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By G. Czupryna, S. Natansohn

Electrowinning is the prevalent technique for recovering silver from wastes generated in silver-tungsten electrical contact fabrication. Such scrap is placed in a permeable plastic basket, which constitutes the anode compartment, and is suspended in a larger cell where the cathode is mounted. A silver nitrate solution is used as the electrolyte and, upon application of an electric field, the silver migrates to the cathode where it deposits. At times, there is tungsten migration to the cathode compartment and consequent co-deposition with the silver. This W contamination, which may amount to several thousand mg/L (oz per gal), necessitates additional purification of the deposited Ag and its removal would, therefore, be advantageous. Eliminating the W impurity also extends the operating period of the cell with corresponding increases in the Ag output and savings in labor costs. The solution to this problem lies in the selective removal of the W from the AgNO3 electrolyte either by liquid-liquid or liquid-solid extraction. Although W extraction in the liquid-liquid mode with a quaternary amine or other reagents as extractants appears feasible, particularly in view of the low pH (about 1.5) of the system, this approach was not considered because it is essentially a batch-type process. It is a more elaborate industrial operation than extraction by ion exchange resins. Implementing the latter process consists of passing the electrolyte from the electrowinning cell through an ion exchange column which would extract the tungsten selectively and returning the tungsten-free stream to the cathode compartment. This process continues until the W extraction efficiency decreases (as a result of resin loading) below an acceptable value. The electrolyte is then directed through another, similar column while the W values are eluted off the first one and the resin is regenerated for another cycle. Thus, two columns in a parallel flow pattern provide the means for continuous processing of the electrolyte feedstock.

Jan 1, 1984

By Davood Moradkhani, Maryam Rasouli

"The present study reports solvent extraction and separation of zinc from concentrated nickel sulfate solution. In our previous study, Zn and Cd in cold filter cake (CFC, 45-55% Zn, 11-16% Cd, and 2.5-4% Ni) were extracted by alkaline leaching followed by electrowinning processes respectively. The strip solution from electrowinning process, was a pregnant liquor solution (PLS) containing 16.8 g/l Ni and 4.2 g/l Zn. In this work, PLS is subjected to the solvent extraction process for separation of traces of Zn from high grade nickel. The zinc free raffinate containing 16.6 g/l Ni sulfate solution will flow to nickel electrowinning plant for cathode production. Di-(2-ethylhexyl) phosphoric acid (D2EHPA) is the organic phase used in this study. The effects of various parameters such as organic phase concentration (10- 40 % v/v), pH (1.5-3), and temperature (25 and 45°) are investigated. Zinc over nickel separation factor increases by increasing pH from1.5 to 3. The results show that 98.7% of Zn was loaded at25°C O/A ratio of 1, pH 3, and 10% vol. D2EHPA. A 10 M NaOH is used for pH control. However, replacement of pH controlling agent with ammonia solution (25%. v/v) results in interesting observations. Ammonia addition, reduces the solubility of Ni2+ cations in solution. Nickel starts to precipitate from raffinate in a crystalline bluish-green form. The obtained crystals are characterized by XRF and identified to be nickel-bearing analog of boussingaultite, a rare sulfate mineral of nickel with chemical formula of Ni (NH4)2(SO4)2.6H2O. The crystallization rate of crystals is studied. It is a slow precipitation process with 75% in 4 days.INTRODUCTION Annually, many hydrometallurgical processes such as electrowinning, solvent extraction and leaching are applied to waste filter cakes in Zanjan, Iran. These additional processes on filter cakes are critical from both environmental and economic points of view. CFC (cold filter cake) is a toxic but valuable filter cake (45-55% Zn, 11-16% Cd, 2.5-4% Ni, and 0.2-1.5% Cu) produced in many zinc production plants. Our research group at Research and Engineering Company for Non-ferrous Metals (ReCo.), developed a series of hydrometallurgical methods for extraction of zinc, cadmium and nickel from CFC (Moradkhani, & Rasouli 2012). The recovery of zinc and cadmium with alkaline leaching and electrowinning was successfully studied. The strip sulfate solution from cadmium electrowinning unit, contains high concentration of nickel with some traces of zinc impurity. This pregnant liquor solution (PLS) was analyzed and reported to contain 16.8 g/l Ni and 4.2 g/l Zn. In this work, PLS is subjected to the solvent extraction process to separate Zn from nickel (Fig. 1). The organic phase stripping solution containing 4.2 g/l zinc and 200 ppm nickel can be circulated and used in the acidic leaching step. This generates nickel solution with high purity suitable for nickel cathode production."

Jan 1, 2017

By M. J. Collins

A hydrometallurgical process has been tested at the Sherritt Research Centre for separation of zinc, lead, copper and precious metal values from the zinc oxide fume leach residue stockpiled at the HBM&S operations in Flin Flon, Manitoba. Zinc is found primarily as zinc femte and sphalerite in the residue, lead as lead sulphate and jarosite, and copper as sulphide and jarositic compounds. Gypsum and siliceous gangue are significant diluents. The separation process includes hot acid, atmospheric and reducing (sulphur dioxide) leaching; zinc, copper and lead flotation; and atmospheric iron removal steps. Zinc is recovered to a high strength zinc sulphate solution and to a zinc sulphide concentrate suitable for feed to a roaster or to a zinc pressure leach plant; lead to a high grade lead sulphate concentrate suitable for feed to a lead smelter, and copper to a high grade copper sulphide concentrate suitable for feed to a copper smelter. Precious metals are divided between the zinc, lead and copper concentrates.

Jan 1, 1993

By E. A. Brocchi

This work is dedicated to the chemical treatment of a typical EAF dust by chlorine content agents. The residue contains 48% Fe2O3 and 28% ZnO mainly in the franklinite structural form. The others less compounds are: 4.00% CaO, 2.60% MnO, 2.20% MgO, 1.08% PbO, 0.36% Cr2O3, 0.21% B2O3, 0.03% CdO and 0.01% NiO. The experiments purpose is selectively to remove zinc from the residue leaving the iron behind. Thermodynamics studies have indicated some roasting possibilities. The experimental results have shown that both HCl leaching and carbochlorination will remove iron and zinc without any selectivity. However, either the chlorine or the calcium chloride roasting seems to form, preferentially, the zinc chloride. So, it will be presented how the EAF Dust can be treated by the using of chlorine roastings process such as carbochlorination in order to separate the zinc and iron present in the frankilinite, main EAF dust content.

Jan 1, 2014

By E. H. Cho, P. K. Dai

An acid mine drainage sludge sample, containing 3.6% Mn, 0.18% Ni, 0.15% Co and 0.39% Zn, was obtained from central West Virginia. The sample was leached at room temperature with a pH 3 sulfuric acid solution. Air was then bubbled through the leach solution at pH 4.5 to precipitate the iron. The leach solution contained 258 ppm Zn, 105 ppm Ni, 90 ppm Co and 2050 ppm Mn. Three solvent extraction studies were performed to selectively extract these metal ions in a sequential operation with DEHPA. Cyanex 272 and Lix 84. The A/O ratio was one, and the extractants were diluted ten times with kerosene. A study of solvent extraction with DEHPA showed that Zn could be extracted without Ni or Co up to a pH of 2.75. A pH of 2.75 is considered to be the optimum pH to maximize the selective separation of Zn from Mn. The separation factor of Co over manganese suggests that the two metal ions should be extracted together by Cyanex 272. The separation factor of Co over Ni indicates that these two metal ions can be separated most effectively around pH 5.3. Also, Ni can be effectively extracted by Lix 84 above pH 5.5.

Jan 1, 1995

By E. H. Cho

An acid mine drainage sludge sample containing 3.6% Mn, 0.18% Ni, 0.15% Co and 0.39% Zn was obtained from the central part of West Virginia and leached with sulfuric acid at pH 3 and at room temperature. Air was bubbled through the leach solution to precipitate the iron content from the solution at pH 4.5. The leach solution contained 258ppm Zn, 105ppm Ni, 90ppm Co and 2050ppm Mn. Three distinctive solvent extraction studies were performed to selectively extract these metal ions in a sequential operation with DEHPA, Cyanex 272, and then Lix 84. The AIO ratio was 1 and the extractants were diluted 10 times with kerosene. A study of solvent extraction with DEHPA showed that Zn could be extracted without Ni or Co up to pH 2.75. A pH of 2.75 is considered to be the optimum pH to maximize the selective separation of Zn from Mn. The separation factor of Co over manganese suggests that the two metal ions be extracted together by Cyanex 272. The separation factor of Co over Ni indicates that these two metal ions can be separated most effectively around pH 5.3. Also, Ni can be effectively extracted by Lix 84 above pH 5.5.

Jan 1, 1993

By J. He

The separation performance of a raw coal sample from South Africa based on the dense gas-solid fluidized bed beneficiation technique was investigated. Analysis showed that the raw coal sample distributes mainly in the -50+6 mm size range with the -1.6 g/cm3 density fraction accounting for 78.20% of the cumulative mass distribution. A flow sheet of the sequential coal beneficiation technique was introduced. The experimental and simulation results of the hydromechanical and density distribution stability indicate that the dense gas-solid fluidized bed could provide a stable separation environment with uniform density fluctuations under suitable operating conditions. With the application of this technique on the raw coal sample, the separation results show that three products, consisting of 73.25% clean coal with an ash content of 9.59%, 11.31% middlings, and 15.44% gangue were effectively separated at two different densities of 1.82 g/cm3 and 1.57 g/cm3. This study provides an effective flow sheet for achieving the high-efficiency separation performance on raw coal from arid areas. Keywords dense gas-solid fluidized bed, dry coal beneficiation, density distribution, separation efficiency.

Jul 1, 2013

By T. Suzuki

Acid leach liquors from spent lithium-ion batteries employing lithium cobalt dioxide (LiCoO2) as a cathode active material contain valuable metals such as cobalt and lithium, as well as common metals such as aluminum and copper. The objective of this study was to develop a hydrometallurgical process providing a high selectively between aluminum, cobalt, copper, and lithium from hydrochloric acid solutions. Batch experiment results showed that selective hydroxide precipitation approach was not suitable because a significant amount of cobalt and lithium was co-precipitated with aluminum and copper hydroxides. In contrast, batch experiments and the subsequent data analysis indicated that the sequence of solvent extraction circuits using PC-88A, Acorga M5640, and tri-n-octylamine (TOA) extractants effectively separates four metals of interest. In the first circuit of the proposed process, copper was selectively extracted by Acorga M5640 at the pH 1.0-2.0. Aluminum, cobalt, and lithium remain in the aqueous phase. In the second circuit, aluminium is selectively extracted by PC-88A at pH 2.0-2.5. For the separation of remaining cobalt and lithium, Acorga M5640 provided higher separation efficiency compared to PC-88A single extractant and PC-88A/TOA mixed extractant. However, the stripping efficiency of cobalt from the Acorga M5640 organic phase was found to be low (less than 10 %). Therefore, PC-88A/TOA mixed extractant rather than Acorga M5640 is an advisable choice for the separation of cobalt and lithium. Keywords: spent lithium-ion secondary batteries, solvent extraction, lithium, cobalt

Sep 1, 2012

By S. H. Lucas

"This paper describes the development and application of single-stage fluidized bed continuous ion exchange to the following processes: removal of base metals from mine water; cyanide removal from gold mill waste; uranium removal from sulphuric acid leach liquor; and uranium removal from mine water. IntroductionThe recovery of metals by fixed bed ion exchange has been practiced for many years. However, the operation has been semi-continuous and can only treat clarified liquors. Substantial savings could be made if the operation were continuous, thus reducing the resin inventory, and if the costly filtration and clarification steps could be eliminated. Work has been in progress over the past twenty-five years on the recovery of metals from unfiltered ore leach slurries, heap leach liquors and mine drainage waters containing a small amount of suspended solids. In the case of dilute waste solutions, with suspended matter and high flow rates, filtration is usually prohibitively expensive.Continuous ion exchange operation using fluidized beds has made it possible to reduce considerably the resin inventory and to allow the treatment of unclarified liquors and slurries. The subject has been reviewed on several occasions (1. 2. 3, 4). Although fluidized beds in multi-stage columns are an improvement over previous resin-in-pulp treatment, they have complicated internal structure and I or involved automatic control systems to move the resin stages.Slater and Prud'homme performed studies on multi-stage fluid beds, and concluded that the performance of single, deep fluidized beds of resin with continuous or semi-continuous flow of both phases would be worth investigating, particularly in the case of very dilute metal solutionsoi, A pilot plant was designed and built at CANMET in a cooperative project with Dr. Slateroi. The use of deep fluidized beds of resin (with a large height/diameter ratio) for treating dilute solutions has the attraction of engineering simplicity. Extreme simplicity of equipment design is considered of paramount importance when treating, for example, mine waters at high flow rates, possibly of the order of several million gallons per day. This paper describes the development of the concept of single-stage fluidized bed continuous ion exchange arid reviews the pilotplant development of processes for the treatment of base metal mine water, cyanide waste water, uranium leach liquor and uranium mine water."

Jan 1, 1980

By Donald S. Bubar

The Separation Rapids property, located approximately 60 km (37 miles) north of Kenora, Ontario, Canada, is host to one of the largest rare metal pegmatite deposits in the world. This pegmatite was named the Big Whopper by its discoverer, Dr. Fred Breaks of the Ontario Geological Survey, because of the exceptional length and width of the surface exposure. The property is readily accessible from a main, all-weather road via a network of secondary logging roads. The main line of the Canadian National Railway passes through the village of Redditt, just 40 km (24 miles) south of the property. The Big Whopper is one of the complex-type (petalite sub-type) class of rare metal pegmatites that are geochemically the most highly evolved in the spectrum of granitic pegmatites. Such deposits are economically important as resources for the rare metals lithium, tantalum, cesium, and rubidium. While complex-type pegmatites are found in many areas of the world, most are too small to be profitably mined, however, with an inferred resource in excess of 15 million metric tons (mt) (17 million short tons [st]), the Big Whopper is only the fourth deposit of its type in the world with the size required to be of major economic importance. The other three deposits, which are currently in production, are the Tanco mine in Manitoba, the Bikita mine in Zimbabwe, and the Greenbushes mine in Western Australia. The Big Whopper is situated in the eastern continuation of the Archean Bird River greenstone belt which also hosts the Tanco pegmatite, approximately 60 km (37 miles) to the west. The principal commodities identified in the portions of the Big Whopper pegmatite explored to date are petalite (LiAlSi4010) and rubidium-rich K-feldspar. These are industrial minerals with important applications in the glass and ceramics industries. The Big Whopper also contains substantial quantities of lepidolite, a lithium, rubidium-mica which is the principal ore mineral for rubidium metal. Tantalum and cesium occur in anomalous levels with the petalite mineralization, and the possibility of finding zones of high enrichment in these valuable high-tech metals elsewhere in the deposit is excellent. Such enriched zones are typical of pegmatites as highly evolved as the Big Whopper. To date, the geological mapping and diamond drilling work completed by Avalon have delineated the Big Whopper pegmatite system over a strike length exceeding 1.5 km (0.9 miles), over widths ranging from 10 m to 80 m (30 feet to 260 feet) in thickness, and to a vertical depth of close to 300 m (1,000 feet), where it remains open. The pegmatite system consists of a vertically oriented massive petalite pegmatite dyke striking 280 degrees that is flanked by amphibolites containing a swarm of narrower albite and petalite dykes which have all undergone intense deformation in a high strain zone resulting in folding and intense shearing. The Big Whopper exhibits a mineralogical zonation pattern which is complicated by the superimposed deformation that has both folded and stretched the deposit. Drilling has defined a dilute geological resource totaling 7.1 million mt (7.8 million st) grading 1.283 percent Li20, 0.346 percent Rb20, and .007 percent Ta205 over a strike length of 600 m (2,000 feet), and to a vertical depth of 250 m (820 feet), where it remains open. The petalite and rubidium-rich-K-feldspars contained in the Big Whopper all appear to be of superior quality. The grades are consistent with a petalite content averaging 25-30 percent and a rubidium-rich-K feldspar content averaging 15-20 percent. The remainder of the rock consists mainly of albite, several types of mica, and quartz. Accessory minerals include columbitetantalite, cassiterite, apatite, garnet, and gahnite. Market studies show that the deposit is located close enough to existing transportation infrastructure to access rapidly growing, major markets in the glass and ceramics industries, both in the northeastern U.S. and Europe. There are only three competing lithium minerals producers in the world with the Bikita mine in Zimbabwe being the only major producer of petalite. The other lithium deposits produce spodumene, a less desirable lithium mineral. Lakefield Research Limited has successfully designed a process to produce an ultra-pure petalite concentrate containing up to 4.65 percent lithium oxide, and as little as 0.014 percent iron oxide, low levels of soda and potash, and negligible amounts of other trace elements. With these specifications, Avalon will have an excellent quality product for glass-ceramics applications such as Corningware®, CERAN® stove tops, and other thermal shock-resistant products. Lakefield is presently designing the balance of the flowsheet to produce separate concentrates of rubidium-rich K-feldspar, albite, mica, tantalum, tin, and high-purity quartz. Initial test work indicates that the feldspar products will be of exceptional quality. The K-feldspar concentrates contain 11-12 percent K20 and 1 percent Rb20, while the sodaspar contains over 10 percent Na20, and both have very low iron (0.01 percent Fe203). A preliminary economic model shows that with an initial production rate of 90,000 mt (100,000 st) per year, expanding to 170,000 mt (190,000 st) per year over 5 years, the project is capable of generating over CDN $10 million per year in pre-tax cash flow by year 5 on a capital cost of approximately CDN $30 million.

Jan 1, 2001

By E. C. Goodhue, R. D. Coleman, I. E. Puddington, A. F. Sirianni

Iron ore samples containing apatite are fractionated to acceptable products by spherical agglomeration methods. The phosphorus is upgraded by a factor of 10 at an elevated pH. High-grade iron products are obtained in good yields from the tailings in a second step of the operation. The grade of the phosphorus separated appears to reach a maximum at about pH = 10; recovery of iron decreases below about pH = 4. Both the recovery and grade of iron are found to be reduced sharply above about pH = 8.

Jan 1, 1968

By Ahmed I. A. Salama

The problem of estimating separation characteristics by utilizing mass distributions of feed and products is considered. The separation characteristics include mass yield and partition curve. The cases of mass balance and partition functions balance are identified. The developed techniques can handle the cases of two- and three-product separation. Actual data are utilized to demonstrate the proposed techniques.

Jan 1, 1992

By B. Li

"Sepiolite is a magnesium-rich, porous, and fibrous clay mineral with broad applications. These applications take advantage of sepiolite’s high specific surface area and unique nano-sized channel structure. Over the past years, numerous materials and techniques based on the structure and property of sepiolite have been developed. This article reviewed the recent achievements in developments and applications of sepiolite in environmental protection, catalysis area, and nanocomposites. The prospectives of sepiolite applications is also discussed. INTRODUCTION Sepiolite is a magnesium-rich, porous, and fibrous clay mineral. The typical chemical formula of sepiolite is Si12Mg8O30[OH]4[OH2]4·8H2O [1]. Sepiolite has a high specific surface area and nano-sized channels because of its special structure [2]. This unusual crystal structure projected onto the [001] plane is shown in Fig.1 [3]. As it can be seen, silicon and oxygen atoms make up for the tetrahedral structure and magnesium and oxygen atoms form octahedral structure, respectively. Those tetrahedral-octahedral-tetrahedral ribbons expand boundlessly in one direction, leading to the special fibrous structure of the sepiolite. Every tetrahedral-octahedral-tetrahedral ribbon is combined to the next by an inverted Si–O–Si bond, which causes an interlaced layer with a continuous tetrahedral sheet and a discontinuous octahedral sheet. Terminal MgO at the edges of the ribbons is coordinated with two structural water molecules, which can form hydrogen bonds with zeolitic water molecules that are located in the channels running parallel to the fiber axis [3,4]. The Mg2+ ions inside the channels can be exchanged with other cations in solution. Therefore, sepiolite can be used to remove metal cations from aqueous solutions [5,6].It can be seen from Figure. 1 and Table 1 [7], the unique structure endows sepiolite with a substantial surface area and nanoscale channels. Its specific surface area reaches 320 m2/g and channel dimensions up to 3.7×10.6 Å. There are three active adsorption sites on the surface of sepiolite: one is oxygen atoms of the Si-O tetrahedron. The second is the structural water molecules paired with magnesium ions at the edges of the ribbons, which can form hydrogen bond with adsorbent. The last one is the Si-OH groups due to the breaks of the Si-O-Si bonds on the surface of Si-O tetrahedron. Thus, sepiolite has strong adsorption ability which has been widely used in environmental protection. Moreover, sepiolite has many significant features such as good mechanical and thermal stability, rheological property, porous, and easily modification, making it be a hotspot in chemical and biological researches. However, with the development of technology, broader and more advanced applications of sepiolite have been carried out in recent years. In this study, the recent achievement in sepiolite applications is summarized, the further tendency is also discussed."

Jan 1, 2018

By Patrice Favreau, R. F. Ph. D. Professor Emeritus Favreau

After Nobel invented dynamite, blasts were carried out by trial and error, without the use of equations based on the fundamental principles of Chemistry and Physics, although everyone accepted that shock waves were involved; but no equations were available to calculate their intensity. Around 1950 however, M.A. Cook proposed a procedure to calculate the thermochemistry of an explosive, and thus evaluate theenergy released during the explosion of a fully confined explosive. Some blasters then began to use theEnergy Factor; the latter, however, does not subdivide the energy into its three kinds of activities necessaryfor a successful blast, as demonstrated in 1968 by reference 2, namely i) the shock wave which weakensthe rock; ii) the stress field SS which fully fragments the rock; and iii) the release of the stress SS whichcauses the fully fragmented rock to burst out with the velocity necessary to move the rock into the muckpile. The article explains that the way to achieve a successful blast is by insuring that the explosive energy divides properly into the weakening action of the shock wave which is necessary to prevent large collar blocks and toe, and into adequate stress field intensity SS which is necessary to obtain good fragmentation size distribution at the crusher, as well as adequate burst out rock movement to achieve easy mucking.

By Favreau R. F. Ph. D, Favreau Patrice

The article explains that the way to achieve a successful blast is by insuring that the explosive energy divides properly into the weakening action of the shock wave which is necessary to prevent large collar blocks and toe, and into adequate stress field intensity SS which is necessary to obtain good fragmentation size distribution at the crusher, as well as adequate burst out rock movement to achieve easy mucking.

Feb 1, 2020

By George M. Fowler, J. P. LYDEN

T HE relationship of geological structure to orebodies and to the great masses of chert in the Tri-State mining district is of such significance that it prompts a brief recital of the existing information regarding the various periods of structural deformation. We began to note a relationship between structure and orebodies early in our studies of the field, and further work revealed a like relationship of structure to the great masses of chert so commonly associated with the ore-

Jan 1, 1934

By Guillermo V. Borquez

INTRODUCTION The basic factors that influence the selection of a mining method are the geometry of the ore deposit and the strength of the ore and surrounding rock. These factors usually determine the ore recovery and dilution of the deposit under analysis. For the purpose of this paper we will consider a generalized case of a block caving system and the sequence of the analysis needed to optimize ore recovery. The analysis presented is intended for the preliminary planning stage. The critical factors involved in the caving process are identified and the methods for evaluating the suitability of a particular deposit to the caving method of mining are outlined. The mine design is approached by identifying the two most important, dimensions: the size and spacing of the draw points. These dimensions are estimated from a simulation model which takes into account the ore recovery, the incremental value of the ore recovered, and the safety conditions of the operations. The model estimates the number of draw points required to extract efficiently and safely the broken ore from the caving area. This number is found in the range where the net incremental value of the ore is maximized and the probability of pillar failure does not exceed the limit which is acceptable for a safe mining operation. A mining method and its design are in a continuous process of optimization. This process is unique for each deposit because of the many parameters involved; therefore, a back analysis of the operations should be performed when possible. These analyses will indicate if the ore body responds in a predictable manner to the parameters contained in the geological, fragmentation, and draw point models. SELECTION OF ALTERNATIVE MINING METHODS The selection of a mining method best suited to define the minable ore reserves of a given deposit involves consideration of the geometry of the ore deposit and the strength of the ore and surrounding rock. These two factors are usually sufficient to determine the ore recovery and dilution of the deposit. After analyzing the geological reserves and defining the limits of the deposit, either an open pit or an underground mining method can be selected to define the minable ore reserves. The sequence of steps to follow in the selection of a mining method is presented in Figure I. From an economic point of view, the strip- ping ratio is the determining factor for the selection between an underground or open pit method. In turn, the stripping ratio is a function of the grade distribution and con- figuration of the ore body. In order to obtain a preliminary estimation of the stripping ratio, a parametric analysis could be applied, such as the approach used by Professor E. P. Pfleider (1972) and summarized in Table 1.

Jan 1, 1981

By J. D. Gillaspie, D. T. White

"Sulphuric acid atmospheric leaching of nickel saprolite / garnierite ores was proposed in the early twentieth century, prior to the development of the pressure acid leaching process for limonite ores. Sulphuric acid leaching with alkali addition for iron precipitation as jarosite has been considered periodically since the early 1970s, shortly after the development of the jarosite process for iron control in zinc refining. In the late 1990’s a series of flowsheets were proposed and tested involving simultaneous atmospheric leaching, iron precipitation and acid regeneration. These were considered particularly suited to sequential leaching of limonite and saprolite ores. A number of laterite sulphuric acid leaching processes utilising this chemistry have been put into commercial operation. Similar sequential leaching flow sheets involving precipitation of iron as goethite were also developed and piloted. This paper will firstly describe some of the important historical developments of both the jarosite and goethite processes. It will then provide descriptions of the commercialised processes involving alkali jarosite/alunite precipitation, with special emphasis on the downstream impurity removal requirements necessary for efficient recovery of nickel and cobalt from solution.INTRODUCTION This paper is a summary and extension of previous works with which the authors have been involved, on the subject of sulphuric acid leaching of nickel laterites (Adams, et al., 2004), (Liu H. , Gillaspie, Lewis, Neudorf, & Barnett, 2004), (White, Miller, & Napier, 2006), (White, 2009), (White & Gillaspie, 2013), (White & Gillaspie, 2015). Some of the historical developments of both the jarosite and goethite leaching-precipitation processes will be described, followed by a description of the commercialised sequential nickel laterite leaching processes, involving alkali jarosite/alunite precipitation and acid regeneration, with a comparison to the equivalent goethite based processes. These will include Jarosite Atmospheric Leaching (JAL), Goethite Atmospheric Leaching (GAL) and combined Pressure Acid Leaching – Atmospheric Leaching, termed EPAL®. There will also be an explanation of the downstream impurity removal requirements necessary for efficient recovery of nickel and cobalt from solution."

Jan 1, 2016

By Lazar Kricak, Zoran Teodorovis, Dragan Bogunovic

Scattering of delay interval around nominal value is an important issue in blasting practice. The scatter interval, as becomes wider, increases a probability of time delay overlapping adjacent blastholes. An excessive flyrock, ground vibrations, air blast and poor fragmentation are the symptoms for scattering related problems. Utilization of the short delays in the blastholes and longer delays on the surface are recommended solution to overcome mentioned problems. In contrast to global trends manifested through the application of electronic detonators and non-electric shock tube systems, revitalization of inexpensive electric detonators in combination with an accurate initiation system may be an acceptable solution. An extensive work was carried out at the Blasting Center Laboratory at the College of Mining and Geology, University of Belgrade, to develop a precise initiation system. As a result, Sequential Blasting Initiation System – Radio Frequency (SBIS-RF) was developed. It is expected that this system will increase blasting productivity, and reduce negative environmental impacts and total blasting costs.

Jan 1, 2008

By S. N. Bhatnagar, S. K. Haldar

"Mineral exploration is an expensive speculative venture and calls for optimization of sampling. Exploration is generally planned in grid pattern to be conducted in phased manner. The objective of any such campaign is to assess the quality, quantity and its reliability. The traditional estimation methods are unable to provide a means of calculating the degree of reliability. Application of mathematical models can quantify the precision of estimation and bring forth decision-making criteria at any stage.In this paper three statistical and geostatistical models, viz., Frequency, Probability, and Semi-variogram, have been employed to evaluate sequential exploration data of RampuraAgucha zinc-lead deposit. The optimization in exploration for a specific objective has been highlighted.IntroductionAny exploration campaign demands heavy resource investment starting from the collection of the first sample to the production of the first ton of ore. The best estimation would be if the entire orebody is taken out and sampled. However, this is neither possible, nor desirable. Drilling is the primary tool of geological sampling. It shares the major portion of the investment and, thereby, warrants critical analysis at every stage during the ongoing exploration projects.The exploration program should be conducted as a well planned scheme based on a sequential evaluation approach in a phased way to achieve a specific objective. It would then be possible to evaluate exploration data at the completion of each phase by employing computer-based mathematical techniques to quantify the precision of the estimate for fulfillment of objectives. The subsequent drilling program should be modified accordingly."

Jan 1, 1987