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By Fernando Concha

"Thickening development and innovation in the 20th Century occurred in two decades periods. From 1900 to 1920, Dorr invented the continuous thickener and the first method of thickener design was proposed by Coe & Clevenger. From 1920 to 1940, the operating variables of continuous thickening were discovered by Comings at the University of Illinois. From 1940 to 1960, the first theory of sedimentation was proposed by Kynch and several authors, Talmage and Fitch, Ysioka and Hasset used this theory to develop new thickener design methods. From 1960 to 1980 Shannon and Tory demonstrated that this theory applied exactly for the sedimentation for small rigid spheres in water and Scott showed that it is not valid for flocculated suspensions. From 1970 to 1990 the phenomenological theory of thickening was developed based on the theory of mixtures. Papers by Kos, Bascur, Thacker and Lavelle, Concha and Bascur, Concha and Barrientos, Auzeris, Buscal and White, Landman, White and Buscal show the application of this theory to thickening and Adorjan presented the first thickener design method based on this theory. In 1980s the mathematical analysis and solutions of Kynch’s model for batch and continuous thickeners was obtained by Bustos. The new Century brought solution to the phenomenological thickening model which was applied to thickener design, simulation and control. The method of thickener parameter determination was improved by designing new laboratory and on-line instruments and advanced control strategies was designed and validated in plants. Structural thickener design produced improved feedwells and feed dilution mechanism, stronger raking mechanism and different thickener shapes according to new ambient requirements, such as high rate thickener and paste thickeners that increase the water recovery. New challenges include better flocculation and advanced control strategies."

Jan 1, 2014

By Marco Le Roux, Quentin Campbell

"Most coal supply contracts include strict limitations on the moisture content of the products. This limitation exists because of calorific, transport cost, and handling reasons. Managing the retention time on product stockpiles to facilitate drainage commonly controls the moisture content in coarser coal products. A large-scale pilot rig has been commissioned to facilitate research in this regard. A stockpile of about 15 tonnes of a typical washed power station feed coal from the South African Witbank coalfield was loaded onto a 6 x 6 m platform on load cells. The platform was shielded from the elements, like rain and wind, and an artificial rain sprinkler system was installed. A variety of rainfall intensities and durations were simulated and the water accumulation and run-off was measured. The material was also sampled for moisture and particle size before and after the series of tests. At the same time, the drainage characteristics of the same coal were determined in smaller scale drainage columns for comparison. Data is presented in this paper on the drainage rates, maximum accumulated moisture, final moisture, and the effect of evaporation. The results showed that for all rainfall patterns tested, drainage ceased after 2 - 3 days, but evaporation continued for at least the time period under investigation, albeit at a low rate. Evaporation contributed to dewatering at rates that depended on wind, temperature and humidity conditions. This caused a slow reduction in moisture content even after the run-off had ceased. The value of the final moisture depended on the coal PSD. It was also found that the intensity and duration of the more intense rainfall events did not significantly affect the maximum water accumulation, with drainage rates equaling the corresponding water addition rates which suggested that water ran off over the surface of the stockpile with limited penetration. During more intense rainfall scenarios, some penetration was apparent."

Jan 1, 2014

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 as shown in figure 1."

Jan 1, 2003

By V. Avzian Belozersky reg 453524, Vasily V. Lodeischikov, Adolf F. Panchenko, Til E. Vidusov, Olga D. Khmelnitskaya, Vladimir E. Dementyev

"Presently, there is a number of gold deposits which cannot be developed using the conventional technologies with acceptable economic indices due to low gold content, high depth of ore occurence and other conditions. For such ores, an innovation technology of underground gold leaching in the place of their occurrence is of special interest. The technology excludes ore mining and hoisting to the surface. This enables the production cost of 1 g of gold to be considerably reduced in comparison with conventional methods of gold recovery.The Maminskoye deposit oxidized ores were chosen as the subject of underground leaching because they are ideally suited for this purpose due to their composition and hydrological parameters. The ore reserves within the underground leaching block are 90, 000 t with 0.7 g/t average gold grade.An experimental system for underground leaching with a throughput of 15 cubic m solutions per hour was developed. A combined filtration-and-infiltration system of the ore massif development by holedrillingwhere pregnant solutions are air-lifted was worked out.The underground leaching flowsheet includes Au dissolution by chlorine-containing solutions, Au adsorption from solution onto active carbon, carbon concentrate ashing and calcine melting to produce Dore metal."

Jan 1, 2003

By Jyotirmoy Sarma

The surface-active agent, like sodium humate, in the present systems produces a very stable suspension. Saline water up to 15% is used as a flocculating agent for recovery of water from the dispersed phase containing 1% solid. The recovered water in the presence of humate at different concentrations (0.05-0.2%) and at pH 8 also produces a stable dispersed phase of iron ore fines and tailings. The beneficiation techniques adopted in the present systems can increase ~ 2-3% iron value in the concentrate of iron ore fine, however, for tailings the increase in iron value is not observed. Keywords: beneficiation, flocculation, iron ore fines, tailings, saline water, surface-active agent

Sep 1, 2012

By Jacques Eksteen, Sharif Jahanshahi, Neill J. Bartie, Shouyi Sun

The gradual increase in Cr2O3 levels of UG-2 ores has become a matter of concern for platinum producers in South Africa due to the formation of chromite spinels in melters. The effect of bulk FeOx/MgO ratio, Cr2O3, temperature and oxygen partial pressure on phase stability and composition in the system FeOx-MgO-SiO2-Cr2O3-Al2O3-CaO has been studied between 1400°C and 1600°C using the drop-quench technique. An increase in temperature was found to result in a decrease in [Fe/Cr]spinel. This effect is shown in Figure 1 for the six slag compositions investigated.

Jan 1, 2003

By A. N. Lazaridi, S. I. Afanasenko, S. A. Safonov

"This scientific work presents the results of comparative tests of centrifugal concentrators with a vertical and horizontal axis of rotor spinning. The article also describes a qualitative analysis of the behavior of mineral particles in a centrifugal field with fluidization for concentrators with a horizontal axis. The analysis was aimed at studying the effect of the quantity of fluidization holes in the basket on the degree of magnetite recovery and the effect of the basket rotation speed. The experiment demonstrated that concentrators with a horizontal axis are not inferior in terms of the qualitative recovery indicators to concentrators with a vertical axis. The researches also provided facts about the wide practical application of concentrators with a horizontal rotor arrangement. INTRODUCTION Methods of gravity concentration, attractive for their ecological purity, see their revival. First of all, this is due to the invention of centrifugal concentrators with the fluidization of the layer of mineral particles. The creation of a new generation of gravitational centrifugal concentrators made it possible to achieve a tangible breakthrough toward the direction of reducing the grain size of mineral particles effectively divided by density, up to 10 µm at a high production capacity of the vessels reaching 500-600 tons per hour. Such separators enable dressing a crushed ore of flotation grade and in some cases replace the flotation process, especially in primary operations, in joint cycles. In the long term, this method of concentration should significantly expand the scope of application, replace partial flotation processing, ensure the introduction of environmentally friendly energy-saving technologies at ore-processing plants. Summarizing the international experience for the last 30 years, it can be confidently asserted that centrifugal methods of concentration have become widely used in processing schemes for the separation of finely dispersed products in the mining industry. An interest in the creation of installations using centrifugal acceleration for the separation of mineral particles by density has particularly grown. Dozens of centrifugal concentrator designs have been created and continue to be improved. Centrifugal concentrators have a number of obvious advantages, which are especially relevant today. Losses of fine gold during the process of mining are well known. Gold washers screw, spiral separators, jigging machines and sluices traditionally used in the production processes do not solve the problem of fine gold recovery. Table 1 shows the results of comparative tests of a centrifugal concentrator (ITOMAK) and other concentration equipment. According to the research conducted by the company GMRS, Krugersdorp, South Africa (2000), the initial material was gold ore crushed to 150 micrometers (90%)."

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 H. Matsuoka, M. Kawata, K. Haga, K. Mitsuhashi, C. Tokoro, A. Shibayama, R. Kawarabuki

"The mineral surface characteristic is important for the flotation because it affects the flotation behavior of the mineral. Conventionally, in the flotation process for Cu sulphide ores, the degree of ore floatabilities varies depending on the mineral phase such as chalcopyrite, bornite, or chalcocite. Therefore, mineral surfaces are treated with activators and depressants to control the floatability of target minerals and obtain high recovery performance. There are lots of flotation reagents which have many effects on ores depending on ore characteristics and conditioning methods. For example, sodium hydrosulphide (NaHS) as activator is known to have the effect of resulfurizing and activating oxidized surface of chalcocite and bornite. On the other hand, NaHS is reported to depress chalcopyrite floatability. In this study, the evaluation of mineral surfaces was conducted to predict flotation behavior. Activation of Cu minerals (chalcocite, bornite, and chalcopyrite) by NaHS was experimented. The three Cu minerals were treated in two steps. Firstly, they were oxidized in the drying oven. Secondly, they were treated with NaHS solution. After each treatment step, each mineral was measured by contact angle apparatus then analyzed by X-ray photoelectron spectroscopy (XPS). The results of surface characteristics after oxidation and NaHS treatment were utilized to accomplish the study’s objective. INTRODUCTIONCopper sulfide minerals such as chalcocite (Cu2S), bornite (Cu5FeS4), chalcopyrite (CuFeS2) are well-known abundant copper-containing minerals. Generally, these minerals are recovered and concentrated via collector flotation. In flotation, the wettability of the mineral surface is an important factor in its recovery. When the surface is hydrophobic, the mineral has high floatability and when the surface is hydrophilic, it has low floatability. Usually, Cu sulfide minerals have high floatability. However when their surfaces are oxidized, their floatability becomes lower in flotation. In this case of surface oxidation occurrence, activators are added before flotation to modify the surface property of the Cu sulfide minerals so that they can be floated at a certain flotation condition. Sodium hydrosulfide, NaHS, is known as one of the activators. Zhao et al. (2018) reported that NaHS is often used as sulfurizer for oxidized sulfide minerals (Becker et al., 2014, Clark et al., 2000, Newell et al., 2006, 2007, Newell and Bradshaw, 2007). Although the floatability of Cu sulfide minerals was raised by adding NaHS generally, the results of this flotation study showed that the flotation behaviors of mixed chalcocite, bornite, and chalcopyrite were different. Therefore, it was considered that NaHS works differently for each mineral."

Jan 1, 2018

By E. D. Gálvez, L. A. Cisternas, R. Acosta-Flores, F. A. Lucay

The optimal design of flotation circuits is a complex task as the mathematical model is nonconvex and includes binary variables, making it difficult to obtain the global optimum solution. Furthermore, there are epistemic uncertainties in the evaluation of the performance of flotation and grinding units. This epistemic uncertainty is generated because semi-empirical models are used with parameters that are a function of the circuit design. Therefore, it is not possible to adjust these models based on experimental data as the experimental conditions are not known at the design stage. This paper analyzed the effect of the uncertainties in flotation units, grinding units, and both units simultaneously. As a result, this paper demonstrated that, for a specific mineral, few structures exist that are optimal for a wide range of values in flotation stage recoveries and liberation fractions in the grinding units. A remarkable aspect of this work is that it introduces grinding, one of the operations that most influences energy consumption in mining and has not been studied in depth in the design of flotation circuits. To demonstrate that there are few optimal structures for a specific mineral when there is high uncertainty in the process units, an a priori approach or first principles was used. Two methods of proof were used: proof by construction and proof by exhaustion. Proof by construction is the construction of a concrete example with a property to demonstrate that something having that property exists. In this work, two examples were considered in the proof by construction. In proof by exhaustion, the conclusion was established by dividing it into a finite number of cases and proving each one separately. Here, 30,000 cases were included for each example. This knowledge can be useful in circuit design including grinding, e.g., after selecting the circuit (or the most promising circuits), and the equipment design parameters and operating conditions can be defined based on simulation and laboratory tests.

Jan 1, 2018

By Y. G. Zhu, G. B. Zheng, Z. B. Qian, X. R. Zhang, A. J. Ren

"Based on PMA polymer, two kinds of novel macromolecular depressants, PMTA and PMA-PDTC, were designed and synthesized. PMTA was used for desilication in hematite reverse flotation. PMA-PDTC was used in the separation of Cu-Pb concentrate as galena depressant. The micro-flotation results showed that PMTA exhibited a good depression performance compared to starch. We found that 98.29% of hematite was depressed with a grade of 65.24% at pH 11 with 12 mg/L dosage. The flotation separation results of Cu-Pb concentrate demonstrated that PMA-PDTC was an effective depressant for galena in separating galena from chalcopyrite. When IPETC was used as collector, at pH 8.5, the flotation separation tests demonstrated that Cu concentrate was obtained with a grade of 28.89% containing 12.12% Pb.INTRODUCTIONThe role of depressants is crucial in the separation of oxide or sulfide minerals, for example, the desilication in hematite concentrate and the separation of Cu-Pb concentrate.It is well known that, in current reverse flotation processing of hematite, starch is a traditional depressant for hematite. Most research focus on the modification of starch [1-3]. Although its depressing ability is enhanced in some extent, its own shortcomings is still existed. So, it is important to develop a new hematite depressant instead of starch.In the separation of chalcopyrite and galena, because of their extremely similar floatability, it is difficult to realize the separation of these two minerals. Traditionally, there are two methods widely used in the separation of Cu-Pb concentrate by depressing chalcopyrite with cyanide [4] or depressing galena with dichromate [5] in most cases. However, these two inorganic depressants are toxic and hazardous. Their use could cause heavy pollution problems and a large loss of precious metals. In order to adapt to the requirements of sustainable economic development and environmental protection, the research and development of green and environmental depressant has gained more and more significance."

Jan 1, 2018

By Zhou Yucai, Zhou Jing, Zhu Yimim

"In this article, research on the beneficiation process for a carbon containing copper-molybdenum deposit in Henan province has been carried out. The elemental composition, mineral composition and relative content, characteristics of embedded and symbiotic relationship, the forms of the copper-molybdenum particlesand their disseminated features, and particle size distribution have been worked out. Flotation behavior of copper-molybdenum mixed flotation and separate flotation in the systems of different collectors and depressants, and the influence of different flowsheets on the copper-molybdenum separation have been investigated by actual ore flotation tests, The separation problem of copper-molybdenum for a carbon-containing copper molybdenum ore has been solved.INTRODUCTION In this paper, technological mineralogy, small experiment of flotation recovery of copper Molybdenum ore, and expansion experiment of mineral recovery of copper Molybdenum ore are introduced. (1) The results of process mineralogyshow that the content of copper is 0.42%, molybdenum is 0.021% in the raw ore, they mainly exist in the form of sulfide, with low oxidation rate. (2) The results of flotation tests show that the combination of butyl xanthate and isopropyl ethylthionocarbamate reagents is effective collector of copper-molybdenum mixed flotation for carbon copper molybdenum, (3) When the grinding fineness is -0.074 mm taking 71%, and regrinding fineness is -0.038 mm taking 85%, the flowsheet of first copper-molybdenum mixed flotation and then regrinding rough concentrate and last separation flotation of copper and molybdenum was adopted, which realized the effective separation of copper and molybdenum for carbon-containing copper molybdenum ore. Good flotation indexes were gained with copper concentrate containing 20.96% copper and copper recovery 90.06%, and molybdenum concentrate containing 46.78% molybdenum with molybdenum recovery 72.68%. (4) The seven days average concentration index of scale-up the continuous dressing experiment is as follows, the copper content in copper concentrate is 20.16% and the recovery is 89.61%; the molybdenum content in molybdenum concentrate is 24.58% and the recovery rate is 74.09%."

Jan 1, 2018

By S. M. Marr

"Water availability and water-related environmental impacts are now one of the most significant challenges facing primary production within the global minerals industry. Whereas the industry has, to a large degree, been able to demonstrate a collective response to greenhouse gas emissions and global energy demand, no such strategy exists for water. In part, this is because water quality and availability are regional issues, but also too because water quality cannot be defined by a single attribute, and extrapolation of any set of water quality indices to their associated environmental impact is no trivial matter. In design situations specifically, the situation is exacerbated by the fact that process engineers typically focus on site water management, and consent authorities are primarily in catchment level issues. As a result, an integrated approach to water management is difficult. This does not stem from an unwillingness by members of the professional team to engage with the total picture of water quality and availability, but principally from the fact that there is no common language through which to approach the problem.It is imperative for process engineers to extend their design system boundary, to reflect the geographic scales over which water impacts need to be quantified. It is not suggested that chemical engineers and metallurgists should become experts in geochemistry, hydrology and limnology. Rather, the challenge is to develop tools through which better communication of the relevant environmental impacts can be achieved.It has been estimated that 50% of the capital for new plants is devoted to handling wastes. The integration of mass exchange processes, adopting a life cycle analysis framework for considering environmental impacts and shifting the emphasis from effluent concentration to environmental impacts should be adopted in the framing of the design problem have been suggested as techniques of Environmentally Conscious Design. Shifting the focus from effluent concentrations and legislative requirements to evaluating the environmental impact is an important step towards improving the overall environmental performance of a process facility."

Jan 1, 2003

By Felix M. Zhuravlev, Vitalij P. Lyalyuk

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By Steven Wright, Sharif Jahanshahi, Shouyi Sun

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By Qingming Feng, Dongbo An, Jinhong Zhang

The application of the geopolymerization of mine tailings as construction materials is hindered by the fact that conventional geopolymerization of mine tailings usually involves a large consumption of alkali, i.e., sodium hydroxide (NaOH), and a long curing time, because of the relatively low reactivity and the high Si/Al ratio of mine tailings when compared to other active resources such as fly ash and clay minerals. The present study is focused on efficiently activating mine tailings, reducing alkali consumption, decreasing curing time and improving compressive strength. We firstly investigate the temperature effect on the alkali activation of mine tailings during geopolymerization. Secondly, the impact of additives, i.e., calcium hydroxide and aluminium oxide, on the compressive strength of samples is studied. Thirdly, different forming pressures have been applied to remove excessive alkali and its impact on samples' strength has been investigated. The results show that a 40 MPa unconfined compressive strength (UCS) can be achieved with the geopolymerization samples after mine tailings are activated by sodium hydroxide at 170°C for 1 hour with the addition of calcium hydroxide and alkali dissolved aluminium oxide, further compressed with a 10 MPa forming pressure and finally cured at 90°C for 3 days. The findings of the present work provide a novel method for the geopolymerization of mine tailings as construction materials.

Jan 1, 2014

By S. J. Ma

To look into the ionic release laws of the sulphide tailings in some copper mine, the effects of the soaking period, temperature, oscillating speed, and pH of the solution on the heavy metals ionic re-lease were investigated through static soaking leaching experiments. The results show that Cu2+, Zn2+, Pb2+, Cr3+ are of different leaching behaviors with soaking period. The ionic concentration in soaking solution for four ions has a great gap. Zn2+ is more easily released out of sulfide tailings than Cu2+, Pb2+, Cr3+ , and the ion release concentrations are not proportion to their contents in the tailings. The time for each ion to reach its maximum concentration are different. The minimum concentration do not occur at the first day. Each ion concentration shows different fluctuation characteristic during soaking process. Solution pH, oscillating speed, and temperature affect the releases of Cu2+, Zn2+, Pb2+, Cr3+ out of the tailings to different extent. The effect of pH value on ion re-lease for the four ions ranks as Zn2+>Cu2+>Pb2+>Cr3+, the oscillating speed ranks as Pb2+>Zn2+>Cu2+>Cr3+, and the temperature ranks as Cr3+>Pb2+>Zn2+>Cu2+.

Jan 1, 2014

By Ming’an Wei, Chuanyao Sun, Zhenguo Song, Zengwen Dou, Yangge Zhu

"The effects of pH value and dissolved oxygen(DO) in the pulp to the Eh and flotation recovery of pyrite were investigated using potential online detection system. The effects of DO on the Eh and flotation recovery were tested which DO changed from 0.5mg/L to 19.5mg/L in six different pH buffer flotation systems (pH: 4.00, 5.62, 6.86, 9.18, 10.01 and 10.96 respectively). Results of potential detected online demonstrated that the pyrite electrode has nearly the same response to potential as with platinum electrode for the FeS2-xanthate-H2O flotation system under the study conditions. Pulp potential and recovery of pyrite were remarkably influenced by pH, both of them decreased with the increasing of pH and, especially, sharp decreased after pH above 9. As the concentration of DO was increased, the Eh was also increased in all the six different buffer solution, while the flotation recovery was decreased. And the amplitude of variation for Eh was not obvious in acidic and neutral solutions, but on the contrary, the change for flotation recovery was remarkable in neutral and alkaline conditions. This may indicate that the flotation of pyrite will be deteriorated in alkaline and high DO concentration conditions."

Jan 1, 2014

By M. Schneider, J. Regler, Th. Neesse, AKW A. + V GmbH & Co., J. Dueck, F. Donhauser

"In practice the advantages of high-quality hydrocyclone separation are often reduced due to varying feed suspension properties, solids concentrations and particle size distributions. Therefore, a new control system for a battery of hydrocyclones with a nominal diameter of 150 mm was developed for bentonite slurry regeneration associated with tunnel excavation using bentonite slurry as a transport aid and lubricant.As can be seen from Fig. 1, the hydrocyclone battery has been modified for regulation purpose. The battery was manufactured with an overflow collector as a closed pressure chamber (1) with a discharge pipe (2) where the throttling valve (3) was installed.With increased stepwise throttling of the overflow, the volume split is changed with the effect that the solids discharge in the underflow is increased. To stabilize the total throughput, the feed pump speed is simultaneously increased so that the pressure drop ?p across the hydrocyclones (9) remains approximately constant. However, this causes the increasing pressure inside the hydrocyclones, which results in a further intensification of the underflow discharge. After the detection of the break through of the air core (transition state) by the capacitive sensor (4), the throttle valve is shortly opened, and a new regulation interval can start again. The control concept is focused on maintaining the optimum operating state, characterized by the underflow discharge shape at the transition state between rope and umbrella discharge."

Jan 1, 2003

By Hakan Gürsoy, Kemal Bilir, Bahri Öteyaka, Ali Uçar, Oktay Sahbaz

"Jet diffuser flotation column (JDFC) is a flotation device resembles Jameson flotation cell (JFC) in terms of operational principles, but it has significant difference structural points of view. This difference is an essential change in the downcomer to provide an increase of performance of the cell for the coarse particles. By using the hydrodynamic and fluid mechanics principles (Sahbaz et al., 2013) downcomer was changed as a diffuser, therefore turbulence at the end of the downcomer was reduced to provide better aggregate stability and homogeneous dissipation of the bubbles in the separation tank. According to experimental results carried out in pilot scale JDFC, not only higher flotation performance comparing with Jameson cell was obtained specifically for coarse talc particles in a certain condition, but also quiescent froth layer was acquired. In these tests, vertical pipe of JDFC having inlet diameter of 6 cm and outlet diameter of 11.5 cm was used with separation tank having the diameter of 39 cm. By using the data, pilot scale JDFC having 410 cm vertical pipe integrated with separation tank was produced and flotation tests were carried out by using the naturally hydrophobic mineral (talc). Finally, the talc recovery obtained as 94% by the use of JDFC, while Jameson cell has the recovery of 77% for the particle size of 250 µm."

Jan 1, 2014