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By Qian Xu, Xionggang Lu, Xingli Zou, Zhongfu Zhou, Shuhua Geng, Hongwei Cheng

Herein we report an electrochemical route from oxides (SiO2, TiO2)/sulfide (MoS2)/carbon mixture precursors to Mo-Si-Ti-C composites in molten salt. The electrochemical process was carried out in molten CaCl2 at 800 °C with a potential of 3.1 V, and a graphite-based anode was used in the electroreduction process. The reaction mechanism and the influences of the stoichiometry of the initial mixture precursor on the morphology and phase composition of the final products were investigated. It has been found that Mo5Si3/MoSi2/Mo3Si/Mo, Mo5Si3/MoSi2/Mo2C/SiC, and Mo5Si3/Mo2C/Ti3SiC2/TiC/Mo composites can be electrosynthesized directly from their oxides/sulfide/carbon mixture. The preliminary experimental results suggest that the electroreduction process can be used for the controllable electrolytic production of alloys/carbide-containing composites from oxides/sulfides/carbon precursors.

Jan 1, 2016

Direct electrowinning is an electrochemical process which has been developed for the extractive recovery of copper from dilute, acidic leach solutions. In passing through a thin particulate coke cathode. copper is reduced and electrodeposited in a thin layer forming a solid copper-coke sheet. The product is melted in a furnace to separate the coke, then the molten copper is fire refined to produce a salable product. A description is given of a prototype unit which is designed to treat 20 gpm leach liquor and recover over 100 pounds of copper per day. Economic analyses show that in a commercial facility producing about 25.000 TPY copper.*the direct electrowinning process has lower overall cost than either conventional iron-cementation or the newer liquid ion exchange-electrowinning when the leach liquors contain less than 2 grams per liter of copper.

Jan 1, 1976

By Siemon J. R

It is possible to use compacted anodes formed at low pressure for the direct electrowinning of lead from sulphide concentrates. An anode containing 5.5 per cent graphite, formed into a compact at 34 MN m-2, enables lead to be won at a current efficiency of about 80 per cent; the power consumption .at 150 A m-2 current density being 0.4 kWh per kg lead. It would be necessary to use diaphragm cells to obtain a sufficiently pure product. A conceptual design for a full-scale plant is presented, together with a cost estimate for the production of lead by a process based on direct electrolysis of compacted concentrates. It is unlikely that the proposed process will be economically competitive with the conventional lead smelting.

Jan 1, 1976

By M Paakkunainen

Sampling is often the most critical part in process analysis. When the mean of a continuous lot, such as a process stream within a given time interval, is estimated by calculating the mean of several samples, the uncertainty of the estimated mean depends usually on the sampling strategy. Three different strategies can be used: systematic, stratified or random sample selection. The standard deviation of the lot mean is different for these methods so that systematic sampling usually gives the smallest and random sampling the largest standard deviation. Pierre Gy developed a method where the sampling errors of one-dimensional lots for these three strategies can be estimated from a variogram, which is calculated from a variographic experiment, ie from a set of one-dimensional time series measurements with fixed sampling interval. While variogram is an excellent tool for characterising and displaying the variability of autocorrelated processes as function of time, the sampling error estimation from it is quite laborious, however. In this work an alternative method for estimating the sampling errors was developed. The process is divided into strata of equal sizes, varying the length of strata similarly to variographic analysis. For stratified sampling the standard deviation estimates are obtained by analysis of variance and for systematic sampling by fitting a linear regression line to each stratum and calculating the residual variances. By pooling the individual variance estimates the variance estimates that are corrected for autocorrelation are obtained. The new method was tested with simulated and real data sets having periodic, random and/or drift fluctuations. For stationary processes the results agreed well with the results of the GyÆs variographic method. The method can probably be generalised to estimate the sampling variance of higher-dimensional lots.

Jan 1, 2005

By K. Turgeon, J. -F. Boulanger, L. Whitty-Léveille, C. Bazin

"Several companies targeting Rare Earth Elements (REE) around the world have completed a Preliminary Economic Assessment showing positive project financial results. Although REE ores may contain high REE value, financing these projects remains difficult in part due to the number of processes required to valorize the elements and the relatively low yield of these processes. Direct ore leaching without the production of a concentrate may reduce the capital and some operating expenditures through process simplification, and increase revenues because of REE recovery improvement.A silicate REE ore is used to test various hypotheses regarding whole ore processing and to develop a simplified flowsheet. An atmospheric leach (200 kg/t H2SO4) of crushed whole-ore containing 0,6% REE, followed by impurity removal and REE precipitation recovers 80% of the REEs in a solution containing REE and Th. The high value contained in most REE deposits makes whole-ore approaches worth considering for most projects, given the losses of valuable species typically associated with the production of a REE mineral concentrate from the ore. INTRODUCTIONRare Earth Elements (REE) find applications in modern and clean technologies, whether civil, industrial or military. With China producing more than 90% of the world’s REE, the embargo imposed by this country in early 2010 increased REE prices for a three year period. With prices having shrunk back many-fold from their peak value, projects put forward during the hike are struggling to secure financing. Owing in part to unconventional processes such as acid-baking and solvent extraction, REE plants are high capital cost and high operating-cost units, and present technological risks (Verbaan et al., 2014)."

Jan 1, 2016

By T. Watanabe, T. Fujiwara, S-Wan Song, Y. Nakagawa, R. Teranishi, M. Yoshimura

"We have tried a new method to fabricate cobalt oxide films onto porous films directly by an electrodeposition method at low temperature. The porous substrate films (membrane-filters and paper) were placed between a LiOH solution and a CoSO4 solution and treated under electricalcurrents of ~1 mA/cm2 using carbon cathode and anode. Effects of temperatures, flow rates and electrical currents have been investigated. As a result, we found that the LiCoO2 films and Co3O4 films were successfully fabricated on paper and/or porous PTFE filters at 120 °C, while CoOOH films were deposited at 80 °C. Furthermore, LiCoO2 films did not cover whole the substrates, but was fabricated just over the anode. The chemical reactions of electrodeposition were also considered from these results.IntroductionRecent agreements on global environment require rapid progress of low energy and low waste processings for the fabrication of thin films that succeed to the mass production with the high consumption of the fossil fuels (such as physical vapor deposition°, chemical vapor deposition2), dip coating with firing3) etc.). The total energy consumption in aqueous systems should be the lowest among all these processing routes, because a much lager excess of energy is necessary to create melts, vapor, gas, or plasma than to form an aqueous solution at the same temperature4)The ""Soft Solution-Process"" (SSP) was one of the best concepts to realize the low energy and low waste fabrication of inorganic materials. 5) Generally speaking, SSP gives similar results to any other processing using fluids (such as vapor, gas plasma) and/or beam/vacuum processing. However, the total energy consumption of all these possessing routes should be the lowest in aqueous systems, because a much larger excess of energy is necessary to create melts, vapor, gas or plasma than to form an aqueous solution at the same temperature (Fig. 1). This idea can be easily demonstrated, using the simple example of BaTiO3. The energy diagram for the formation of BaTiO3 from various precursors is shown in reference 4. The driving forces (AG) for representative synthesis reaction of (1), (2), (3), (4) and(5) are 38, 727, 3685, 17 and —14kcal/mol, respectively at room temperature. 6' 7"

Jan 1, 2000

By J. E. Dickinson, R. Campbell, M. J. Cole, K. P. Galvin, M. C. Cole

It is common practice in the beneficiation of iron ore ultrafines to apply reverse flotation, and hence float the gangue minerals, leaving behind a high-grade iron ore product. Given the difficulty in floating particles finer than 20 μm, especially those finer than 10 μm, it is necessary to subject the feed to desliming ahead of flotation. There are two problems with this approach. Firstly, given the known inefficiencies of cyclones, desliming needs to be aggressively applied to particle sizes upwards of 45 μm to ensure all the offending ultrafine gangue is removed. Secondly, desliming also removes a significant portion of the valuable iron ore. Thus, a significant portion of the ultrafines, including the iron ore, is lost to the tailings. This paper examines the direct flotation of iron ore fines using the Reflux Flotation Cell. The RFC operates using a concentrated bubbly zone, in the absence of a conventional froth zone, for delivering strong counter current washing through an inverted fluidised bed of bubbles ahead of the product discharge. A system of parallel inclined channels in the lower part of the cell provides enhanced segregation of the bubbles from the tailings, preventing the loss of air bubbles into the tailings. Experiments conducted under semi-batch conditions produced encouraging results with concentrates of ~65% Fe grade in a single flotation stage. Comparatively, up to five successive batch flotation stages were required when using a laboratory scale batch mechanical flotation cell to achieve similar Fe grades. For a given grade, lower Fe recoveries were typically achieved by the RFC. This was attributed to the effective prevention of ultrafine particles succumbing to hydraulic entrainment under a positive bias flux, as well as an increasing feed volume involved in the semi-batch operation using inverted fluidisation water for cleaning. Overall, similar grade-recovery curves were produced, however, significantly higher grades were obtained by the RFC and with fewer separation stages. Keywords: Desliming flotation, iron ore, modified release analysis, reflux flotation cell

Jan 1, 2020

By J. L. Huiatt, D. G. Foot

The Bureau of Mines devised a direct flotation method for treating low-grade carnallite (KCl*MgC12*6H20) and sylvinite (a mixture of KC1 and NaCl) ores characterized by a high content of water-insoluble impurities. The procedure consists of flocculation-depression of the insoluble slimes; decomposition leach (carnallite only); potash rougher, cleaner, and recleaner flotation; and final product leaching. The technique eliminates the necessity of removing the water-insoluble slimes by mechanical or flotation desliming. Laboratory results demonstrated that the direct flotation procedure, when compared with conventional techniques, yielded equivalent or improved flotation results with equal or less reagent consumptions. The direct flotation method recovered 81.6% of the solid phase KC1 from camallite ore and 79.7% of the KCl from sylvinite ore in final products containing 60.5 and 60.9% K2O, respectively. A flowsheet incorporating the described flotation procedure is presented.

Jan 1, 1983

By Donald G. Foot

The Bureau of Mines, U.S. Department of the Interior, devised a direct flotation method for treating low-grade carnallice (KC1.MgC12. 6H20) and sylvinite (a mixture of KC1 and NaC1) ores characterized by a high content of water-insoluble impurities. The procedure consists of (1) flocculation-depression of the insoluble slimes, (2) decomposition leach (carnallite only), (3) potash rougher, cleaner and recleaner flotation, and (4) final product leaching. The technique eliminates the necessity of removing the water-insoluble slimes by mechanical or flotation desliming. Laboratory results demonstrated that the direct flotation procedure, when compared with conventional techniques, yielded equivalent or improved flotation results with equal or less reagent consumptions. The direct flotation method recovered 81.6% of the solid-phase KC1 from carnallite ore and 79.7% of the KC1 from sylvinite ore in final products containing 60.5% and 60.9% K20, respectively. A flowsheet incorporating the described flotation procedure is presented.

Jan 1, 1981

By D Feng, G C. Lukey, J S. J, R R. Dagastine

An ongoing challenge in the coal industry is pyrite removal using the flotation method for fine particles. The use of a æseparation in frothÆ (SIF) technique to recover coarse particles of up to a few millimetres in size has presented a potential solution. However the commercial success of SIF is still limited due to the poor understanding of the separation mechanism. The particle floatability in SIF is determined by the constructive or destructive stresses resulting from the forces exerted by the particle on the bubble films. Hence, the surfactant properties and the mineral characteristics (eg hydrophobicity, shape, etc) are the key factors that control the interaction between the particle and froth. In the present work, the interaction forces between different particles (ie crushed glass, pyrite and coal) and a single air bubble are measured in a non-ionic surfactant solution (pentaethylene glycol mono n-dodecyl ether, C12E5) using an atomic force microscope (AFM). Experimental results show that the interaction is primarily dependent on the contact angle (ie the degree of hydrophobicity of the minerals). For crushed glass, the transition from repulsion to attraction was observed as the contact angle increased. As the concentration of C12E5 increased, the adhesion force was suppressed due to wrapping effect. For coal and pyrite with much higher hydrophobicity, long-range attraction was observed. The effect of shape factor was only visible at high surfactant coverage where sign of destabilisation became noticeable on the retracting force curve. These results provide a fundamental understanding of the direct interaction forces of a complex real system between an irregular particle of either crushed glass, pyrite or coal and a single air bubble. Theories developed for interpretation and modelling are thus far based upon the study of ideal systems of glass or polystyrene spheres and an air bubble and were incapable of explaining the interactions in complex real systems. The presented measurements were a first approach to the problem of real mineral-bubble interaction in flotation systems. Thus, the results can only best be interpreted semi-quantitatively.

Jan 1, 2005

By R-H Yoon

Increasing the upper particle size limit of flotation would allow operators to increase mill throughput, save operating costs, simplify flow sheets, and increase rate of return. However, coarse particle flotation is difficult for various reasons, of which bubble-particle detachment has been identified as the major limiting factor. In the present work, the process of bubble-particle detachment has been studied by directly measuring the forces required to detach air bubbles from hydrophobic solids. The measurements were conducted using a hanging balance apparatus. The measured forces were converted to detachment energies that can be compared with the work of adhesion, which in turn is a function of hydrophobicity, surface tension and contact area. The results were converted to detachment energies using the Young-Laplace equation for capillarity. In the present work, the solid surfaces were hydrophobised by coating them with octadecyltrichlorosilane and cetyltrimethylammonium bromide. The results show that bubble-particle detachment occurs in two steps. In the first step, air bubbles are deformed, while the three-phase contact (TPC) line is pinned, until contact angle reaches advancing angle. In the second step, the TPC line moves along the surface until solid/air interface is replaced by solid-liquid interface. In general, the energies of detachment measured in the second step are in good agreement with calculated work of adhesion. Thus, the energy required during the first step effectively represents an energy barrier for bubble-particle detachment. It has been found that the energy barrier varies with surface roughness and surface tension. The results of the present investigation suggest, therefore, that coarse particle flotation can be improved by control of particle hydrophobicity, surface roughness, and surface tension.

Jan 1, 2005

Using a Mark IV surface force apparatus, adsorption of dodecylamine hydrochloride on mica has been studied at pH 5.7 in the presence and absence of octanol. With amine alone, the repulsive forces between mica surfaces are reduced due to charge neutralization and attractive hydrophobic force; however, the latter is of relatively short-range with its decay length less than 2 nm. In the presence of amine and octanol, the hydrophobic force becomes more significant, its decay length being extended up to 6.5 nm. The increased hydrophobicity in the presence of octanol is attributed to the co-adsorption of the neutral surfactant with dodecylamine, which increases the packing density of hydrocarbons on the mica surface. The increased hydrophobicity due to co-adsorption is reflected in the improved floatability of silica when using octanol as an extender.

Jan 1, 1993

By K. A. Foo

Minproc Resources Inc. is developing a tailings retreatment project in Bolivia at the Colquiri zinc-tin-silver mine, in the Bolivian Andes. Mining at Colquiri has been continuous since 1940 and COMIBOL, the Bolivian state mining company, has owned and operated the underground mine since 1964. There are 10 million tonnes of tailings at Colquiri containing 0.51% Sn, 4.21% Zn and 50 g/t Ag. Minproc plans to mine and concentrate the tailings at 1.15 million tonnes per year to produce feed for pyrometallurgical upgrading. The tailings are moderately complex, with cassiterite grain size ranging from 200 to 5 pm. The marmatite (ZnFeS) in the tailings is relatively easily recovered, with zinc recovery of 80% to a 40%-45% Zn concentrate. The major problem with the zinc metallurgy is that the marmatite is complexed with iron, which results in iron contents of 14%-17% Fe in concentrate. This makes the product very difficult to sell because of the low Zn:Fe ratio. Most zinc smelters require a Zn:Fe ratio of 6:1 or 7:1, whereas this concentrate has a 3:1 ratio. Concentrate sale terms are therefore unfavorable, and the high cost of transportation from Bolivia to the smelters is an additional burden.

Jan 1, 1992

By Iley J. D

The production of zinc from low grade concentrates, particularly those high in iron, has a poor economic return due to a combination of higher concentrate transport costs, increased processing costs and an increased cost for disposal of iron residues. The production of iron-bearing residues from the electrolytic zinc process is also seen as undesirable for both cost and environmental reasons. This paper details the pilot plant testwork conducted at the Pasminco Research Centre in conjunction with Minproc for the production of zinc oxide fume from a high-iron zinc sulphide concentrate from Colquiri in Bolivia. The concentrate is derived from a tailings retreatment operation and has a high iron content (40 - 45 per cent Zn, 14 - 17 per cent Fe) due to the presence of marmatite. Pilot plant trials were conducted using the slag bath reactor at Pasminco Research Centre at rates up to 500 kg/h of pelletised zinc concentrate, using a natural gas/oxygen-enriched air lance. The fume produced averaged 76 per cent Zn at recoveries up to 96 per cent and was recovered in a baghouse after gas cooling. The iron was oxidised into a fayalitic slag that passed the TCLP test (toxicity characteristic leaching procedure) and the sulphur was oxidised to produce a high SO2 gas which would be suitable for sulphuric acid production. The process produces a zinc oxide fume that can replace calcine as a feed for electrolytic zinc plants. The iron reports to a stable slag for disposal, thus eliminating the need for its disposal as jarosite from the leach plant. Only a simple leaching operation on the fume is required as there is no zinc ferrite present.

Jan 1, 1993

Those who have had an occasion to apply geostatistics to the estimation of ore reserves are familiar with the conventional approaches of estimating average grade of quality values for arrays of regular grid squares or blocks. Further, the technology to perform geostatistical analysis and prediction for irregular two-dimensional areas has been available in a practical, readily useable form for some time. However, the generalization of this technology into three dimensions has been hampered by lack of a suitable geometrical characterization for realistic, three dimensional shapes and volumes. The extension of these estimation techniques into three dimensions has now been made possible by a successful integration of the volume modelling capabilities of 3D Component Modelling with proven 3D Geostatistical Techniques. The result is a powerful new technique for mineral ore reserves estimation in a practical, useable form. The technique allows direct geostatistical estimation of grade and quality values for precisely defined, irregular, realistic, geological, and mining shapes or volumes. This paper describes the two technologies, their integration and their application in practice to the estimation of mineral ore reserves. The paper closes with a discussion of the benefits provided by the new approach, in terms of accuracy, efficiency and practicality.

Jan 1, 1991

By Y V. Haarlem, P Coghill, N G. Cutmore, M Millen, I Carter

Due to the low concentration of gold, ore sorting for gold has always required measuring indirect properties, such as density or colour, to allow separation on a rock by rock basis. This only works on an ad hoc, by deposit basis and is likely to be least reliable where it is most required in low grade ores. One candidate technique, Gamma Activation Analysis (GAA) has been demonstrated to directly measure gold extremely accurately in powder samples of around 500gm and is used in the Chrysos PhotonAssay assay system. This method uses a linear accelerator (linac) to generate the X-ray radiation that activates the ore at low levels to emit a characteristic gold gamma ray that is then measured by a radiation detector. CSIRO is adapting this measurement technique to application on a belt at up to 800tph to measure pods of ore. The principal differences between measuring samples and an on-line measurement of pods are the requirement to measure coarse ore (<50mm), to measure a larger volume of ore as uniformly as possible, and to have an open-ended geometry with adequate shielding of the linac radiation source. These issues have all been addressed with measurements of coarse ore reported down to around 1ppm in sub-optimal test conditions. They have been demonstrated for <75mm particles with detectors capable of measuring a large volume geometry continuously. These are projected to allow measurements down to 0.3ppm in more optimum conditions for a 10 second measurement time period where the U/Th content of the ore is less than ~10ppm. Extensive Monte Carlo modelling has demonstrated that adequate shielding can be provided in relatively simple geometries with only 1 right angle transfer in the material handling. Steady progress is being made to reach the technical position required to trial an ore sorting system.

Nov 10, 2020

By John D. VandeKraats

Yielding rock bolt support systems have been developed to accommodate ground movement in shifting ground such as in coal operations; in creeping ground such as salt, trona, and potash; and in swelling ground associated with some clays. These systems, designed to remain intact despite ground movement, should enhance mine safety and help contain costs in areas where rebolting of rigid non-yielding systems is typically required. Four such systems were tested in straight tensile pulls in the laboratory. They include the Slip Nut System from Dywidag Systems International USA, Inc., Ischebeck&apos;s bolt mounted Titan Load Indicator, Rocky Mountain Bolt Company&apos;s Yielding Cable Bolt, and a rock bolt installed variation of the yielding steel post developed by RE/SPEC Inc. The first two systems are currently marketed products and the latter two are prototype systems. Each system responds to load and displacement by yielding in a unique manner. All are designed to yield at predetermined loads. A description of each system and its yield function is provided. Each system was tested over its prescribed yield range in a test machine. At least five tests were performed on each system. Each system yielded and continued to provide support according to its design. Each shows promise for ground control use in shifting or creeping rock. This work helps to illustrate the comparative differences in performance between these specialized systems and the applications where they may be most useful.

Jan 1, 1996

By Kurt Svens

"As zinc consumption is gradually increasing all the time in the world, many zinc plants are looking for possible expansions. As the sulphuric acid market as well as the fertilizer market in many areas is saturated, the costs for the roasting-acid plant section of the zinc plant will create problems when making feasibility calculations for possible expansions. For solving the customers' situation, two proven zinc concentrate direct leaching technologies are available. A third one, which is possible, does not have any industrially proven applications in the zinc area.IntroductionTime to time, the zinc plants in the world scrutinize their operations with respect to their cost efficiency, the total zinc recovery and the near-by or far-away consumer markets, especially the for zinc and sulphuric acid. Their preliminary profitability calculations can show that a possible total expansion is not economical, when they are looking to expand the whole roasting/acid plant - leaching - solution purification - electrowinning - casting route, as the possible market price for the sulphuric acid is low or the sulphuric acid market is too far away. Based on these facts, their economical study often indicates that implementing a zinc concentrate direct leaching option could be more favourable. Regarding this situation, several aspects concerning three alternative direct leaching methods will be discussed, of which pressure leaching and the atmospheric leaching processes have several industrial applications in the zinc industry, as the ultrafine grinding - conventional leaching process at present has none. As iron is also dissolved during zinc concentrate direct leaching, iron removal processes generally used in the zinc industry are additionally discussed below."

Jan 1, 2012

By Sharma T

Laboratory investigations have been carried out on direct leaching of lead-zinc complex sulphide ore and zinc concentrate using ferric chloride. From the investigation it was observed that the use of 0.15 M and 2 M FeC13 gave an optimum dissolution of metal ions from ore and concentrate respectively. The kinetic studies conducted showed that the leaching reaction followed a shrinking core model for galena and sphalerite. The presence of impurities such as pyrite had increased the dissolution rate and reduced the consumption of ferric chloride. Further, the studies were conducted to recover dissolved zinc from the chloride solution by selectively precipitating it as zinc ammonium chloride.

Jan 1, 1992

By R. W. Adams

A new process has been developed for leaching zinc sulphide concentrates at atmospheric pressure using air or oxygen plus nitrogen dioxide in dissolved or gaseous form as? an oxygen carrier. The process is flexible and can cope with variations in the degrees of retention and regeneration of dissolved nitrogen dioxide in the leaching phase. By controlled addition of concentrate a leach solution with low residual nitrate concentration can be produced. Following separation by filtration of unleached solids and elementa1 sulphur and conventional oxidation/precipitation of dissolved iron, the leach liquor can be treated with metallic zinc to remove the remaining nitrate and other meta111c impurities such as copper, cadmium and cobalt to provide a solution suitable for electrolytic recovery of zinc. Pilot scale testing of the process .has given zinc recoveries in the high nlnet1es, production of a hi.gh grade jarosite-free Pb/Ag residue (50% Pb.) and a saleable high grade of elemental sulphur. .A wide range of concentrates can be treated containing substantial levels of lead and copper as well as zinc. Indicative cost estimates suggest that the new process is competitive with alternative procedures such as pressure leaching.

Jan 1, 1990