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By Wenying Liu, Marek Pawlik

Long-chain ionic surfactants, with fatty acids and amines as the most important representatives, are widely used as collectors in froth flotation. These surfactants, generally referred to as weak electrolyte collectors, frequently form solid insoluble precipitates within a certain pH range. However, the action and behavior of these colloidal precipitates in froth flotation systems are not well understood and has not been well decoupled from those of dissolved forms. The objective of this study was to understand and compare the role of colloidal precipitates with that of the dissolved forms at gas-liquid and gas-liquid-solid interfaces. Using primary amine as the model surfactant, experiments were carried out to measure: the maximum volume of bubbles generated in the presence of dissolved and precipitated forms of the surfactant; surface tension for both forms of the surfactant using the drop shape and maximum bubble pressure methods; and contact angle using the sessile drop method. The experimental results showed that: in the presence of the precipitate, bubble volume was smaller, which was correlated with the lower surface tension and higher contact angle measured; the surface tension of solutions containing surfactant precipitate was affected by the surface age of liquid droplets. This study enhances understanding of the interfacial behavior of long-chain ionic surfactants under different conditions, which would help improve the efficiency of reagent utilization in froth flotation.

Jan 1, 2014

By B. Aydin, G. Erturk, G. Bulut

Coal is a complex material because of its structure consisting of carbon constituents and mineral matter. Flotation is readily carried out due to hydrophobic characteristic of coal. Therefore coal flotation is the most commonly used technique for fine coal beneficiation in order to acquire clean coal separating ash forming mineral matter from carbonaceous matter. In the coal flotation, floatability of coal both depends on characteristic of coal and used flotation reagent. Frothers as a flotation reagent are used to facilitate air dispersion into fine bubbles by prevention of bubble coalescence in the pulp phase, and stabilization of flotation froths. In this study, the effects of frother types and concentrations at different pH and particle size on the coal flotation were investigated. The representative sample was taken from Zonguldak, Turkey. It contains 9.1% ash and 7313 (kcal/kg) calorific value. Pine oil, Dowfroth 250, MIBC and Isooctanol were tested as frother types. It is determined that 1.92 ppm frother concentration is sufficient to float coal, except pine oil. In mixed frother tests, it is seen that Dowfroth 250 is more effective than others. No significant difference in coal recovery was observed for flotation tests at the different pH values. The best results are obtained in size of coal (-0.1 mm) with all frothers.

Jan 1, 2014

By M. Gonzalez Fernandez

"If the final goal of metallurgy is the extraction of metals from a given ore, everyone knows that this achievement relies on many separation steps involving much more than a biphasic material made of metal and gangue. Nevertheless, still today, most of the design and follow-up of industrial processes only consider metal grades; in best case metal grades per size class. Working with such heterogeneous material, made of several minerals, with different grain size and texture, implies to consider finer ore models to be able to predict the behavior of material and the subsequent operation performances. Analytical tools are now available to provide such relevant and accurate data which are applied in geometallurgy. To achieve a true geometallurgical approach for process design, but also for optimization, accurate process simulation absolutely needs to combine these fine ore models with the corresponding unit operation models. Based on generic, but configurable, criteria of description and a hierarchical approach, the material modelling tool embedded in the USIM PAC process simulator provides the necessary flexibility to describe any type of ore with almost no limit in the level of details and associated physical properties that can be met in the mining industry. The simulation engine and the mathematical models of unit operations are designed to manage such a diversity of material modelling. The paper present various ore models and some of their applications in flowsheet simulation. It proves the interest of this detailed and advanced approach to accurately predict plant performances. INTRODUCTION Process simulation used for process design and optimization requires accurate unit operation modeling. During the last decades, main efforts of the different teams developing software simulators have been dedicated to those models for comminution and physical separations such as flotation or gravity concentration (Kojovic et al., 2012, Soni and Yoon, 2014). In most case, the ore is modelled with a size distribution and metal or mineral grades. A water phase is considered for wet processes. The association of unit operations and such material models generally allows, after some parameters adjustment, to achieve good agreement between experimental and simulated data. The use of the obtained simulator to test alternative configuration of circuits or settings of equipment devices can be accurate enough as far as the ore processed is close to the ore used for simulator adjustment. However, when the ore gets very different grades, coming from different mineral proportions or grain sizes, the simulation results can be of poor accuracy. Because any physical separations such as gravity, magnetic or flotation, do not sort metal and gangue but particles, made of minerals, containing metals, the only way to be accurate enough for allowing extrapolations, is to model particles. The phase model tools embedded in the USIM PAC process simulator gives this possibility. However, as such level of details are not always available for analyses, the tool has been designed with a high level of flexibility, to adapt the model to the available data. In this paper, the principles of this conceptual modelling tool are presented, and some application cases illustrate the material model flexibility and its interest."

Jan 1, 2018

By Shankar Narasimhan

Data Reconciliation is a technique that has been developed more than fifty years ago and has been refined over the past few decades and is now being deployed routinely in several process industries such as refining, petrochemicals, mineral processing and hydro-metallurgy. The main aim of data reconciliation is to obtain accurate estimates of all process variables and parameters using plant measurements that are invariably corrupted by random errors and may also contain biases and other gross errors. These reconciled estimates are subsequently used either in offline exercises such as simulation, retrofitting, and de-bottlenecking or in online optimization and control applications. In this presentation, a brief history of data reconciliation, its assumptions, and uses are reviewed. New developments in this area are highlighted such as robust data reconciliation to handle model uncertainties, and reconciliation using identified models derived purely from operating data. Several case studies in the mineral processing industries reported in the literature are also highlighted to indicate the potential benefits of this technique.

Sep 1, 2012

By Ventsislav Stoilov, Angel Angelov, Ivan Korolev, Antti Remes, Todor Pukov, Stoyan Gaydardzhiev

"The process modeling and simulation studies aim for better process operation and favorable environmental impact. The objective of this work is to set up a simulation model of the Assarel copper flotation circuit for planning and optimization of the process with the aim of increased metal recovery. The model is parameterized based on laboratory flotation tests of ore samples from the Assarel mine together with mineralogical studies by X-ray diffraction (XRD) and Mineral Liberation Analysis (MLA). A combination of several techniques allows reliable identification of mineralogical composition of the selected samples. The changes in the proportion between sulfide minerals in various products along the circuit have been also documented. Further, this information is used in the simulation model and in data reconciliation procedures for establishing mass balance. The simulation accuracy is studied by comparing the simulations with the plant survey based mass balance. Constructed simulation model can be used to run alternative process scenarios. As an example, the flotation circuit performance with different reagent regimes is simulated based on data from batch flotation tests. Two reagents were considered for this purpose, dithiocarbamate-based one and xanthogen formate/mercaptan blend. It has been revealed that these reagents do not have the same influence on the selective flotation. The one improves flotation of gold and overall recovery but without notable selective action. In contrast, the other showed better selectivity towards copper, rejecting more gangue and lowering gold recovery. The model is also applicable for process optimization studies with different feed compositions, flowrates and circuit configurations. INTRODUCTION In the pursuit for elevated technological indicators, finding an optimal set of process parameters fully utilizing the potential of the equipment is a key factor to success. Over the past decades, there has been a substantial development in flotation circuit optimization through performance benchmarking using metallurgical modelling and steady-state computer simulation. Practical simulation engineering tools are based on experimental data obtained from laboratory tests and from plant surveys and operation data. With the help of flotation models, it is possible to study grade-recovery response with different feed compositions and throughputs, and with various cell-operating parameters, as well as to evaluate and design different flowsheet configurations. Outotec’s HSC Chemistry® software, which incorporates simulation and mass balancing modules alongside with several other tools, can assist with this optimization (Mattsson et al., 2015)."

Jan 1, 2018

By L. Y. Feng

The percentage of leached zinc in massive zinc oxide ores (11.7% Zn, >2 mm) is up to 73.3 % in 120 days. Alternatively, crushed to -0.154 mm, mixed with cement 5% (wt), pelletized and solidified. The diameters of the pellets obtained are between 5 mm to 8 mm. When the pellet solidification periods are 3 days, 10 days and 45 days respectively, the alkaline leaching rates of zinc in the pellets are up to 95.8%, 89.1% and 76.4% respectively. Decreasing the solidification time can re-duce reaction time, increase dissolution of zinc in pellets and lower the effect of initial zinc concentration on leaching rate. The experiment results show that the minimum solidification time is three days, and the kinetic study indicates that alkaline leaching of the low grade zinc oxide pellets is controlled by the diffusion of the leach liquor through the gangue layer in the whole leach process, and the apparent rate constants are 3.33×10-2 day-1, 9.18×10-3 day-1, 5.82×10-3 day-1 respectively.

Jan 1, 2014

By Lev Malytin, Alexander Dyachenko, Roman Kraydenko

"A new method of processing beryllium raw materials based on interaction of beryllium mineral with ammonium fluoride was developed. Possibility of extracting ammonium tetrafluoroberyllate from bertrandite and phenakite minerals is proved. The developed technology allows replacing the outdated sulfate technology developed by Materion Corporation. INTRODUCTION Beryllium is a strategic metal, which is produced only in three countries: the USA, China and Kazakhstan. The world production of beryllium amounts to only 300 tons per year. More than 75% of the world beryllium production is manufactured from oxide beryllium ores in the USA by Materion Corporation. In Russia there is a large-scale beryllium ore deposit «Ermakovskoye» containig over 1 million ton of ore. Concentration of beryllium in ore amounts to 1,2%. Since the mineral base of ore is calcium fluoride, the sulfate technology of Materion Corporation cannot be used. After alloying and sulphuric acid leaching the fluorine-ion found in fluorite and beryllium together evolve into the aqueous phase. At the stage of Be(OH) precipitation fluoride-ion prevents complete isolation of beryllium into the phase of hydroxide, which increases the loss of beryllium. A new ammonium-fluoride method for processing high-fluoride beryllium ores has been developed at Tomsk Polytechnic University. The ammonium-fluoride method involves fusion of ammonium fluoride (NH4HF2) with a beryllium concentrate. The concentrate containing 10% of beryllium oxide and 25% of calcium fluoride is produced by flotation from the ore of Ermakovskoe deposit. The main component of the ore concentrate is fluorite CaF2, which does not react with NH4HF2. Beryllium in this concentrate is represented by two minerals: phenakite Be2(SiO4)2 and bertrandite Be4[Si2O7](OH)2. Ammonium fluoride, unlike sulfuric acid, does not react with calcium fluoride and is not consumed for decomposing the mineral base of the concentrate. Ammonium fluoride is an accessible and environmentally safe reagent, worth up to $ 1,200 per ton. The total amount of ammonium fluoride is recycled and is used for decomposing a new batch of beryllium concentrate."

Jan 1, 2018

By Lev Malytin, Alexander Dyachenko, Roman Kraydenko

"A new method of processing beryllium raw materials based on interaction of beryllium mineral with ammonium fluoride was developed. Possibility of extracting ammonium tetrafluoroberyllate from bertrandite and phenakite minerals is proved. The developed technology allows replacing the outdated sulfate technology developed by Materion Corporation. INTRODUCTION Beryllium is a strategic metal, which is produced only in three countries: the USA, China and Kazakhstan. The world production of beryllium amounts to only 300 tons per year. More than 75% of the world beryllium production is manufactured from oxide beryllium ores in the USA by Materion Corporation. In Russia there is a large-scale beryllium ore deposit «Ermakovskoye» containig over 1 million ton of ore. Concentration of beryllium in ore amounts to 1,2%. Since the mineral base of ore is calcium fluoride, the sulfate technology of Materion Corporation cannot be used. After alloying and sulphuric acid leaching the fluorine-ion found in fluorite and beryllium together evolve into the aqueous phase. At the stage of Be(OH) precipitation fluoride-ion prevents complete isolation of beryllium into the phase of hydroxide, which increases the loss of beryllium. A new ammonium-fluoride method for processing high-fluoride beryllium ores has been developed at Tomsk Polytechnic University. The ammonium-fluoride method involves fusion of ammonium fluoride (NH4HF2) with a beryllium concentrate. The concentrate containing 10% of beryllium oxide and 25% of calcium fluoride is produced by flotation from the ore of Ermakovskoe deposit. The main component of the ore concentrate is fluorite CaF2, which does not react with NH4HF2. Beryllium in this concentrate is represented by two minerals: phenakite Be2(SiO4)2 and bertrandite Be4[Si2O7](OH)2. Ammonium fluoride, unlike sulfuric acid, does not react with calcium fluoride and is not consumed for decomposing the mineral base of the concentrate. Ammonium fluoride is an accessible and environmentally safe reagent, worth up to $ 1,200 per ton. The total amount of ammonium fluoride is recycled and is used for decomposing a new batch of beryllium concentrate."

Jan 1, 2018

By Kari Niiranen, Patrick Krolop, Viktoria Töyrä

Luossavaara-Kiirunavaara AB (LKAB) operates an iron ore mine, three beneficiation plants and three pelletizing plants in Kiruna, Sweden. The current methods of separation in the beneficiation plants are wet low intensity magnetic separation (WLIMS) and reverse apatite flotation, where the WLIMS stage is regarded as the crucial part of silica separation from the ore. Currently, the high-grade iron ore deposit of Kiirunavaara has a low grade of silica in situ, of approximately 3.0% SiO2, but the silica grade is expected to increase in the deeper parts of the deposit with a complex silicate mineralogy. An increasing silica grade will affect especially the production of direct reduction pellets (LKAB DR Oxide Pellets), one of the main products of LKAB in Kiruna with an average silica grade of under 0.75% SiO2. The silica grade in the magnetite concentrate is normally limited to a maximum level of 0.55% SiO2 in order to achieve the required chemical composition in the pelletizing process. In order to study the possibility to reduce silicate content in the magnetite concentrate (pellet feed) in the beneficiation plants in Kiruna obtained by wet low intensity magnetic separation (WLIMS), enlarged mineral processing test work was carried out in the LKAB’s mineral processing laboratories in 2016. The test work consisted of comminution, low intensity magnetic separation and reverse cationic silicate flotation at laboratory scale, complemented by screen analysis, chemical assays and mineralogical investigation. The results of these mineral processing test work showed that the magnetic separation is not sufficient to reduce the silica content in the magnetite concentrate, particularly in the case of the high-SiO2 ore type B2-a. The reverse cationic flotation can be regarded as a possible solution to reduce the silicate content in the magnetite concentrate obtained by WLIMS and to achieve the required silica grade of less than 0.55 % SiO2 in the pellet feed. The results of this study provide a good basis to understand the behaviour of different silicates in the existing beneficiation process. Furthermore, it serves valuable information for development of new beneficiation processes and for investment in new equipment at the beneficiation plants in Kiruna in the future.

Jan 1, 2018

By V. V. Trusova, Shiyrav Gandandorzh, K. V. Fedotov, V. A. Domracheva

"In Russia, 1.1 million tons of mercury-containing waste are stored. The amount of mercury in soils and heaps of industrial enterprises is estimated at 3000 tons; in dumps, concentration tailings, sludge dumps of the gold mining industry up to 6000 tons have been accumulated. Mining is one of the main sources of mercury emissions in the natural environment that poses a real threat to the natural environment. At present, the process of amalgamation is prohibited and not applied in gold extraction. The application of amalgam technology in gold mining led to the accumulation of highly toxic products in the form of gold-bearing amalgam and metallic mercury, with a mercury content of dozens and even hundreds of times greater than the maximum permissible concentration for soil (1). Therefore, mercury-containing tailings constitute an ecological hazard to the environment. According to published data, at waste storage sites of industrial enterprises, up to 50 tons of mercury enter soil annually, and up to 3.5 tons pollute the air. At the same time, mercury-containing waste due to the reduction of the country's mineral resources is an additional source of mercury. Therefore, it is necessary to recycle mercury-containing waste. INTRODUCTION Paperobjective: demercurization of technogenic raw materials (amalgamation tails, discharge outlets of tailing dumps) to be recycled. Under natural conditions, the mercury usually migrates in the form of three most common compounds: elementary mercury (Hg); mercuric mercury ion (Hg+2); methylmercuryion (Hg (CH3) +. Chemical compounds Hg (II) are found in natural conditions much more often than Hg (I).The authors propose to conduct demercurization of technogenic raw materials (mercurial tailings) in two stages. At the first stage, the mercury in the form of small balls, the finest particles (drip mercury) can be extracted by applying the gravitational method. We propose a method for centrifugal processing of crushed waste in the apparatus of Itomak company, Novosibirsk, as the basis of the technological process for mercury recycling. The extraction of metallic mercury out of mercury-containing waste into commercial products, according to the technological characteristics of the Itomak apparatus, reaches 86.0% (2, 3). At the second stage, the extraction of mercury out of technogenic raw materials is performed applying a sorption method and using carbon sorbents based on fossil coals. Hard coal of the Kuznetsk Basin has been already used for industrial production of active carbons (carbon sorbents) of the KAD-iodine type, which is one of the prerequisites that determine the choice to use it for the aims of this study."

Jan 1, 2018

By V. A. Chanturiya

The application of High-Power Nanosecond Electromagnetic Pulses (HPEMP) irradiation in dressing of resistant gold- containing ores appears attractive as this technique provides for a significant increase in precious metal recovery (30-80% for gold and 20-50% for silver), therewith helping reduce both energy consumption and the cost of products. The prospects for applying pulsed power technologies in the mineral processing in order to increase the contrast between the flotation properties of raw minerals dictate the need to perform special studies of the effect of HPEMP on the structural state, phase composition, and physical and chemical properties of surfaces of sulfide minerals as the main carriers of precious metals. This study deals with plausible mechanisms of the micro- and nanophases formation on the surfaces of sulfide minerals under the effect of high-power nanosecond electromagnetic pulses (HPEMP) with high electric field strength ~107 V/m. This paper discusses reasons for the changing of flotation behavior of natural pyrite and arsenopyrite during the pulsed-power treatment. It was concluded that HPEMP treatment may be regarded as a tool to change the surface composition and sorption ability of collector (xanthate) onto mineral surface, and, therefore, it allows to control the hydrophobic-hydrophilic surface balance. Single mineral flotation tests have been shown the significant increasing of the pyrite recovery while the arsenopyrite was depressed. The results of gold containing pyrite (pyrrhotite)-arsenopyrite ore flotation in batch approve the decreasing of arsenic recovery in sulfide concentrate. Keywords: pyrite, arsenopyrite, high power nanosecond pulses, activation, selective flotation

Sep 1, 2012

By Eduard Stefanescu, James M. Mwase, Egon Sehner, Stoyan Gaydardzhiev

"The study is focused on investigating the production of synthetic rutile of 95% purity or higher from a Norwegian ilmenite ore concentrate. This was to be achieved using a sequential process involving oxidative-reductive roasting at 975°C, followed by magnetic separation to further concentrate the sample and finally leaching with 20% HCl to remove mostly Fe-based impurities leaving behind synthetic rutile. The key process parameters considered were roasting time (1 and 2 hrs), leaching temperature (100, 125 and 145°C), pulp density (15 and 25wt%) and leaching time (3 and 6 hrs). The option of directly leaching the concentrate without pre-treatment was also evaluated. Test work was conducted using samples screened to a size fraction of 105-212 µm. Samples were leached in a 1 L glass autoclave. The reactor was hermetically sealed and thus the resulting pressure was not controlled but was dependent on the process temperature. The product from leaching was calcined and analysed via XRD and SEM to determine the purity of the product. The magnetic separation step was found to be ineffective and discontinued early in the study. Thus far a product with 97% synthetic rutile purity as analysed by SEM has been produced by directly leaching the concentrate at a temperature of 145°C, pulp density of 25wt% and leaching time of 6 hrs.INTRODUCTION The study investigates the use of a process route comprising oxidative-reductive roasting, followed by magnetic separation and then hydrochloric acid leaching and calcination to produce synthetic rutile from ilmenite concentrate. The grade of synthetic rutile produced should be suitable for chlorination to produce pigment grade TiO2. Hence the proposed process should produce a product consisting of no less than 95% synthetic rutile. Additionally, the product should have no more than 1.5wt% combined content of CaO and MgO (Zhang et al., 2011) as they tend to form high-boiling-point chloride leading to the malfunctioning of fluidized bed chlorinator (Cardarelli, 2008). Finally the product should have no more than 2.1wt% of Fe2O3."

Jan 1, 2018

By Sami Makni, Ahmed Bouajila

"To remain competitive in an open market, mining companies are more and more requested to optimise their complex and highly multivariate processes. The increasing amount of process data, brought about by intensive use of instrumentation and the progress in computer data acquisition, enhances the need for more and more powerful and robust tools for the analysis and modelling of the process data. Now, with the emergence of multivariate data analysis (MDA) techniques, on-line clustering of mineral process behaviour is possible and could help in the extraction of valuable process operation indicators. MDA can be used in different mineral processes such as flotation, grinding, smelting etc.In this paper, we focus on grinding process considered as the most costly and power consuming node of the mineral processing scheme. Different objective-oriented levels of data clustering will be addressed.First, the evaluation of grinding operation taking into account data and constraints from all processes before (e.g. blasting, crushing) and after (e.g. flotation, leaching, etc.) the grinding node. The second level deals with process diagnosis by differentiating operation level and identifying behaviour drifts. Control and optimisation are two other levels concerned here with the selection of optimum trajectories in the drift correct.Industrial plant data are processed in order to illustrate the feasibility and the potentiality of MDA applications in addressing the clustering and interpretation of grinding mill behaviour."

Jan 1, 2003

By I. Gombköto

To sustain the growth of industrial production necessitates the increasing of the amount of processed mined raw and secondary materials. As a side effect of raw material production, the amount of tailings material, - coming from mineral processing operations ? increases as well. Using wet technology, these tailing materials are generally transported and stored with the water used for previous operations. This habit could lead to dangerous situations. Enough to remember the disasters had happened to the last decade, safe tailings operations need to remove as much water from the solid as possible. Environmental, technical and economical needs should meet, therefore using a technology combining these specific needs could be necessary. Paste technology was introduced generally into tailings operations at the last ten years, as an answer to the call of the above mentioned challenges, however there are still many questions to answer within the frame of using it. The Institute of Raw Materials Processing and Environmental Process Engineering of the University of Miskolc has been dealing with mechanical process engineering, especially by the support of the EU project TAILSAFE Contract nº EVG1-CT-2002-0066 and the NKFP 3A/068/2004 Hungarian fund extensive research has been carried out about paste technology. Different test methods and materials were applied. Tested tailings materials were as follows: different fly ashes from coal power plants from Pécs ? Hungary, Gyöngyös ? Hungary, Neyvil ? India, different Croatian ashes, hazardous waste incineration plant ashes from Dorog ? Hungary, different kind of sands from Hungary, fine glass sand from Fehérvárcsúrgó ? Hungary, flotation ore processing tailing from Gyöngyösoroszi ? Hungary. Different additives, as gypsum, bentonite and cement were tested as well. The applied test methods were as follows: physical material testing (laser sizer, piknometer, different microscopes), X-ray diffraction analysis, settling columns, the three measuring pipes tube viscometer (16~27 mm pipe diameter), a hydraulic test loop with different pumps (40~53 mm pipe diameter, 0.2~4 m/s average slurry velocity range, 0%~52 % volumetric concentration range), rotational viscometer, Cassagrande yield locus tester and the slump cone test. This paper summarizes and systematizes the results of the aforementioned research as well as reports about the developed paste pipe transport design method based on the fine suspension ? coarse mixture model. The paper also reports about a systematic experimental series with different materials to establish criteria when high concentration suspensions behave like paste.

Jan 1, 2014

By Alireza Javadi Nooshabadi, Kota Hanumantha Rao

Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was examined. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn, Fe) S), and galena (PbS) generated H2O2 in pulp liquid during wet grinding and also when the freshly ground solids are placed in water immediately after dry grinding. Pyrite produced more H2O2 than other minerals and the order of H2O2 production by the minerals was found to be pyrite > chalcopyrite > sphalerite > galena. The pH of the water influenced the extent of hydrogen peroxide formation with greater amounts of H2O2 produced at highly acidic pH. Furthermore, the effect of mixed sulphide minerals, i.e., pyrite-chalcopyrite, pyrite-galena, chalcopyrite-galena and sphalerite-pyrite, sphalerite-chalcopyrite and sphalerite-galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction. There is clear correlation of the amount of H2O2 production with the rest potential of the sulphide minerals; the greater the rest potential of a mineral the greater the formation of H2O2. This study highlights the necessity of revisiting the electrochemical and/or galvanic interactions between sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behavior in the context of inevitable H2O2 existence in the pulp liquid.

Jan 1, 2014

By S. N. Shahcheraghi

Grinding is one of the most important and expensive steps in mineral processing allocating about 70% of the total energy in the plants. Nowadays, for optimizing the energy consumption in these equipments, the rock behavior and its size distribution during the crushing have been normally predicted using selection and breakage functions to be able to optimally model and design grinding circuits. For this purpose, the case of Miduk copper mine in Kerman province, Iran was studied. The samples were taken from primary ball mills feed and were subject to screen analysis to achieve uniform sizes. Then, selection and breakage functions were determined at the different grinding times in laboratorial ball mill. Finally, ore size distribution in different grinding times was estimated using MATLAB software

Sep 1, 2012

By A. J. Ren

The ore, dealt with in this paper, is a certain rutile in Shanxi province, China. A trunk flowsheet with gravity concentration is used, and the TiO2 grade of rutile concentrate is about 90%, but the TiO2 recovery is less than 50%. In order to increase the rutile recovery, the laboratory test, using the truck flowsheet with flotation, is carried out. The flowsheet is that firstly floatable gangue is re-moved by flotation; secondly rutile is recovered by flotation; and thirdly impurity is separated by LIMS, HIMS, gravity concentration and flotation from the rutile concentrate. In rutile flotation, BK423, a phosphonic acid, is used as rutile collector, and Na2SiF6 is used as silicates depressant. The experimental result shows that the recovery of rutile concentrate is improved obviously, and the recovery of ultimate rutile concentrate is 68.01% TiO2 and 84.88% rutile. The TiO2 grade of first concentrate is 95.98%, and the TiO2 recovery is 45.60%. The TiO2 grade of second concentrate is 80.53%, and the recovery is 22.41%. Furthermore, magnetite and ilmenite are recovered in the flowsheet.

Jan 1, 2014

Three high-grade limestone ores (CaO content over than 53 %) with different geological origin have been characterized for the synthesis of aragonite precipitated calcium carbonate. Limestone ores examined in this study showed different crystal particle size that cause difference in disintegration degree and surface area of the synthesized slaked limes. The limestone ore with smaller crystal particle size showed higher disintegration degree and also yielded slaked lime with smaller surface area, which is good for the synthesis of aragonite precipitated calcium carbonate. The synthesis of aragonite PCC using the limestone with a small crystal particle size was yielded needle-shaped aragonite with the aspect ratio of about 14. Keywords: limestone, lime, slaked lime, precipitated calcium carbonate, aragonite, accelerated carbonation

Sep 1, 2012

By M. Kargar Razi

According to the unique properties and many applications of nano gold powder, it was prepared from an acid leaching tail solution. The low value residual gold ions were existed in acid leaching tail solution in Mouteh gold mine in Iran with fellow heavy metals as pollutants (FeP2+P, Cu P2+P, and ZnP2+P). Preparation of nano gold particles is achieved by two steps. First, by addition of ammonia as separation step and second using suitable reagents for precipitation/extraction step. Conditions control includes pH, residence time, temperature, agitation, and molar ratio of reducing agents to reagents. The optimum results show that nano gold powder was prepared by temperature of 250 °C, urea to gold molar ratio of 3:1, and reduction time of 20 min. Also, heat treatment improves dispersion of nano sized gold particles. The properties of gold nano powders were evaluated by SEM, A.A, and UV- Vis spectroscopy methods. The size of prepared particles were about 80nm to 90nm.

Jan 1, 2014

By J. Sachweh

The last few years have shown a significant increase in the demands for fine products, which means everything under 10µm. There are more and more applications requesting extremely fine ground powders leading to the development of several milling machines which are better adapted to the new products. Cost saving machinery is important for the worldwide competition. Especially the energy consumption is extremely important for any fine ground product. A worthwhile field for mechanical engineering companies is the dry milling machinery as there are not as many solutions available compared with wet milling applications. In many cases wet mil-ling is the only solution to reach the requested quality, but there are also a lot of applications which can be performed in a dry way on the premise that the right mill and classifier are available. Dry milling permits to save energy and allows to influence the particle size distribution more easily, as a classifier can be used in that case. This will significantly reduce production costs and can enable the production of new products with better properties. The Tower Mill is well-known already in the field of wet grinding. Recent research work and end user reports have shown clearly that this mill has a lot of advantages compared to a ball mill. Up to now this advantage is limited to wet grinding as only few mills have been used for dry grinding. There is no doubt about the efficiency of dry grinding, but new and improved feeding and discharg-ing systems have to be developed in order to enable the full range of its advantages. Especially the high production capacities are a challenge for mill and classifier designers. The Tower Mill with its special grinding mechanism allows to cope with very high feed rates, because machines with more than 1 MW are available. Improved grinding efficiency will decrease the circulation load through the classifier and also lead to smaller classifier sizes. Together with im-proved mills the classifier can work more efficiently because the feed material is finer. Results of a test center plant production will be shown together with a comparison between stan-dard machinery and the MaxxMill®. Preferential applications for each mill will be explained.

Jan 1, 2014