and, of course, fuel in the on-site generation of power and steam. Many of the techniques for using computers to perform energy management and conservation tasks have been known for some time, but they have not been widely implemented because the return on investment has hitherto been insufficient. Recent trends, however, have not only caused the dollar return to rise due to higher purchased energy costs, but the cost of the investment to fall. One trend has been the generally lower cost of computer control systems themselves, due to vendors having adopted the newer, lower-cost and more reliable electronic component technologies now becoming universally available. The second trend is the development of low-cost, but high-speed, remote multiplexing devices, which have reduced considerably the expense of bringing data from points scattered over a wide area, which is typically the case when applying energy management systems to a minerals processing plant. By taking advantage of these and related developments, computer control systems can now have a rewarding impact on plant energy management and conservation practices, and recent advances in application technology have opened up new and additional opportunities for energy cost reduction and control.

Jan 1, 1978

By H. M. Lima, L. A. Gomiero, C. A. de Morais

ABSTRACT The Caetité-Brazil uranium facility produces approximately 400 tonnes U3O8 per year from an ore averaging 0.29% U3O8. The process steps are ore crushing, heap leaching with sulphuric acid, uranium separation and purification by solvent extraction (SX) with a tertiary amine, followed by stripping with a sodium chloride solution, precipitation as ammonium diuranate (ADU), and product drying. A change in the milling process is being evaluated to increase production and uranium recovery. Heap leaching will be replaced by conventional agitated tank leaching of the –590 µm ground ore slurry in a sulphuric acid medium. Replacement of the stripping reagent is also being considered.

Jan 1, 2012

By W. L. Uglow

In recent years, the microscope has come to be recognized as a valuable aid to engineers whose business is the discovery, development, concentration or reduction of metallic ores. In many of the large treatment plants microscopical research has already an important status; and in the hands of efficient operators the microscope has indicated methods of solving ore problems resulting in the saving of thousands of dollars that otherwise might have been lost in tailings, or expended in freight rates or smelter charges. Failure to procure all the information that the microscope can supply concerning the physical and mineralogical characteristics of complex ores has in many cases led to the construction of unsuitable concentrators, just as the failure to investigate in detail the geological conditions of an ore deposit has often results in the squandering of money on shaft sinking, and mechanical installations which subsequent development of the deposit failed to justify. The great progress recently made in the selective separation of minerals by means of the various processes of flotation and leaching brings strongly to our attention, the scope of the field in which microscopical methods may be advantageously employed. The microscope enables one to observe the minute details of rock, ore, gangue, concentrate, tailing, matte, slag and metal, that are invisible to the naked eye. Furthermore, the technique of microscopal research has been so highly developed that a microscopist can not only observe, but deter-mine the compositions and measure the diameters of the mineral constituents in the ore to be treated. Magnifications of these constituents up to 3,000 diameters may be produced by the proper combination of eyepiece, objective, and projection apparatus, and the images thus viewed may be photographed through the microscope, and permanent records made of them.

Jan 1, 1924

By K. Östcrdahl

We are developing a cost-saving and environmentally friendly process with low energy consumption to recover the metal content and recycle the acid from spent pickle baths. The metals are precipitated as metal fluorides, which can be turned into metal oxides and hydrofluoric acid by thermal hydrolysis. In the proposed process the solution is concentrated by nanofiltration and the metals are precipitated in a subsequent crystallization step. Different nanofiltration membranes have been evaluated in experimental batch tests. The membrane MPF-34 from Koch Membrane Systems offered both good flux and acceptable metal retention even after long periods of acid storage. The solubility of B-FeF3.3H20 has been measured at varying acid concentrations and at different temperatures. The solubility decreases with increasing hydrofluoric acid and nitric acid concentrations. This suggests that the crystallization process should be operated at high acid concentrations. The crystal growth rate of B-FeF3.3H2O is found to be quite low but increases significantly with temperature. This suggests that the crystallization process should be operated at higher temperatures (40 - 60°C). Within the project a pilot plant has been constructed where the spent pickle bath solution is concentrated by nanofiltration to the desired metal fluoride supersaturation, and B-FeF3.3H2O is crystallized in a 10 L continuous crystallizer.

Jan 1, 2006

By E. Krause

The oxidation of Fe(II) with S(IV) and oxygen can be used advantageously in leaching systems requiring both Fe(III) and sulfuric acid; e.g., heap leaching of sulfide-containing ores or concentrates. The oxidation of Fe(II) with SO2/air is several orders of magnitude faster than with air alone; rates of up to ?20 g/L'h were observed at 25°C. The reaction proceeds via an intermediate orange-red mixed-valence Fe(II)/Fe(III) sulfite complex, which is present at an essentially constant concentration throughout most of the oxidation process. The oxidation rate of Fe(II) is approximated by the formula -d[Fe(II)]/dt = k [S(IV)] [02]0'3. The concentration of the mixed-valence complex, and thus the iron oxidation rate, decreases with increasing acidity. Certain impurities, such as SCN- and thiourea, negatively affect the rate. The oxidation reaction 2Fe2 + SO2 + 02 4 2Fe3+ + S042- is accompanied by the side reaction H20 + SO2 + 1/202 4 H2SO4. The side reaction becomes particularly important at low temperatures and whenever the oxygen mass transfer is insufficient with respect to the rate of addition of S(IV). Thus, effective oxygen mass transfer is essential for achieving high oxidation rates with minimal acid formation. The activation energy was estimated as 141 kJ/mol over the temperature range 10-35°C.

Jan 1, 2006

By M. Medraj, M. A. Arafin, D. P. Turner, P. Bocher

In this study, the kinetics of isothermal solidification during TLP bonding of Inconel 625 and 718 superalloys with nickel based multi-component filler alloy BNi-2 have been studied through a combination of direct experimentation and mathematical modeling in order to determine the effect of process variables and, thus, to optimize the process. The predicted isothermal solidification completion times from the boron diffusion model and migrating solid/liquid interface model were in good agreement with the experimentally determined values and, they were found to be significantly less than those of other nickel superalloys with different nickel based filler alloys. Further, significant reduction of holding time was observed with increasing bonding temperature and with decreasing joint gap. No significant grain growth in the base metals has been observed in the temperature range being investigated (1325 - 1394K).

Jan 1, 2006

By Roger V. Pierce

IN the last few years, much study has been devoted to increasing stoping efficiency. The reasons for this are shortage of manpower, shorter working hours, operating regulations, and shortages of essential materials. More efficiency in mining methods is needed to offset these adverse factors. In the past, attention has been centred upon haulage, hoisting equipment, mill flow-sheets, etc., but recently the spotlight has been focused on the cost of stoping and the cost of handling materials. Scrapers and scraper hoists have thus assumed greater importance. In order to make it possible for certain orebodies to be mined at a profit, it has been necessary to apply assembly-line practices to the streamlining of mining methods. All cycles of the mining operation must balance, and the sooner broken ore is removed from the working face, the sooner another round can be started. Since the introduction of scraper hoists in metal mines more than thirty years ago, much progress has been made in their design and application. Alert operators are now constantly watching for methods in which to apply scrapers. The saving of manpower in development work and the changing of mining methods that once employed manual handling of materials are increasing vital wartime tonnages. Essential ores from newly discovered orebodies are moving rapidly from stopes to reduction plants-because it is now possible to develop an orebody and place it in production with a minimum of manpower and development work. Today, speed is one of the prime essentials o mining. Everywhere, there is a concentrated effort to use experienced miners in breaking rock, in timbering, and in carrying out the various phases of mining that require many months of training. Much effort is being made to mechanize the working places because experienced men naturally prefer to work where the majority of materials are handled mechanically. In line with this policy, many new men are being taught the details of scraper work, because even green men can quickly become acquainted with scrapers and thereby release older heads to perform the more skilled phases of mining.

Jan 1, 1943

By John Knox

The object of this paper is to present an analysis of the major problems encountered both in the design and application of downhole drills. Particular reference is made to design features governed by: physical size limitations; metallurgical material standards; physical operating environment; range of commercial materials to be drilled; high power output required, coupled with simplicity from the operating and maintenance viewpoint; and bit design. The history of attempts to develop downhole drills is discussed, and representative mechanical configurations are illustrated. The general design and application aspects of the different types are commented on. The application of this drill type in various commercial fields, such as (a) oil-well drilling, (b) open-pit mining and (c) underground mining, is discussed, along with the limitations presently imposed by the commercial state of the art in the manufacture of accessory equipment; e.g., drill rods, air compressors, mountings, etc. Suggestions on the possible future direction of the development and application of this tool are put forth.

Jan 1, 1965

By L. S. Rivett, U. M. Oko

The Strathcona mine, with a large new underground operation and concentrator, was recently brought into\ production by Falconbridge Nickel Mines Ltd. The tailings disposal and water utilization problems at Strathcona are described in this paper, with sections devoted to the Moose Lake watershed system, tailing impounds, the construction of dams and the operation of the Moose Lake neutralization plant for the elimination of acidity, including the equipment used.

Jan 1, 1971

By C. R. Larsen, R. G. Bradburn, A. L. Mular, B. C. Flintoff

"Following the acquisition of grinding- circuit data at Brenda Mines Limited for simulation purposes, various methods to mass balance the circuit were evaluated. Unfortunately, different answers were obtained for different balancing methods. A data-adjustment procedure was therefore developed. The procedure assumes that all ""best-fit"" criterion. A search is conducted with the minimum set of water masses and narrow size ranges of solids around the circuit (called search variables), which permit the computation of any of the measured data points. Search variables are used to calculate measured values for an objective function of the form: [ ] where W, is a weighting factor•, M1 is a measured value and M. is a corresponding . value calculated from search vanables. The best values of M, are those which minimize R. The above method is compared with others that have been utilized for mass balancing circuits."

Jan 1, 1976

By K. C. Lahiri, H. P. Das, G. G. Sarkar

"The petrographic components of any particular rank of coal often show some distinct physical and chemical characteristics. Therefore, their pro-portionate distribution in a coal greatly influences the technological behaviour of the coal.In any size or gravity separation of a coal, the petrographic constituents become segregated to different degrees in some preferred size or specific gravity range. Therefore, a knowledge of the petrographic nature of the washed and/or screened fractions is often considered essential in selecting the most efficient method for coal preparation and in exploring the proper evenues for its utilization.A graphical method is described in this paper to indicate the mineral-matter content and the petrographic components of the washed fractions of a coal at different ash levels of separation in the form of a new washability diagram. This diagram can be conveniently used to assess or predict the petrographic composition of clean coal or sinks corresponding to any desired ash level in the cleans. With the help of this diagram, it is also possible to locate the yield level of cleans at which maximum enrichment of vitrinites and exinites occurs with the least amount of mineral-matter content."

Jan 1, 1971

By Paul Johnston

or is it? Mining Automated Loader Technology Application ? Intelligent software mounted in hardware on a remotely controlled scoop. ? Led by INCO, Sudbury, with ? C-CORE, St. John?s ? Automated Mining Systems, Aurora ? Bell Canada, Sudbury ? Laurentian University, Sudbury ? University of Western Ontario, London Completed: March 2000

May 1, 2008

By T. R. Roy

"This paper presents an objective model for the estimation of fire risk potential of an underground coal mine panel by the bord-and-pillar method of mining where extraction is carried out by depillaring with caving. The fire risk rating procedure for the estimation of fire risk potential involved in the panel is described in detail. The model developed thereof is tested with the field data collected from different collieries across the country along with the laboratory data for critical assessment of fitness of the model. Some results showing the validity of the model are also cited here. Introduction The occurrence of fire in underground coal mine panels is quite common in India and elsewhere. These fires cause extensive loss of resources.Although substantial work has already been done by different research organizations, the problem has not been fully eliminated.The causes of panel fires depend not only on the spontaneous heating susceptibility of coal, but also on the extraneous conditions created in the panel due to the mining operations. From the studies done so far by various organizations, it is apparent that two types of parameters associated with spontaneous heating of a coal mine panel exist: intrinsic and extrinsic. Parameters related to the nature of coal, i.e., its physicochemical characteristics, petrographic distribution and mineral make-up are intrinsic, while the extrinsic parameters are related to the atmospheric, geological and mining conditions such as scope of leakage of air into the panel, wetness of mines, heat dissipation capacity of the panel, gassiness of seam, geological disturbances, method of extraction, size of panel, etc.Banerjee (1982) identified twenty-two mine parameters such as spontaneous heating susceptibility of coal, method of mining, geological disturbances, quality of overburden, degree of consolidation of coal mass, scope of accumulation of coal fines, etc. that can influence spontaneous heating in underground mines. He used the terms “high” and “low” for fire risk rating of each of these parameters. Since this approach is oversimplified, it can not determine the exact fire risk potential of an underground working."

Jan 1, 2003

By C. H. G. Bushell, G. Brown, C. J. Krauss

1N FLOTATION, copper sulphate is used universally for the activation of sphalerite and marmatite and, to some extent, for the activation of other minerals. The copper activated surface readily adsorbs standard collectors such as the xanthates and dialkyl dithiophosphates whereas untreated zir~ sulphide does not. With the host of reagents available, making a mineral float is seldom difficult. The problem is to achieve selective flotation. In zinc flotation, the great value of copper sulphate results from its ability to activate zinc sulphide preferentially to gangue minerals. At the Sullivan Concentrator of The Consolidated Mining and Smelting Company of Canada Limited, the most abundant gangue mineral in the lead-zinc ore is pyrrhotite, and the zinc mineral is marmatite of somewhat variable zinc to iron ratio. For this Cominco study, selectivity between marmatite and pyrrhotite was therefore the most interesting aspect of copper sulphate activation.

Jan 1, 1961