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By S. M. H. Haddad, M. Saghaeian, R. Hejazi

"Due to limited water resources in mineral processing plants, increasing the efficiency of iron tailing dewatering system in order to obtain more usable water is very important. In this research, flocculation process, as one of the best method for water recovery, was investigated by using one organic cationic and three anionic types of flocculants with different dosages (10, 20, 30 g/ton) and iron ore tailing slurry (8 % w/w solidity) in series of graduated cylinders and obtained results were compared. it is observed that for tailing sample, the anionic flocculants have a significant effect on the settling rate of solid particles compared to the cationic types. Furthermore, effects of temperature (10, 23 and 43 ºC) and pH (2, 7 and 12) on the rate of settlement and turbidity was studied by utilizing the best flocculant. Findings showed that by increasing the amount of flocculant dosage, sedimentation rate went up and turbidity decreased. Comparison the out-put data of tests indicated that at temperature 23 °C and pH=7, the best results for both of faster settling rate and producing more clear water were reached. It should be considered that lower temperature has a more negative effect on the sedimentation rate and overflow water quality. So it will be predicted that for cold seasons, the consumption of the flocculant powder would be higher than the moderate or warm seasons. Also, coagulation/flocculation was evaluated by applying organic cationic coagulant added to anionic flocculant in three different times. Different ratios of coagulant to flocculant was examined (70%:30%, 50%:50%, 30%:70%) as well. Overall, utilizing coagulant had desirable influence on sedimentation experiments and obtaining high quality water. 1- INTRODUCTION In mineral Factories, among the important processes, solid-liquid separation process due to limited resources of water is very important and operates on iron ore tailing in settling tank as thickener in order to gain clear water for preventing water loss and applying again in plant units (Ciftci and Serhat Isik, 2017). Iron ore tailing include inharmonious compounds of iron oxides, silicates and aluminum-silicates. The percentage of solid in the tailing usually changes between 5 and 10 percent (Dash et al., 2011). One of the major problem about dewatering of tailing in thickener is solid-liquid separation of slurries with high stability because of fine particles. It means that rate of sedimentation is very slow and recovered water quality is inappropriate. This issue can be solved to a large extent by factors such as pH control, ionic strength and flocculation/coagulation process which results in aggregation of particles so that settlement by gravity is happened (Dash et al., 2011)."

Jan 1, 2018

By Liu Jin

"In order to reduce the dosage costs and increase the economic benefits of a full floating process in a low grade complex multi-metals W-Mo ore, based on the study of the magnetic characteristics of the main gangue minerals, a joint process by magnetic separation and floatation technology is adopted. The first step of the flowsheet is to recover magnetite by magnetic separation, then remove the magnetic gangues by high gradient magnetic separation (HGMS) and finally recycle the W-Mo from the magnetically separated tailings by flotation. Results of the lab experiment show that in comparison with the full flotation process, the floatation feed tonnage of the combined flowsheet are only 46%. The grade and the recovery of Mo in the rough Mo-Bi concentrate is similar with the original one while the WO3 in the tungsten rough concentrate is increased from 3.85% to 6.64%. The dosage costs are reduced by 53%. The new process can significantly improve the plant economically. INTRODUCTION With the continuous exploitation of scheelite resources, the easily obtained resource is gradually depleting. The ""lean, fine and complex"" scheelite resource has become the main research object of tungsten beneficiation (Lou et al., 2010) (Xiao et al., 2013). The passive situation of tungsten beneficiation can be reversed by enhancing technological innovation and investment in science and technology, changing the business philosophy and mode of production, optimizing the process structure and process conditions, improving various technical indexes and reducing production costs of reagents, which will maximize economic benefits (Jiang et al., 2012 ) (Zhu et al., 2016). A polymetallic deposit of tungsten molybdenum bismuth original process adopts a full flotation process. With the expansion of stope mining, the content of valuable metals in the ore continues to decrease, resulting in a high cost per ton of concentrate produced and a low output value (Li et al., 2015)."

Jan 1, 2018

By B. Courtaud

Uranium plants are generally located quite far from conversion plants, resulting in high transport costs. One way of reducing such costs involves putting more uranium in each drum by increasing the final product?s bulk density, which can be achieved via mechanical compaction of the powder and/or improved precipitation. For most of its plants, AREVA has adopted the latter approach, involving a new precipitation process applicable to all kinds of yellowcake produced (via uranates, uranium peroxide, etc.) that achieves bulk density of 3 for soda uranate, for example, instead of the 1.5-2 density available via conventional processes. At present, this process is being undertaken on an industrial scale and will become the norm for precipitation processes at all AREVA U plants. In addition to producing dense yellowcake, the process also delivers a product with no fines, a spherical particle shape (which allows for high-quality powder flow) and superior post-precipitation filterability

Sep 1, 2012

By Jun Xiao, Yucai Zhou, Yanhong Dong, Daixiong Chen

"In this article research on the mineralogical characteristics and beneficiation process for a copper molybdenum deposits in Heilongjiang province of China has been carried out. Chemical analysis, X-ray diffraction(XRD) and Quantitative Evaluation of Minerals by SCANning electron microscopy (QEMSCAN) were conducted on samples. The bulk analysis results showed that: (1) Comprehensive recovery and separation of molybdenum & copper minerals was feasible by the flowsheet of first copper-molybdenum mixed flotation after grinding the ore samples to 61.88% -74µm and then separation flotation of copper and molybdenum after regrinding the Cu-Mo mixed concentrate to 83.26% -38 µm. The results highlight relationship between mineralogical characteristics of ore and comprehensive benefication method. (2) the combination of "" Sodium sulphide and sodium thioglycolate "" reagents are effective inhibitors of copper-molybdenum mixed concentrate, their selectivity performance is better than that of single reagent. Good flotation indexes were gained that might be directive experience to similar cases. KEYWORDS Mineralogical characteristics; molybdenite; chalcopyrite; flotation separation; inhibitor INTRODUCTION Copper and molybdenum metals are widely used in the field of metallurgy, aerospace, national defence, etc. Copper –molybdenum sulphide ore is an important source of copper and molybdenum metals. At present flotation is still the preferred methods to deal with the copper-molybdenum sulphide minerals (Yin, Z., Sun, W.et al. 2017; Ansari, A., & Pawlik, M. 2007; Hirajima, T., Miki, et al. 2017). Copper and molybdenum sulphide ores usually have the similarly good nature floatability(Miki, H., Matsuoka, et al. 2017), meanwhile most molybdenum sulphide ore has tiny grained-size(Castro, S., & Laskowski, J. S. 2015.) and closed relationship with other sulphide and gangue minerals which result in the coper and molybdenum separation problem. Base on the nature of the ore and flotation properties of the individual minerals contained in the ore, there are two methods are usually used to deal with the copper – molybdenum ore, including bulk molybdenum/copper flotation followed by separation of copper and molybdenum minerals, sequential molybdenum/ copper flotation(Liu, G. Y., Lu, et al. 2012). The latter method has the shortcoming of hard to activate the depressed copper minerals in the previous molybdenum flotation process, which may result in a large loss of the copper from concentrate, this disadvantage would avoid while adopting the former method. bulk flotation followed by separation process is to get the copper-molybdenum concentrate from raw ore first then separate flotation the molybdenum and copper minerals, benefit from the excellent floatability of both molybdenum and copper minerals ( Malhotra, D., Rowe, R. M., & Bhasin, A. K. 1986; Cao, Z., Yue, et al. 2017 ), it really easy to obtain the mixed concentrate. In order to get the high quality copper and molybdenum concentrate, the key technologies and difficulties are to realize the separation of the copper and molybdenum minerals which have the very similar natural floatability (Raghavan, S., & Hsu, L. L. 2015; Nakhaei, F. 2014)."

Jan 1, 2018

By Longjiang Li, Mao Jiang, Zhihui Shen, Guizhou University, Yin Pan

This paper aims to study the influence and action mechanism of collector Y-89 on the flotation of auriferous pyrite. The measurement to identify adsorption quantity, IR and XPS spectrum analysis and flotation experiment made on the pure mineral were conducted. The research results show that when the dosage of MIBC is 20 mg/L, and pH of 8, the dosage of Y-89 over 60 mg/L, the pyrite recovery rate is greater than 80%.And when the dosage of Y-89is 60 mg/L, dosage of MIBC 20 mg/L, the pH in a wider range (pH=4~9), the recovery is greater than 75%. When the pH is above 10, the pyrite is depressed and its recovery rate is found to be dramatically dropping. Adsorption quantity measurement also shows that the strong alkaline conditions inhibit the adsorption of Y-89 on the surface of auriferous pyrite. IR measurement indicates that chemisorption of Y-89 occurs on the surface of auriferous pyrite. XPS spectrum analysis indicates that without adding Y-89, oxidation reaction occurs on the surface of pyrite. But when Y-89 is added, the chemical environment of Fe on the surface of pyrite has changed. Such findings indicate that with Y-89, chemisorption occurs on the surface of auriferous pyrite.

Jan 1, 2014

"A sulfide ore containing 2.19 g/t Au and 13.13 g/t Ag was studied for the deportment and extractive metallurgy of pay metals. The ore is composed mainly of non-sulfide minerals with approximately 10% pyrite and minor amounts of marcasite, chalcopyrite, galena, sphalerite, and arsenopyrite. The deportment study using a comprehensive approach showed that gold and silver minerals in the ore are mainly native gold (Au), electrum (Ag, Au), petzite (Ag3AuTe2), hessite (Ag2Te) and volynskite (AgBiTe2) with minor amounts of acanthite (Ag2S) and argyrodite (Ag8GeS6). At current grinding fineness (P70@150µm), the majority of liberated native gold and electrum particles are larger than 20µm with 85% >40µm, some of them are even larger than 150µm, indicating that extended leach time may be required to dissolve these coarse gold grains to completion and gravity concentration needs to be considered. Petzite is relatively fine with all liberated particles being less than 80µm in size. Gold distribution and leach testwork results showed that liberated gold accounts for 50% of the head assay, exposed gold associated with sulfide and non-sulfide minerals accounts for 30% and 11% of the head assay, respectively. Gold locked in sulfide and non-sulfide accounts for 4.8% and 4.1%, respectively. Thus, it is recommended that cyanide leaching be considered for extracting gold in this ore. This paper presents the detailed results of the gold and silver deportment study and extractive metallurgy testwork, and discusses the impact on process selection and flowsheet development.INTRODUCTION Gold and silver have been found in various geological environments and ore types. In terms of the gold deportment and extractability, gold ores can be classified into two categories and 12 types. Their extractive metallurgy is largely driven by mineralogical factors such as particle size, association with other minerals, coatings and rimmings, presence of cyanicides, oxygen consumers and preg-robbers, presence of refractory gold minerals and locking of submicroscopic gold in sulfide and sulfarsenide mineral structures (Zhou et al., 2004; Marsden & House, 2006; Zhou, 2013a, 2013b). Compared to gold, silver mineralogy is more complex and metal extraction is more challenging due to its presence in over 200 Ag-bearing minerals (a silver-rich ore body would typically contain 5 to 10 different silver minerals) (Gasparrini, 1984), presence of refractory silver minerals (such as pyrargyrite, prustite and stephanite) and submicroscopic silver in sulphide minerals (such as galena and copper minerals) (Chryssoulis et al., 1985; Di Prisco et al., 2002). The presence of these issues often causes problems in Ag-bearing ore processing. Therefore, knowing the deportment of gold and silver in the ores and metallurgical products is important for flowsheet development and process optimization."

Jan 1, 2018

By M. A. Wei

Effects due to some metal cations, such as Ca2+, Mg2+,Fe2+, Fe3+, Cu+, Cu2+, Zn2+ and Al3+ etc., which are produced in the pulp when the complicated ore is dipped in the deioned water, on the flotation of chalcopyrite and galena with the isopropyl thionocarbamate as collector have been studied. It is found that chalcopyrite floatability is not influenced by these cations, and the cations can be di-vided into three types according to their effects on the purified galena?s floatability with the same conditions, as activator such as Cu+; as depressant such as Al3+, Fe3+, Fe2+, Zn2+ and Cu2+ , unaffective such as Mg2+ and Ca2+.

Jan 1, 2014

By P. Altinkaya, O. Celep, H. Deveci, E. Y. Yazici

"Copper-bearing gold ores have proved difficult to treat in conventional cyanide leaching process owing to prohibitively high consumption of cyanide by readily soluble copper minerals interfering with gold extraction. Cyanide leaching of gold from copper-rich ores can be improved by intensive cyanidation or application of physical and chemical pre-treatment methods (e.g., ultra-fine grinding, roasting and leaching). In this study, nitric acid leaching of a copper-bearing pyritic concentrate (0.3% Cu, 62 g/t Au and 110 g/t Ag) as pre-treatment prior to cyanidation was investigated. The mineralogical analysis of the concentrate has revealed that it is composed of mainly quartz (25%), pyrite (32.7%), sphalerite (5.37 %), galena (2.12 wt %) and anglesite (1.13 %). Arsenopyrite, chalcopyrite, covellite and bournonite were also detected as minor phases present. Previous studies carried out by the authors showed that direct cyanidation (1.5 g/L NaCN) of the as-received concentrate (d80: 80 µm) produced low gold (45% Au) and silver (46% Ag) extractions. Nitric acid pretreatment of the as-received concentrate was performed for 5 hours at 1-5 M HNO3 and 80oC. In these tests, 65% Cu/Pb and 94% Zn were leached. The nitric acid pre-treatment improved the extraction of gold and silver in subsequent cyanide leaching (1.5 g/L NaCN) to 82% Au and 87.5% Ag with a concomitant decrease (up to 54%) in cyanide consumption. It can be inferred from these findings that the removal of copper from the concentrate and the liberation of the occluded gold by nitric acid leaching as pretreatments can be appropriately employed for the treatment of the copper-bearing pyritic concentrate with high gold/silver content. INTRODUCTION Refractory ores are often characterized by their low amenability to cyanide leaching resulting low (often <80%) Au/Ag extractions (Bhappu, 1990; Celep et al., 2009; Marsden and House, 2006). One of the most common causes of refractoriness is the physical or chemical encapsulation of submicroscopic or fine-grained gold in sulphide minerals such as arsenopyrite and pyrite (La Brooy et al., 1994; Vaughan, 2004). To render refractory gold ores/concentrates amenable to cyanide leaching, the decomposition of sulphide minerals within a pre-treatment process to expose the encapsulated gold grains to the action of cyanide is required. In this regard, roasting, pressure oxidation and biooxidation are commercially used processes for pretreatment of arsenopyrite/pyrite gold ores/concentrates ahead of cyanide leaching (La Brooy et al., 1994). Presence of cyanide-soluble copper minerals in the ore leads to low gold recoveries and high cyanide consumptions under conventional cyanide leaching conditions. High cyanide concentrations (CN:Cu>4) should be employed to achieve high gold recoveries from copper-rich gold ores/concentrates. This may require recovery of cyanide from effluents. In this regard, SART Process (sulphidisation-acidification-recycling-thickening) employed industrially for recovery of cyanide and copper as a by-product. This process simply relies on the addition of sulphide (as Na2S, NaSH or H2S) to the leach solution prior to the acidification to pH 4-5. This results in the precipitation of copper as Cu2S and occurrence of free cyanide which is recycled to the leaching circuit after S/L separation (Dai et al., 2012)."

Jan 1, 2018

By M. R. Hosseini

Among all gravity concentration methods, Shaking-table is the most effective one from the metallurgical point of view. In this study, Shaking-table was used for concentrating a manganese ore sample from Ghasem Abad area in Kerman, Iran. Experiments were designed by using L8Taguchi design with five variables and each variable in two levels. The influence of each variable and their interactions on the operation of the device was studied. The variables under investigation were: table slope, table frequency, water flowrate, feed rate, and particle size. The manganese concentrate grade and recovery were used as response variables. It was shown that water flowrate, table slope, feed rate, and particle size are the significant variable on concentrate grade; while, all the variables influence manganese recovery.

Jan 1, 2014

By I. Baltov

The Chelopech Mine, located in Central Bulgaria, is one of Europe?s largest underground copper- gold mines, currently mines and processes approximately one million tonnes of ore per year. Copper concentrates containing gold and silver with significant arsenic content are produced and sold on the world markets. The operation was acquired by Dundee Precious Metals in 2003. A modernization program for the Mine and Plant has since been in progress to increase both production and mechanical reliability, and also to optimize metallurgical recoveries. Since then, mine output and plant throughput have doubled, plant availabilities and utilization have increased significantly, and overall metallurgical performance for copper and gold improved. The paper presents the systematic approach for obtaining and analyzing data, communicating it, and making suggestions for changes to the technological parameters and flowsheet in order to maximize metal recoveries. These studies have correlated the ore characteristics (particle size distributions, metal assays, and Cu and Au mineralogical and phase analyses) with the actual plant performance (kinetics, recoveries, and throughput). Cu recovery by size has been calculated from the mineralogical analyses of feed, concentrate and tails, with the most significant losses being identified in terms of fraction and phase recoveries. This has provided the recommendations (some of them tested in plant) for changes in the grinding and flotation circuits.

Jan 1, 2014

By Y. M. Zhao

The dry coal preparation technology based on air dense medium fluidized bed is an innovative dry highly efficient technology of coal by introducing fluidization technology into beneficiation.. A modularized 40-60 tones/hr system (KZX-40) with an air dense medium fluidized bed separator was established. The construction cost of the modularized system is only 1/3 of the traditional wet heavy-medium coal preparation plant construction cost with the same capacity, and the consumption of dense medium can be reduced by 50%-60%. The footprint and volume of the system are decreased by 75% and by 89.81%, respectively. These facts show the improved intensity, efficiency and economy of modularized coal preparation technique. The modularized KZX-40 system can be used to separated 50-6 mm coal over a separation density range of 1.3-2.2g/cm3 depending on the variation of <1mm fine coal content. The separation efficiency of the system was investigated at low and high separation densities by float-sink experiments. The partition curve of the products at low separation density shows that the separation density (d) is 1.33g/cm3, and the probable error (E) value is 0.05g/cm3. Using 50-6mm coal with an ash content of 55.35% as feedstock, at a high separation density of 1.61g/cm3, the clean coal with an ash content of 14.67% and the tailing with an ash content of 71.26% are produced with an E value of 0.06g/cm3. It is clear that the modularized KZX-40 system is suitable for separating 50-6mm coal effectively. During the dry coal preparation process, the dust is removed by the bag collector and recycled back into the separator. So the system has the advantages to work without environmental pollution and water. The modularized system greatly reduces the construction and operating costs and decreases the area and volume required, compared with wet dense medium system. The system is applicable to the coals located in arid regions and also to the coals that tend to slime during wet separation. Keywords: modularized, KZX-40, fluidized bed, beneficiation

Sep 1, 2012

By E. V. Konoplev

Russia is one of the largest manufacturers of potash fertilizers in the world. Processing of potash ores basically is carried out by a flotation way. Potash ores of the Verhnekamsky potash deposit contain sylvite (KCl), halite (NaCl) and water insoluble impurity (up to 8 %), presented silicate- carbonate minerals, easily slimed by ore crushing in rod mills. Ore of some deposit sites, developed by ?Uralkali? Co, contains a significant amount of a carnal-lite (KCl.MgCl2.6H2O). Work practice of flotation potash plants has shown, that increase of car-nallite content in ore leads to increase magnesium ion content in a circulated plant brine (saturated salt solution of potassium and sodium chlorides-sylvite brine) and to deterioration of indicators of ore processing. In the Paper we present results of researches of floatability of various size sylvite and influences silicate-carbonate slimes on sylvite flotation in brines with the various magnesium chloride content. It is established, that with increase magnesium chloride content in the brine mainly decreases the flotation of coarse sylvite and amplifies adverse effect of silicate-carbonate slimes on sylvite flotation. In this connection deterioration of potash ore flotation is caused not only decrease sylvite floatability but also strengthening adverse action of water insoluble impurity on potash flotation. It is shown that the application of frothing agents, which render the stronger dispersive action on the colloidal-micellar structures of amine in the salt solution increases the absorption of collector on the sylvite and improves the flotation of sylvite into magnesium the containing salt solutions. The application of an organic depressor to the greatest degree of decreasing the absorption amine on the slimes and to the smallest degree possessing anionic properties proves to be most effective with the flotation carnallite of the containing potassium ores. It is offered the application the frother on a basis of glycol ethers and organic depressor on a basis of urea-formaldehyde resin. Results of industrial tests developed reagent modes for flotation treatment of ores with the raised carnallite content (to 6 %) on ?Uralkali? Co plant have shown that their application has allowed to receive equivalent indicators potash flotations at change of magnesium chloride content in the brine from 0.7 % up to 2.4 % .

Jan 1, 2014

By V. A. Luganov

The processing of pyrite concentrates may be accomplished with a preliminary enrichment due to withdrawal of iron sulfides from initial concentrate. The withdrawal of iron sulfides is possible with the help of magnetic separation when the magnetic iron sulfides (pyrrhotites) separate from non-magnetic fraction?the sulfides of non-ferrous metals. The conversion of non-magnetic pyrite to magnetic pyrrhotites of adjusted structure occurs because of roasting with the withdrawal a part of sulfur. The resultant pyrrhotites are strong ferromagnetic with hexagonal structure. This paper contains results of studying magnetic properties of pyrrhotites, findings of the de-composition of initial pyrite with the consequence homogenizing of product and hardening of product of decomposition in liquid nitrogen. Define the type of pyrrhotites super lattice structure, was established that at increasing of sulfur contention the magnetic properties of pyrrhotites are change ? pyrrhotite begin evince the properties of antiferromagnetic material and magnetic susceptibility of material is interruptive changing from 0 to 1.7×10-3~1.9×10-3 SI units/g system. The paper contains results of technological researching of pyrrhotizing-magnetizing roasting of pyrite concentrates in different condition: in unoxidizing, with the decreasing contention of oxygen and in autogenous regime in air atmosphere. Established that the magnetic susceptibility of products depend of pyrite degree dissociation and achieve 2×10-3~3×10-3 SI units/g system at 40%~45% desulphurization. Defined the conditions of steady regime autogenous roasting, at resultant products possess coefficient of magnetization, and this can be used magnetic separation for separating ferromagnetic pyrrhotites from gangues and non-magnetic sulfides of non-ferrous metals. This technology of pyrrhotizing roasting is recommended for effecting in industrial condition. The initial materials and products of roasting subjected to chemical, thermogravimetric, Moss-bauer spectroscopy and X-ray analysis, defined the magnetic properties.

Jan 1, 2014

By A. Chattopadhyay

Porphyry copper deposits are the most important source of copper and also gold within the orogenic belts. These deposits are characterized by low grade copper ore (0.2% to 1% copper) and host precious metals like gold, silver and platinum. A characterization tool has been developed for complex porphyry type copper-gold deposits with an aim to quantify copper and precious metal deportment. This quantification is a challenging task due to low grades and also due to the presence of both visible and invisible precious metal content in different sulphide and non-sulphide phases. Studies were initiated on various complex, ore deposits and also on samples representing various processing streams such as flotation and leach products. This paper presents the development of a systematic approach to diagnostic metallurgy using state-of-the-art optical and automated mineralogy, micro-probe, dynamic SIMS and TOF-SIMS analysis for quantification of invisible and visible precious metals content in various species. The relevance of this characterization to metallurgy will also be briefly discussed in this paper.

Sep 1, 2012

By Ji Hoe Kwon

Hindered-settling column process has been used for the separation of mineral particles for many years. In this study, a model based on DEM (Discrete Element Method) was used to simulate the movements of particles in a settling column separation process. DEM model is capable of describing the motion of particles and calculating the impact forces between particle and particle, and particle and wall contact under various operating conditions. Drag force, rotation induced Magnus lift force, contact force and the gravitational force acting on individual particles were considered for establishing the mathematic models. The overall motion of the particles in water is satisfactorily simulated by the DEM, and the result shows good agreements with the result of lap-scale experiment.

Jan 1, 2014

By Homero Delboni, Mauricio Guimarães Bergerman

"At the regrind or secondary grinding stages of metal ores, usually the concern about the energy efficiency is extremely important. In general, the specific energy consumption at this stage is higher than in the primary stages of grinding. In addition, it has been observed a tendency of finer grinding in the new projects of metallic ores, due to the finer mineral liberation, which leads to an increasing capital and operational costs. In Brazil, during the last 10 years, regrind P80 sizes of approximately 40 to 20 ?m have been necessary in all new iron and copper plants. The use of conventional ball mills, adequate for primary grinding, can result in low energy efficiency with products below a P80 of 50 ?m. This paper presents an overview of the regrind circuits in Brazil that operate with vertical mills and the results of industrial surveys at selected vertical mill circuits. Where possible, comparison is made between the plant results and the design criteria and laboratory regrind tests used for the mill sizing."

Jan 1, 2014

By Sanjay Agarwal

Polymetallic sea nodule is a good source of manganese having 22.3% Mn, 5.4% Fe, 6.14% Si, 1.1% Cu, 1.15% Ni, 0.076% Co & others. These nodules are available in plenty on the sea beds, generated over thousands of years. Nodules mainly consist of manganese dioxide and iron oxide phases and other metals are entrapped in the complex cage of these Mn-Fe oxides phases.World-wide research is going on sea nodules as an alternative future source of many metals, which has led to intensified efforts to develop practical and economical processes for recovery of these metals. Very little attempt has been made world over on manganese extraction from sea nodules. We at CSIR-NML, India have developed a direct reduction smelting process of nodules to recover manganese as Fe-Si-Mn alloy. This alloy is extensively used in the steel industry as deoxidizer and as an alloying element to make different grades of steel. Lab scale experiments were conducted to study the manganese distribution between Fe-Si-Mn alloy and FeO-SiO2-MnO-CaO-Al2O3-MgO slag at different temperatures by varying the basicity of the slag. The direct reduction smelting of nodules is carried out on 1 kg scale in 50 kVA electric arc furnaces to have Cu, Ni and Co alloy and Mn rich slag. This manganese rich slag is further treated in 50 kVA arc furnace to produce Fe-Si-Mn alloy. The composition of Fe-Si-Mn alloy obtained is in the range of Mn: 60-65%, Si: 14-17%, C: 1- 2%, S: 0.03% and the balance iron which is acceptable grade for industrial use. The equilibrium amount of MnO in the final slag is controlled by the temperature of the final melt and the silica, lime, alumina content of the final slag. Increasing the basicity of the slag plays a major role in minimizing the loss of MnO to slag, increasing the amount of SiO2 of the slag, increases the manganese loss in the slag. Increasing the CaO/ Al2O3 ratio of the slag decreases the manganese loss in the slag. Increasing the MgO/CaO ratio of the slag increases the manganese loss in the slag. Thus by adjusting

Sep 1, 2012

As the high-grade copper deposits in the world become depleted, there is an increasing focus on developing lower-grade ores which contain problematic impurities such as arsenic. Indeed, many high-arsenic ore bodies remain undeveloped because of their high arsenic levels. Separation of arsenic from copper at the beneficiation stage is difficult using the current processing techniques and invariably copper concentrates produced from these ores contain high arsenic levels, sometimes so high as to make them unacceptable to copper smelters. New, sustainable approaches are needed to allow development of high-arsenic copper ores while at the same time minimizing arsenic dispersion into the biosphere. The most promising route for separating arsenic from copper at the flotation stage centers on the use of pulp potential control. In recent years CSIRO has demonstrated at laboratory scale, for a number of high-arsenic copper ores, that crude flotation separation between arsenic-bearing minerals and copper-bearing minerals can be made. The differences in specific minerals responses to changes in pulp potential and pulp pH are exploited for this separation. Such separations nominally produce a low-arsenic high-copper concentrate and a low-copper high-arsenic concentrate. The former can be smelted without penalty, while the latter requires further treatment to capture and stabilize the arsenic and to recover the residual copper. Consideration has now been given to treatment options for the high-arsenic low-copper product. A pyrometallurgical option, involving a low temperature roast to selectively fume off the arsenic and leave a high-copper calcine, coupled with an arsenic stabilization step, has been demonstrated in the laboratory, and a preliminary techno-economic evaluation shows the process to be economically favorable. These results were recently published. CSIRO has now investigated an alternate hydrometallurgical option, involving atmospheric alkaline sulphide leaching of the high-arsenic product, where the arsenic is selectively dissolved leaving a copper-rich leach residue. The leaching testwork was carried at on the same high-arsenic low-copper material used in the roasting studies, enabling a techno-economic comparison of the alternate roasting and leaching treatment options to be made. The alkaline leaching work is briefly described and the results of the process option comparison presented and discussed. Keywords: arsenic, copper, economics, enargite, ore flotation, tennantite INTRODUCTION

Sep 1, 2012

By S. Banisi, A. R. Hasankhoei, A. R. Ghasemi, E. Razi, G. A. Parsapour

At the pelletizing plant of the Gol-E-Gohar mining and industrial company two parallel dry closed grinding circuits, each with the average throughput of 300 t/h, are used. Each grinding circuit consists of a burner, a dry ball mill (6.2 m×13 m) and an air separator. The ball mill consists of a 2 m-long drying and an 11 m-long grinding chambers. The iron ore concentrate is fed to the drying chamber by a feed chute. The first 1.35 m of the drying chamber consists of 56 flights in two rows to decrease the feed moisture from 5% to below 0.2% and the rest of dryer chamber consists of deflectors to pass the material to the grinding chamber. It was found that when the feed moisture content increased from 1.3% to 3.5%, the throughput decreased by 12% (35 t/h) indicating a low performance of the dryer. Monitoring the wear rate of flights for a period of 12 months showed that the first 0.8 m length of the flights length did not experience any wear. This means that about 59% of flights do not contribute in material distribution and drying. A DEM simulation of feed chute and the drying chamber demonstrated that due to the inappropriate design of feed chute most of materials do not get in touch with the first 0.8 m length of the drying chamber. To overcome this problem, various feed chute designs with different geometries were simulated by the KMPCDEM© software. With the aim of arriving at a proper material trajectory, where the total length of the flights are used, a new feed chute was selected. The simulation results indicated that if the height of the flights is increased from 1.60 to 2.26 m and the slope is increased from 45 to 48 degrees the material arrive at the first 0.48 m of drying chamber. In this manner, the unused part of the drying chamber decreases from 59% to 36% of the length. After installation of the new feed chute, during a period of three months the throughput increased by 36 t/h.

Jan 1, 2018

By Johannes Käppeler, Alexander Weh, Harald Giese, Klaas van der Wielen

"High voltage breakage is a novel comminution technology that shows considerable potential for various applications in mineral processing and recycling. It relies on highly energetic electrical pulses to induce electrical explosions in rocks, causing the formation of plasma channels and powerful shockwaves that result in fracturing and fragmentation of particles. There has been considerable interest in this technology for its ability to weaken rocks and preferentially liberate target minerals, as previously demonstrated by other authors. Weakening has the potential to significantly reduce energy consumption during comminution, or alternatively, increase throughput of a circuit. Better liberation may yield an improved concentrate grade and/or recovery, reduce fines losses to tailings or increase the grind size required for liberation, thereby also reducing energy requirements. This review aims to increase general awareness of this technology, and promote further research into it by providing a general summary of the theory underlying high voltage breakage, as well as the envisaged applications of this novel technology. A brief history of the technology is presented before progressing onto the physics underlying the electrical and geomechanical aspects of high voltage breakage. Key results regarding weakening and liberation are discussed, and some mention is made of current challenges and planned developments to increase the throughput of high voltage breakage units to the level where it can be integrated into new and existing processing circuits."

Jan 1, 2014