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By O. V. Suvorova, V. G. Minenko, D. V. Makarov, A. L. Samusev, E. A. Selivanova

"Saponite is a valuable marketable product widely used foundry, metallurgy and construction, in the chemical, food and consumer goods industries, in medicine, and in agriculture.The feasibility of producing of high-quality construction materials with enhanced physical-chemical and decorative characteristics from saponite-containing products has been given a scientific basis and experimentally validated. The compressive strength of the samples in the sintering temperature range of 800-1000 ?? varied between 35 and 80.9 MPa; the bending strength 8 and 26.7 MPa, which is superior to the high-grade ceramic bricks of the M-300 brand. It has been established that the saponite-containing product obtained using the electrochemical separation of process water has a high sorption capacitance relatively the heavy-metal ions such as nickel and copper (40.7 and 99.8 mg/g, respectively). As a result of sintering at 750 ?? the sorption capacity of the sorbent for the nickel and copper ions increased to 194.2 and 212.8. Using the XRD and SEM with energy-dispersive spectroscopy studies, we diagnosed the metal-containing phases and validated the metal sorption mechanism.This study opens up the possibility of practical realization of the saponite-containing product as high-quality building materials and heavy-metal sorbents while addressing the issue of process water treatment. INTRODUCTION The Lomonosov deposit of the Archangelsk diamondiferous province accommodates 10 kimberlite pipes. The anticipated volume of diamond-bearing ore and dump rock to be removed before reaching the open-pit mining limit is 300 mln t (Karpenko, 2009; Bezborodov, etal. 2003). Rock in the deposit is almost completely displaced by clay minerals, predominantly saponite, amounting to 90% in the vent facies (Karpenko, 2009)."

Jan 1, 2018

By Carlos de F. Gontijo

The flotation of quartz particles was studied over the particle size range from 0.5 µm to 1000 µm and for advancing water contact angles between 0º and 83º. Flotation was performed in a column and in a Rushton turbine cell. Particle contact angle threshold values, below which the particles could not be floated, were identified for the particle size range 0.5~1000 µm, under different hydrodynamic conditions. The flotation response of the particles, either in a column or in a mechanically agitated cell with a similar bubble size, was comparable. Turbulence plays a role, as does bubble-particle aggregate velocity and bubble size. The stability of the bubble-particle aggregate controls the maximum floatable particle size of coarse particles. For fine particles, the flotation limit is dictated by the energy required to rupture the intervening liquid film between the particle and bubble. Flotation of very fine and large particles is facilitated with small bubbles and high contact angles. These results greatly extend our earlier observations and theoretical predictions.

Jan 1, 2014

By Alexander E. Sobolev

The kinetics of oxidative dissolution of lead sulfide in solutions of nitric acid and sodium hypochlorite was studied using the rotating disk method and the factor experiment design. The dependences of the PbS dissolution rate on concentration of oxidant (HNO3 or NaOCl), temperature, disk rotation frequency, and duration of interaction were determined. The adequate kinetic models allowing calculating the specific dissolution rate at any combination of values of the influencing factors were obtained. As a result of the complex analysis of kinetic parameters, the modes of the each process proceeding were established, their macromechanisms were determined, and the limiting stages of the oxidative dissolution processes were revealed. The dissolution kinetics of lead sulfide in solutions of nitric acid was found to have several regimes. At HNO3 concentrations below 0.5 mol/L, the process proceeds in a mixed mode close to a diffusion-controlled. It is caused by low oxidative activity of HNO3 and considerable diffusional diffuculties at this acid concentration range. At the nitric acid concentrations more than 0.5 mol/L and less than 6.0 mol/L, the process was found to be kinetically controlled. The limiting stage of the interaction is the surface chemical reaction. The observed relatively high order of the dissolution rate with respect to the concentration of HNO3 was explained by the autocatalytic action of the acid reduction products. At nitric acid concentrations exceeding 6 mol/L, the process proceeds in a mixed mode close to a diffusion-controlled, and the dissolution rate sharply decreased that was explained by a decrease in the solubility of reaction products. When lead sulfide is dissolved in sodium hypochlorite solutions, at first time the amount of PbS transferred into solution remains close to zero. Gradually, as time goes, this amount begins to increase sharply. The dissolution rate rises and reaches the maximum value, further remaining constant. At the steady-state regime, the dissolution rate is independent of the duration of interaction and disk rotation frequency. The process was found to be kinetically controlled. The limiting stage is likely to be chemical reaction on the surface of solid lead sulfide. The obtained kinetic characteristics enlarge the database on hydrochemical oxidation kinetics of sulfides and can be used for the elaboration of the recommendations for the effective comprehensive exploitation of lead-bearing ores and concentrates. Keywords: lead sulfide, nitric acid, sodium hypochlorite, hydrochemical oxidation, dissolution kinetics

Sep 1, 2012

By Svetlana Krashenina, Nebojsa Petrovic, Oscar Lopez

"As easily mined gold deposits are becoming increasingly rare, gold producers are turning to the processing of ores with polymetallic and copper mineralization. These base metals can create metallurgical and waste handling/disposal challenges, which can render the project uneconomic. Base metals such as copper compete for the cyanide reagent used to extract gold, necessitating the need to increase cyanide volumes and consequently, cyanide consumption costs. Copper will also stabilize in tailings water as weak acid dissociable cyanide, a form that is both difficult and costly to destruct. SART (sulphidization-acidification-recycling-thickening) process technology can be applied to remove base metal interferences in gold projects. SART breaks the copper-cyanide complex and precipitates the copper as a saleable copper sulphide concentrate. With the copper removed, the cyanide can be regenerated as free cyanide for recycle to the gold recovery process. This improves cyanide utilization efficiency and mitigates copper presence in tailings. Additionally, copper recovery generates an incremental revenue source to further improve project economics. Engineering Dobersek GmbH together with BioteQ developed a SART and AVR testing program to recover cyanide and maximize operational efficiencies at a gold processing plant in Kazakhstan with high cyanide consumption in the leach circuit. Results from on-site testing demonstrated that up to 99.98% of copper can be removed and precipitated as copper sulphide and 60-90% of cyanide can be recovered from the AVR and recycled for leaching, reducing total overall cyanide use. This paper provides an overview of the testing methodology and results together with a discussion on the potential of SART to improve gold production operations to deliver both environmental and economic benefits."

Jan 1, 2014

Gold thiosulfate complex is difference with gold cyanide complex, which is easily adsorbed on activated carbon while gold thiosulfate complex cannot be satisfactorily recovered by activated car-bon. However, Gold thiosulfate could be recovered by activated carbon if cyanide is added into the thiosulfate solution. To avoid the cyanide solution, this paper will review several possible processes to recover the gold from thiosuldate solution: such as the cementation with metals, ion exchange resin and solvent extraction. The gold cementation with the powders of zinc, copper, iron and aluminium will be discussed. The advantage and disadvantage of each are compared. The consumption of thiosulfate and copper and feasibility of solution recycling will be discussed. The gold recovery by some ion exchange resins will be suggested. The processing conditions, its applications and the resin elution methods will be presented. The gold recovery by using solvent ex-traction will be reviewed. The processing conditions of solvent extraction process are discussed and compared.

Jan 1, 2014

By Leopoldo Gutierrez

Eight oil sand ores with the bitumen content ranging between 3.6 and 15 wt.% and the fines content (-44 micrometers size fraction) between 9 and 69 vol.% were subjected to the standard alkali extraction test used to measure the degree of oxidation of metallurgical coals. The results showed that there is a strong correlation between the absorbance of extracted solutions at a wavelength of 520 nm and the ratio of the fines content to the bitumen content. The latter factor is directly relevant to assessing the extractability of bitumen from oil sand ores. Bitumen extraction tests showed that the bitumen recovery decreased from 90 % to 30 % when the absorbance at 520 nm increased from 0.23 to 0.64. A comparison of the UV/visible spectra of solutions obtained from the alkali extraction tests with solutions of Aldrich Humic Acids suggest that low bitumen/high fines oil sand ores release substantial amounts of organic matter very similar, if not identical, to humic acids. In contrast, high-bitumen and low-fines ores do not release as much of the organic matter, and UV spectra of these leached substances are quite different from those of humic acids. Overall, the results show that chemical composition of bitumen in the tested ores varies greatly with ore type and oxidation, and that the alkali extraction test is a promising technique to quickly assess the extractability of bitumen from a given ore type.

Sep 1, 2012

By J. Menuge, G. Barker

"Liberation (QEM-SCAN) and surface chemical (time of flight secondary ion mass spectrometry -ToF-SIMS) analyses have been undertaken on plant Pb circuit flotation samples with the aim of defining the factors resulting in sub-optimal recovery. Lead flotation circuit feed, cleaner concentrate and tails samples were examined on a size-by-size basis using QEM-SCAN. Sulfide minerals accounted for approximately one third of the flotation feed with pyrite making up 16 wt.%, followed by sphalerite (13 wt.%) and galena (3 wt.%). Non-sulfide gangue minerals comprised calcium carbonates calcite (26 wt.%), dolomite (14 wt.%) with quartz, feldspars and barite all contributing > 5 wt.%. QEM-SCAN analysis indicated that liberation was not the primary cause of the poor flotation response. Principal component analysis was applied to the ToF-SIMS data derived from five samples (as for QEM-SCAN plus rougher concentrate and tails) to enable identification of correlated surface chemical components. Although a significant range of correlations were observed we focus here on those that show a discernible pattern throughout the float circuit. It was found that for the flotation feed the surface of galena were largely obscured but where Pb was observed there was a strong correlation to surface Fe. For the rougher and cleaner concentrates the galena surfaces were relatively clean with no correlation to Fe for the rougher concentrate and minor correlation in the cleaner concentrate. However, in the rougher and cleaner tails the clear correlation of galena with Fe was again apparent. These results suggest, in combination with prior studies of pyrite morphology, that the principal factor effecting grade is the rapid oxidation of the framboidal pyrite resulting in the contamination of the galena surface with a high loading of hydrophilic oxidised Fe species, thus depressing flotation."

Jan 1, 2014

By T. Sun, J. Kou, C. Sun, D. Gong

"HPGRs (high pressure grinding rollers) have demonstrated higher energy efficiency than conventional comminution equipment due to their application of particle-bed breakage mechanism. In the past decades, a large number of studies have been conducted to understand the characteristics of particle-bed breakage, however, few of them concerned the difference between single-particle and particle-bed breakage in compression comminution. This comparison is able to demonstrate possible advantages of particle-bed against single-particle in a breakage event due to the isolation of the effect of force loading mechanism. The present study investigated progeny fragment size and energy efficiency of these two breakage methods at variable energy inputs and particle sizes. The results show that at low energy inputs of around 0.25 kWh/t, the numbers of fines resulting from the two breakage methods are very similar. However, at higher energy inputs of around 1.0 or 2.0 kWh/t, particle-bed breakage generated significantly more fines. From the perspective of newly generated surface energy, particle-bed breakage is obviously more energy efficient than single-particle breakage at energy inputs between 1.0 and 2.0 kWh/t. It was also found that HPGRs demonstrate high energy utilization rate in generating new surface energy within their typical range of specific energy input. INTRODUCTION Rock comminution is an important and energy-intensive process in mining operations. Nowadays, decreasing head grade and mineral liberation size of ore deposits greatly increase the energy cost of comminution process. HPGRs have been successfully applied in the mining industry to improve the energy performance of comminution circuits. Typically, 10 to 30% of energy saving can be achieved due to the application of HPGR (Wang et al., 2013; Casteel, 2005; Günter et al., 1996)."

Jan 1, 2018

By Brian Hart, Liuyin Xia, Staman Dimov

The recovery of rare earth elements (REE) is fast becoming of significant interest globally. Modern research and development programs (outside China) towards the optimization of REE mineral recovery by flotation have been lacking. Due to the mineralogical nature of many hosting deposits, the separation and recovery flotation flowsheets can be very complex. As part of a research project, a selected suite of hydroxamic acid collectors (Naphthoyl, Cyclohexyl, Benzoyl, Octyl, Terta-butyl benzoyl, Salicylic) were tested in order to examine their potential for recovery and selectivity. Micro flotation tests were conducted on a feed sample, recovery and grade of the products were used to evaluate their relative flotation efficiency and selectivity and these results (flotation with the hydroxamate collectors) were compared to a phosphoric acid/sulphonate collector mix(shorten as "mixed"). In the testing, the flotation response with the hydroxamate collectors was significantly better than the "mixed" collectors on freshly ground (no slimes removed) samples. Furthermore the tests revealed that for most of the hydroxamate reagents tested, their dosage requirements to achieve recoveries at the same levels was less than that for the "mixed" reagents. Product assays show that recovery of both La and Ce is significantly better for using the salicylic hydroxamate and that Nb, Y and Zr recovery are favored by using octyl hydroxamate. These results indicate mineral flotation selectivity could be achieved by using selective hydroxamtes. Spectral fingerprinting by TOF-SIMS surface analysis has allowed for the identification of reagent species. TOF-SIMS surface analyses of grains analysis confirmed that REE bearing grains reporting to the concentrates are doing so in response to collector attachment whereas grains reporting to the tails are doing so in response to a lack of collector. The TOF-SIMS surface data bas been used to examine the factors promoting flotation stream partitioning.

Jan 1, 2014

There is a strong demand for nickel metal due to quick development of stainless steel and new electrical materials, most of which is extracted from sulfide resources, resulting in a serious depletion of sulfide resources. Therefore, more attentions have been paying to extracting nickel from nickel laterite, which accounts for some 50% of total nickel metal production. In this paper prereducing?magnetic separation?reduction smelting process is developed to manufacture ferronickel rom two types of nickel laterite DN and GN, assaying 1.42% 1.83% Ni and 49.09% 26.08%Fetotal, respectively. Mineralogy of two types of nickel laterite was determined and optimization of process parameters was conducted to manufacture ferronickel. Prereduction?magnetic concentration process is used to upgrade low grade nickel laterite firstly, two kinds of Ni-Fe concentrates being prepared of DN assaying 2.26% Ni and 69.45% Fetotal and GN containing 2.76% Ni and 38.00% Fetotal. Finally, two types of ferronickel were produced by reduction smelting of prereduced concentrates DN and GN, DN and GN ferronickel contains 2.97% Ni and 6.58%Ni, 89.25%Fetotal and 89.95%Fetotal, respectively. The total recoveries of Ni and Fe of the two ferronickel are beyond 90% and 80%, respectively. There is significant difference in grade of magnetic concentrates and ferronickel between DN and GN laterite due to their different mineralogy, the former belongs to limonitic laterite and the latter saponite one. There are some advantages for this process to manufacture ferronickel from nickel laterite, including effectively upgrading low grade nickel laterite by prereducing?magnetic separation technique, higher throughput and lower energy available for reduction smelting section due to some 40% gangue being rejected in pre-upgrading section, and much less metallurgical coke consumption by using non-coking coal as reductant for prereducing?magnetic separation?smelting process.

Jan 1, 2014

By G. Bulut, Ü. Yenial

"Heavy metals contamination in aqueous systems can be found in a variety of industries, in particularly mining, metal processing, finishing and plating. Water containing more than 0.5 ppm heavy metal considered as a toxic due to harmful effects on health. The major techniques employed for heavy metal removal from industrial wastewaters include precipitation, reduction, ion-exchange, and adsorption. Adsorption is a heavy metal removal method for aqueous systems. As a common adsorbent, activated carbon has been used in treatment but it is ineffective on some elements. Thus, there has been a research for developing highly effective and selective adsorbents. In this study the adsorption of lead and arsenic ions on to coal waste and red mud waste was studied. The experimental parameters were determined as pH, equilibrium time and the initial concentrations, activation by heat treatment and the blend of adsorbents. According to their adsorption capabilities, red mud was found to adsorb lead and arsenic ion with high efficiency but coal waste did not adsorb arsenic ion. However the binary adsorptions of these ions were enhanced the adsorption efficiencies. In addition the blend of coal waste and red mud affected the adsorption efficiency. INTRODUCTION As a result of rapidly developing industry and excessive use of water resources, the need of clean water and wastewater management appear as a problem. Wastewater recycling activities and researches come to the fore for an effective use of wastewater. Recycling of wastewater, which contains heavy metals harmful to the environment and human, is a difficult process. Heavy metals are generally divided into three categories: toxic metals, precious metals, radionuclides. Wastewaters from various industries, containing heavy metals, are the main polluters of water and soil. Although their toxic structures, heavy metals are used extensively in industry as they have technological significance. Usually mining industry, metal industry and industrial organizations produce heavy metal containing wastewaters (Volesky 1990, Ting and Lawson 1991, Bishop 2002)."

Jan 1, 2018

By Máximo Pérez, Julio C. Videla, Rosa Dzioba

"This paper reports the hydrometallurgical treatment of gravity concentrates of gold samples and outcrops associated with graphite ores obtained from:- flotation tailings of crystalline graphite minerals- outcrop samples of graphitic rock associated with auriferous minerals and free gold located in the Villa Union district and surrounding areas - La Rioja Province - Argentina.Previous studies of the flotation potential of crystalline graphite minerals failed to consider the presence of noble metals.The samples represent a mineralized zone of crystalline graphite which is distributed in different formations with a NE - SW trend in Maz, Punta Colorada and Las Ramaditas hills, covering an area of about 300 square kilometers.The most important variables affecting the leaching process, such as grain size, mineralogical species, reagent consumption, and so on, were previously studied. (Weiss, L.1985)On the basis of the present study results, the possible definition of an economically viable mineralization will be analyzed to define an auriferous district."

Jan 1, 2003

By Arkady Senchenko, Andrey Vasiliev, Alexander Aksenov, Metodi Zlatev

"This paper considers the issue of treating argillaceous oxidized ores with low gold grade (around 1.0 g/t). These ores according to their gold recovery level can be treated by heap leaching and agitation cyanide leaching, but low gold grade makes this treatment uneconomic. Heap leaching of argillaceous ores requires high consumption of binding agent for pelletizing to create suitable hydro-physical properties of material, which leads to increasing process cost and makes the material treatment unprofitable. Agitation cyanide leaching of this ore requires grinding processes; these expenses would not be paid back by the amount of gold recovered. Considering the above, poor argillaceous oxidized ores are recommended to be treated according to a combined technology including fractionation of ore, heap leaching of coarse fraction and agitation leaching of argillaceous fraction. Complex approach application allows one to avoid expenses incurred for material grinding and high consumption of binding agent for pelletizing. The success of this technology implementation will firstly depend on the quality of material fractionation, therefore the section of disintegration and division of ore into different components will be the most important part of flowsheet development for gold production. When disintegrating minerals at positive temperatures both standard (mechanical disintegration - scrubbers) and innovative (hydraulic method of disintegration - HydroClean units) equipment is applicable. HydroClean units are characterized by lower energy and water consumption. In a number of cases it is required to treat frozen ores. When treating frozen material standard approaches to disintegration are not reasonable due to their low efficiency. In order to treat frozen material a two stage disintegration flowsheet with screening after each stage is recommended. The first stage is recommended to be performed in a scrubber, the second - in a HydroClean unit. This technological flowsheet will provide high and steady values for ore separation into fractions at low energy consumption for disintegration."

Jan 1, 2014

By T. N. Alexandrova

Results of mineralogical-technological research of carbonaceous rocks of Kimkansky precious metal occurrence are presented, on their basis new typemorphic natural and technological microelement associations were revealed using mathematical statistics methods. Keywords: carbonaceous rocks, processing, microelements, flotation, statistical criteria

Sep 1, 2012

By E. V. Bratygin, S. P. Elnikova

"In the article carrying out technique of experiments for estimation of power efficiency of process of crushing in cone crushers «in a layer». The conformity of the granulometric characteristics of the crushed products obtained as a result of the tests, the theoretical equations for the size characteristics, and the choice of the optimal equation for the prediction of the granulometric characteristics of the final products obtained by crushing ""in a layer"". Optimization of granulometric composition of power of crushers working ""in a layer"" is carried out with the purpose of reduction of the electric power consumption at reception of a product of required quality. As a result of the experiments, dependencies are obtained, the relevance and applicability of the crushing process ""in the layer"" for obtaining crushed stone of improved quality is proved. INTRODUCTION A number of equations have been proposed for calculating the breakage energy. These include equations by Kirpichev and Kick, Rittinger and Bond (Bogdanov, Olevsky, 1982). Another equation for calculating the specific breakage work in breakage by impact and in stratum was proposed by Devyatkin (Boiko, 2003). In these equations, the breakage energy is proportional to the particle size raised to some power. In the equation proposed by Devyatkin, the ore piece thickness or the feed average diameter is used. The relationship between energy and breakage can be expressed as an equation (Lynch, 1981; Tikhonov, no. 3, 2008) where k is a coefficient characterizing the energy parameter depending on the strength of the feed, l is size in microns. The Rittinger formula is obtained for n = 2, the Kirpichev-Kick formula for n = 1, the Bond formula for n = 1.5. For n = 3 or higher, the equation describes fine and ultrafine grinding. R. Hookey proposed a formula (Lynch, 1981, Tikhonov, no. 3, 2008), where the size exponent n is substituted by a size function In all of the above equations, a single value of the feed and product size is used; for example, in the Bond equation, size is defined as the particle size of the 20% oversize (P80 and F80). However, a single size value can represent multiple distributions, making that the Bond method inappropriate for determining the size (Lynch, 1981). Describing an equation for estimating the grinding energy will allow to design the electric drive and ore preparation circuit for the given breakage method."

Jan 1, 2018

By Willis Forsling, Prabhakar Swarna, Hanumantha Rao Kota

The solubility of wollastonite and its Hallimond flotation behavior as a function of dodecylamine concentration and pH are examined. The flotation response is characterized by adsorption and surface energy in the presence of amine. The methods of UV-Visible and FT-IR spectroscopy for adsorption and sessile-drop contact angle for surface energy estimation are employed. The results of wollastonite dissolution clearly showed its non uniform dissolution. The surfaces are depleted with calcium ions and became rich in silicon within 10 min of flotation time. Maximum flotation of wollastonite corresponds to a half monolayer coverage of amine, which is substantiated by the surface energy parameters. A good correlation between flotation, adsorption and surface energy as a function of amine concentration and pH was observed. The sudden increase in wollastonite flotation and amine adsorption at and above 5x10-5 M initial amine concentration is thought to be due to two-dimensional precipitation of amine on surface in accordance with the recent 2D-3D precipitation model of amine adsorption on silicate minerals.

Jan 1, 2003

By Vincent C. Smith, Agit Singh, Bernard Muller

The control of minerals processing plants is by no means trivial. In order to address some of the difficulties, it was decided to develop and test a model predictive controller (MPC) for a milling circuit. A milling simulator was written to test the MPC controller. Due to the large number of tuning parameters in a controller of this type, an optimising algorithm was written that automatically tunes the controller. Good, robust control was obtained when the controller was tuned using this algorithm. The controller was tested for various scenarios that are often encountered in the minerals processing industry. When problems were encountered, the MPC algorithm was specially adapted to accommodate for these difficulties. Control was further improved by using feedforward control as well as quadratic programming. The movement of the manipulated variables was reduced by using a region of uncertainty that allows the controlled variable to vary within a specified region around the setpoint. Integrators were handled by adapting the algorithm slightly and adding an optional integral error portion in the MPC controller. The performance of the enhanced MPC controller was compared to PI controllers on both the simulator and a commercial plant. The MPC controller showed faster responses with less interaction in both cases.

Jan 1, 2003

By Gülay Bulut, Ünzile Yenial

"Water composition has a significant effect on flotation. A better understand the role of water composition in flotation is very important for process quality and performance. Based on source of process water, the concentrations of the dissolved ions, hardness, contaminants and residual composition of reagents can varied in the process water. The various water resources employed during lead and zinc flotation caused activation, depression or selectivity effects on minerals due to the presence of different ions in solution. Metal ions in these water resources were adsorbed on the ore, whilst ions associated with gangue minerals were released into the flotation pulp.This paper includes a research for the effect of different water compositions on lead and zinc sulphide flotation. Experiments were carried out five types of water such as tap water, recycling water, tailing dam water, underground mine water and sea water. The ore sample was obtained from Esan Balikesir-Balya mine. The mineralogical study of ore sample shows that it contains mainly galena, sphalerite, pyrite, hematite, quartz, diopside, epidote, calcite, and chlorite. It is found that recycling water and tailing dam water has a positive effect on recovery of lead and zinc minerals and detrimental effect on content of both minerals. This result may explain with remaining reagents in these waters. Underground mine water and tap water gave nearly the same recoveries of zinc. Sea water affects the recovery lead and zinc negatively."

Jan 1, 2014

By X. F. Cao

There was a large reserve of bauxite resources in China. And these resources were mainly the dias-pore-bauxite and contained plenty of aluminosilicate including kaolinite, illite and pyrophyllite. Reducing the contents of aluminosilicate would be significant to the using of Chinese bauxite resources. In this paper, from the single mineral flotation experiments, the surface properties and flotation beha-viours of aluminosilicate including kaolinite, illite and pyrophyllite, using dodecyl amine (DDA), N, N-dimethyl-dodecyl amine (DRN12) and N-dodecyl-1, 3-propane diamine (DN12) as collectors, had been investigated. The flotation results shown that the flotation behaviours of aluminosilicate had been influenced greatly by three amine collectors; and the effect was in the order of DN12>DRN12 >DDA. On the basis of FT-IR spectra and zeta potential measurements, the ionization of surface hydroxyl, as well as isomorphic exchange of surface ions, accounted for the mineral surface charges. And the mechanisms of interactions between the mineral surface and collectors mainly included two modes: electrostatic adsorption and hydrogen bonding; for the DDA and DRN>DDA. 12, the mechanism was mainly electrostatic adsorption; and for the DN12, the mechanisms of electrostatic adsorption and hydrogen bonding both existed. Results of quantum chemistry research shown that the collecting ability of amine collectors was greatly influenced by the electro-negativity of N in the amine molecules; and The |E(EHOMO| and ?E (HOMO-ELUMO) of N atom in DRN12 and DN12 were lower than that of DDA, and the positive charge of ?N group in DRN12 and DA12 were higher than that of DDA. All these calculation results could explain the collecting power of DN12>DRN12>DDA.

Jan 1, 2014

By M. Canbazoglu

In this study, optimum sulphuric acid leaching parameters of zinc and other valuable metals from two different waelz clinker samples were researched. Origins of the samples are electric arc furnace dust and original non-sulphite zinc ore waelz clinkers. The main zinc mineral determined as zincite in these samples. Collective dissolution of zinc and other valuable metals in waelz clinkers can easily achieved by sulphuric acid leaching of them. 96% of zinc can be dissolved at 80 ºC, 60 g/L sulphuric acid concentration, 60 min. leaching time and 1/5 solid/liquid ratio. Almost 90% of Cd, Cu and Co in clinker dissolved at these conditions while dissolution of Fe was almost 20%. The consumed acid amount for dissolved per ton of zinc is approximately 1.7 tons at these conditions.

Jan 1, 2014