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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 T. V. Vijaya Kumar

Most of the iron ore washing plants set up in India consist of sizing of the ore by dry / wet screening, washing and classification by screw classifiers followed by multi-stage hydrocycloning of screw classifiers? overflow. The underflow of the hydrocyclones forms the concentrate which is used for pellets making. As the cut-point of hydrocyclones is generally at 20 microns, relatively lower diameter cyclones in a cluster with parallel feeding are used. This is found to, often, result in choking of the spigots, at times, by extraneous material as well reporting along with the slurry. It is observed to lead to sub-optimum performance of the cyclones with loss of iron values into cyclone overflow and thereafter into tailings. Analysis of data of an operating beneficiation plant (Beneficiation Plant - 1, BP - 1) of M/s Jindal South West (JSW) Steel Limited, Karnataka, India over a period of one year indicated that weight recovery of the concentrate from two-stage hydrocycloning was 45% (with respect to feed to the cyclones) and assaying 63.43% Fe, 4.43% SiO2 and 2.32% Al2O3. CSIR - National Metallurgical Laboratory - Madras Centre (NML - MC), India developed flotation process for processing screw classifier overflow with a view to improve the recovery of ultrafine iron values from it. Detailed investigations at laboratory scale flotation column and on-site pilot plant trials proved that the production of the concentrate and productivity of the plant could be enhanced by adopting reverse cationic column flotation technology. It was established that a typical screw classifier overflow analyzing 60.54% Fe, 5.95% SiO2 and 3.67% Al2O3 could be upgraded to 64.21% Fe, 2.64% SiO2 and 2.20% Al2O3 with 82.7% weight and 84.7% Fe recovery. This translates into an increase of 80% in weight recovery of the concentrate compared to what is being obtained, at present, by two-stage cycloning. It had an added advantage of reduced SiO2 in the concentrate. Thus, SiO2 in the final concentrate could be reduced to 2.64% from 4.43% by adopting column flotation. The concentrate so generated meets the stringent specifications, especially in terms of alumina in it(should be less than 2.50%) for pellets making. Later, JSW Steel Limited set up the flotation plant in BP - 1 based on the process developed by NML - MC and was commissioned in 2010. Keywords: iron ore, washing, ultra fines, alumina reduction, column flotation

Sep 1, 2012

By E. N. Levchenko, K. V. Bulatov, G. I. Gazaleeva

"The article highlights the new procedure of choosing special methods of disintegration for very complicated rare ores. Especially, there is a complex approach to criteria of comparing different types of crashing machines. Criteria are as follows: degree of crashing, degree of grinding, degree of slime creation and the new creations criteria for selection. The procedure is tested on tantalium and zelestinium ore. A new methods was shown the preference impact crashing machine MD-7 for zelestinium ore. In product of crashing 0,050 mm size are 9.4 %, 0,005 mm size – 2.5 %.This is gut result, and show the best indexes of slime creation. The check of the proposed new methods was done when comparing the crushing of tantalium ores of the Otboynoe deposit in cone inertial and centrifugal impact crushers and roller press. A research was conducted different crashing machines under laboratory and semi-industrial conditions. Different operation modes of crushers were also checked. The comparison of the crushing selectivity criterion and degree of crashing showed that the centrifugal impact crusher (CC) and roller press (RP) has the lower indexes than the cone inertial crusher (CIC). Degrees of crashing for different crashing machines are:centrifugal impact crusher – 10.1;roller press – 10.8;cone inertial crusher – 14.2. The crusher CC the selection criterion is 0,632 units, RP – 0.730, where as for the crusher CIC – 0.848 units. The best crushing mode for the crusher CIC-300? is as follows: turning angle of debalance is 0°, peripheral speed of a cone rotation – 1420 r.p.m, specific energy consumption- 4.99 kW*t/t. Crusher CIC-300? is the best for tantalium ores of the Otboynoe deposit. The new methodic are verification.INTRODUCTION Rare metal concentration circuits are complex, energy-intensive, and costly. The reasons for this include the multimetal nature of the rare metal ores, their volatile material composition and low grade. That is why careful selection of ore preparation circuits is of great importance for minimizing the energy inputs and maximizing the liberation of the valuable minerals. Recent research (Khopunov, 2011, Gazaleeva, 2015) has demonstrated that the most efficient comminution processes available to the mineral processing industry are selective disintegration, crushing and grinding, allowing to liberate the mineral particles along the cleavage in the comminution process. These methods are indispensable for the processing of finely disseminated refractory ores prone to slime formation. Many rare metal ores found in Russia are of this type."

Jan 1, 2018

By V. Sheikhzadeh, M. Saghaeian, R. Hejazi, A. Abazarpoor

"The morphology of pellet feed is known to influence several properties of iron ore pellets, including drop number, compression strength and reduction properties. The effect of grinding method on the particle size, specific surface area and particle shape of an iron ore concentrate has been studied. Grinding methods considered were ball mill and high pressure grinding rolls (HPGR). 5 different samples (Blaine= 1800, 1900, 2000, 2100 and 2200 cm2g-1) were used to compare the size, shape and specific surface of particles for each grinding methods. Particle size was meticulously examined by sieve analysis, laser sizing and image analysis. In samples ground by ball mill, it has been observed that the fraction of particles <10µm increases from 39% to 46% as the Blaine number increases from 1800 to 2200 cm2g-1. For HPGR samples, corresponding figures are 41% and 49% respectively. It is evident that for samples ground by HPGR, the fraction of fine particles is higher than that of those ground in the ball mill. To measure the specific surface area of particles, Blaine and BET were used. A higher real surface area was observed from HPGR treatment in comparison to ball mill grinding, provided by a higher porosity, cracks, roughness and new surfaces. Shape factor of particles was determined using three descriptors: circularity, roughness, and aspect ratio. It was also observed that particles produced by HPGR were more elongated, less circular, and rougher than those processed by the ball mill. 1. INTRODUCTION Particle size and shape are important parameters that can significantly affect mineral processing plants and pelletizing plant performances (Brozek and Surowiak, 2007). It has also been proposed that green pellet quality is directly related to the amount of fine particles in a pelletizing plant (Dwarapudi et al., 2008; Umadevi et al., 2008; Gul et al, 2014; Van der Meer, 2015). Particle size distribution and specific surface area (SSA) are also key factors that can control the quality of green and fired pellets. Processing circuits generally reduce the fines fraction to lower than the required specific surface area (SSA). It is therefore necessary to regrind concentrates to produce suitable pellet feed for further treatment in the balling and firing steps [1, 2]. Two main methods are used for re-grinding of concentrate; high pressure grinding rolls (HPGR) and ball mill. Given the different forces used to reduce size in these methods, it is expected the shape of particles will be different (Pourghahramani and Forssberg, 2005)."

Jan 1, 2018

By G. M. Savov

Copper recovery via Solvent Extraction (SX) process is a well-known technique, which is practiced commercially for approximately 40 years. The main advantages of the SX process are: an absence of using high temperatures (not more than 50oC), complete recycling of barren solutions, very high selectivity at removal and purification of metals, metal conversion to a form, suitable for easy subsequent recovery. Despite the wide acceptance of Solvent Extraction, it does have some shortcomings, namely SX employs large quantities of volatile organic substances. This leads to huge plant ?footprint? of SX circuits. The SX is very sensitive to low temperatures and quantity of total suspended solids in feed solutions and this causes bad phase disengagement times and therefore poor process performance. And most important ? low grade solutions (with copper concentrations below 1 gram per litre) are very unattractive for treatment via Solvent Extraction process, because capital costs for SX plants are not a direct function of copper tenor in the feed. The combined process with Ion Exchange (IX) and Solvent Extraction ? further called IX/SX by IONTECH Engineering was developed. This combination eliminates SX negative characteristics shown above. The key findings of the IX/SX process are; (1) recovery of copper from pregnant leach solutions with metal concentrations, varying on a broad range; (2) can be operate effectively with very low-graded copper pregnant leach solutions; (3) the construction of large ponds for pre-filtration of the pregnant leach solutions is not necessary; (4) working with turbid solutions and solutions containing large quantities of suspended solids; (5) winter operation without troubles; (6) strongly reduces organic losses; (7) technological approach lead to plant sizes reducing. This paper reviews the main stages of the combined IX/SX process, the applications for metal recovery and introduces projects where the process is applied. Key findings of the process are also presented. A comprehensive technical comparison between processes has been carried out. This paper also evaluates the advantages of using combined IX/SX technology as a purification and separation step in copper production process, with the aim of further increasing the overall copper recovery and to ensure that a suitable copper concentration in the advanced electrolyte is achieved. The implementation of a Ion Exchange (IX) process in the classic Solvent Extraction-Electrowinning process(SX/EW) is found to be beneficial for processing of pregnant leach solutions both from spent and new copper heaps and dumps, acid mine drainage and other industrial streams without any preconditioning as well as in terms of achieving improved recoveries and copper cathode quality.

Sep 1, 2012

By John Starkey

"The development of the Minnovex SAG Power Index (SPI) Test in Canada has created a new and accurate way to measure ore hardness in a SAG environment. Coupled with the standard Bond Ball Mill Work Index Test it is now possible to determine from these two relatively inexpensive laboratory tests, the amount of energy required to grind an ore from the primary crushed size to the desired liberation size. By testing multiple samples to discover where the hard and soft ore zones are located, the representative hardness distribution for the ore body can be determined and grinding equipment selected that is capable to meet production targets for all mining areas of the deposit.In the last century, it was deemed prudent to select up to 50-tonne samples of ore for Semi Autogenous Grinding (SAG) pilot plant tests and a second 5-tonne sample for metallurgical pilot plant work. The methods outlined in this paper show how the 50-tonne sample is no longer required for SAG design because it is better to measure the representative hardness distribution of the ore body than the hardness of a “chosen” representative sample. This is demonstrated to be true by the fact that the lower horizons of an ore zone cannot initially be accessed to recover 150mm pieces, and the bulk samples taken, rarely represent the desired design hardness of the ore. Median hardness will not yield a good result in most cases and a clear understanding of the location and extent of the hardest zones is imperative to achieve a truly successful grinding circuit design.This paper discusses two plants (Kubaka and Agnico Eagle) that have been designed using SPI and Bond with no pilot plant test backup. Both have been very successful in that they can grind design tonnage on the hardest ores in the mine. This guarantees that the mine’s profitability is protected in any mining sequence. It also allows the extra power to be converted to higher throughput on softer ores if additional cash flow is needed. The advantages from the owner’s point of view are dramatic. This procedure will be attractive as well when considering the costs for test work, and the samples for testing. Existing diamond drill core is normally used and this allows the inclusion of data from the deepest regions of the mine. In addition, it is also clear that the decision to use a crusher can now be made at the design stage of a project based on measured parameters and the estimated capital costs of alternative designs."

Jan 1, 2003

By Thi Hien Tran, Ngoc Anh Tran, Thi Thu Hien-Dinh, Duy Anh Dao

"Natural graphite exists mainly in three types of amorphous graphite, flakes graphite and lump graphite. Graphite from Bao Ha mine, Lao Cai province is with 11.8 carbon-graphite %, forms in flakes and dense crystals. The paper presents the comparison among enrichment of Bao Ha’s graphite by two types of froth flotation with 90 g/t kerosene oil as collector, 50 g/t montanol 800 as frother, in natural pH, and at 20 wt.% of solid. The differences among three flow charts indicated that using middling flotation and re-middling flotation for the cleaning tails and the scavenger concentrates, the final concentrate was improved and containing up to 94 carbon-graphite % of flakes graphite, up to 82 carbon-graphite % of fine graphite with 93.48 % total recovery, moreover, carbon-graphite in the final tail is 0.86 % with 6.52 % recovery. However, a refining mill and a classification stage were applied before doing the middling flotation stages.1. INTRODUCTIONGraphite (black lead or plumbago) is used in many industrial applications due to its high electrical and thermal conductivity, lubricity, and chemical inertness. High quality graphite can be used in metallurgy, molding, brake linings, batteries, and fuel cells. Graphite, composed exclusively of the element Carbon, is rather common mineral, but fine crystals are rare. Graphite is soft (hardness: 0.5-1 Mohs scale); has a layered, planar structure; occurs in platy or acicular (occasionally) morphology (Minerals.net; Pierson, 1993).USGS (2017) reported, global consumption of natural graphite continuing increase in the period of 2012-2016. In 2016, China used up 35% while produced 66% of the world’s graphite. Graphite mining are being developed in Madagascar, Mozambique, Namibia and Tanzania, but production reduced in Canada from 2015. Based on the statistic of Statista.com shows from 2011 to 2014 and a forecast to 2020, the global graphite demand is mostly in refractories, batteries, others, lubricant/crucible, foundry and pencil, and that is increase every year.Graphite Vietnam lies on the fault of the Red river from Yen Bai to Lao Cai, the reserves is estimated around 29.3 million ton and mostly in Lao Cai province where has Nam Thi graphite mine and Bao Ha graphite mine (MOIT, 2008). Vietnamese graphite products could reached to ~92% graphite carbon (Cg) (Tran Thi Hien et al., 2008). Studied samples was collected in Bao Ha graphite mine, is located in Lao Cai province. Graphite Bao Ha exists in two types as graphite flakes and amorphous (Hoang Thi Minh Thao et al., 2017)."

Jan 1, 2018

By J. K. Wen

The Mojiang nickel deposit is situated in Mojiang county, yunnan province, a tropical area in the southwestern China. The known geological resource (measured and inferred) includes 4 million tons of nickel sulphide ore at an average grade of 0.50% Ni and 0.34% As. Nickel minerals are mainly gersdorffite (NiAsS), millerite(NiS), morenosite (NiSO4?7H2O), and more than 50% nickel minerals is gersdorffite which contains arsenic. Pyrite is the dominant sulfide mineral, 22% in the ore and part of nickel minerals are locked in pyrite. It is difficult to attain the nickel concentrate quality allowed by a smelter and the acceptable nickel recovery through flotation due to the high arsenic in the ore and the nickel minerals locked in pyrite. The shake flask tests and column tests for nickel recovery from the Mojiang nickel sulfide ore had been finished by the end of 2003. The column bioleaching tests were carried out using the columns with diameter of 120 mm and height of 1200 mm, and 70% of nickel dissolution was attained in 300 days. A 10,000t pilot test heap was built in the Mojiang Nickel Mine by June 2003. The ore used in the test heap was supplied from the underground mine and was crushed to -20 mm before being feeding to the test heap, which is 4~5 m high, 30 m wide and 50 m long. The naturally adapted mesophile and moderate thermophile have been inoculated into the leaching pile by June 2003. The operating temperature in the heap ranges from 38? to 55?, varying with season. 30% of nickel dissolution was achieved in 3 months and the typical pregnant leach so-lution (PLS) contains 2~3 g/L Ni and 20~30 g/L Fe. The Ni sulfide which was produced by us-ing Na2S 10 t/t?Ni contains 16%~20% Ni and Ni recovery was 98%. 70% of Ni dissolution is expected in 12 months. The test heap was scaled up to 60,000 t of ore per year by mid of 2005. In this paper metallurgical studies and performance of the pilot test heap will be described. The results of mineralogy, shake flask tests and column tests on bioleaching of the nickel sulfide ore are reviewed. The key test heap operation parameters, the performance of bio-heap leahing-Na2S precipitation, ferrous oxidation-ferric precipitation-Na2CO3 precipitation, bacteria in the pile, water and acid balance, effluent disposal, capital and operation costs are also discussed.

Jan 1, 2014

By J. C. Wilmot

The chemistry of aqueous cyanide containing species in the presence of reduced sulfur sources was investigated. Reactions were studied as a function of time and tracked using novel analytical techniques developed for the analysis of the species of interest in the sample matrix. Species of interest in this work were ?free? cyanide, thiocyanate, cyanate, metallo-complexed cyanide, and various sulfur and sulfur-oxygen species. The studies were conducted under constant temperature conditions at pH values above 10.5 and in the absence of light in order to eliminate photochemically induced cyanide reaction effects. The reaction products of cyanide when in contact with pure sulfur, pyrite, pyrrhotite, sphalerite, and cinnabar are reported. These data combined with data for cop-per, arsenic and lead sulfides, which have been previously published, are used to predict the fate of cyanides in refractory gold ores containing sulfides. A mass balance on the cyanide reaction products was completed for all experiments. From these studies insight into the chemical relationships of cyanide, the specific metal sulfide, thiocyanate and cyanate have been gained. When looking at the results of the experiments, it was immediately apparent that even though the overall conversions of cyanide to thiocyanate were different, most of the samples exhibited a similar pattern of reactions. The recovery of cyanide in these experiments was better than we expected, averaging in most experiments to be over 90 percent. Based on this research, the formation of cyanate was probably the main route of loss. The pathway choice for the various minerals was also elucidated in these experiments. The competitive oxidation pathways that cyanide can take have been tracked for sufficient time to determine the ?final? fate of cyanide in a process. Initially, it was thought that the cyanate may be involved in a hydrolysis reaction to form ammo-nia and carbonate ion, but upon further inspection of the data, there was not the drop-off in total cyanide recovery that would manifest itself were this to happen. On the other hand, the hydrolysis of thiocyanate to cyanate and sulfate would require oxidizing conditions which did not exist in solution at the beginning of the experiment. If the solutions were to be sufficiently oxidizing in nature for this to happen, the initial amount of cyanide would have been converted to cyanate quickly. The data support premise that the ?hydrolysis? of thiocyanate to cyanate and sulfate occurs to some extent in the experiments conducted. Knowledge gained from the experiments with pure metal sulfides helped interpret the results of experiments with refractory ores found in the State of Nevada.

Jan 1, 2014

By A. Kamali Moaveni, A. Boehm

"The current research study addresses a physical separation process of basic oxygen furnace slag “BOF-Slag” (also called Linz-Donawitz “LD” slag) for the removal of phosphorous. In order to recycle the LD-Slag into the steelmaking process, the physical dephosphorization methods are considered. The systematic geometallurgical investigation carried out on particle classes of decreasing size, which were further subdivided into susceptibility classes, shows intimate intergrowth beyond 40 µm due to slag´s fast cooling rate. Generally, LD-slag is constituted of oxide and silicate phases, among which di-calcium silicate “C2S” dominates. About 0.5 mass% of phosphorous was detected in the LD- slag, which is exclusively situated in the lattice of the C2S. The comparatively high grinding resistance (Rittinger index 12.9 cm²/J) of the LD-slag reflects the energy demand to liberate the phosphorous carrying phase from the iron containing minerals, while separation efficiency decreases in the very fine particle size range. Hence, the conversion of the phase intergrowth pattern was considered alternatively to re-build the mineral phase structure of the material to a separable interlock pattern. Carbon dioxide, as the most convenient substance in the steelmaking plants, was applied to digest the C2S in the presence of water and to reduce the basicity of the slag slurry. As a result of C2S-CO2 chemical reaction, calcite and calcium silicate are precipitated at the surface of the oxide particles and carried between 25% and 75% of the total phosphorous, respectively. Accordingly, the mobility of the phosphorous content phase to the top of the grains serves the possibility to separate the precipitated layer by mechanical attrition process. This paper is focused on the impact of the intergrowth pattern alteration on the slag´s separability characteristics. Moreover, the operating conditions of the carbonization process, such as temperature and pressure, on the deconstruction kinetics of the C2S and the generated products will be considered. INTRODUCTION The steelmaking BOF-Slag is a stony aggregate waste from the steel manufacturing process. The investigated slag by the current research work was provided by voestalpine steelmaking plant in Linz, Austria. The current legislative situation in Austria restricts the historical application of slag as construction material, due to the high level of heavy metals content and corresponded land-filling and transportation costs of slag, therefore the demand for recycling of BOF-Slag to the steelmaking process has raised, as an alternative application. Statistically, voestalpine steelmaking plant produces about 130 kg BOF-Slag per ton of final steel product, which serves annual BOF-Slag production of 700,000 ton, approximately."

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 Haghi. Hamed, Javad Vazifeh Mehrabani, Mehdi Sedigh

In this investigation, the possibility of physical separation methods such as attrition scrubbing, hydro-cycloning and also reverse flotation were investigated for silica and calcite removal from relatively low grade kaolin. The head sample was provided by the Iran China Clay Company and the major oxides contained in the head sample were Al2O3, SiO2, CaO and Fe2O3 with grades of 15.07%, 71.05%, 4.35% and 0.52% respectively. The liberation degree of calcite and iron impurities from kaolin particles were over 80% for particles finer than 210µm and 53µm respectively. Attrition scrubbing tests were carried out with scrubber cell of the conventional Denver flotation cell model D-12. Laboratory scale hydro cyclone was applied for the beneficiation of kaolin. Attrition scrubbing tests with control screen of 106µm did not achieve the desirable yield for purification of this kaolin sample, but hydro cyclone experiments resulted in better results. Hydro cyclone with 2.54 cm diameter used and the only variable was the cyclone inlet pressure ranging from 68.65kPa to 98.07kPa. The best inlet pressure was found to be 68.65kPa. Based on the optimized condition, Al2O3 and CaO amount in the hydro-cyclone overflow was 23% and 1.47% respectively. However weight recovery of the overflow in each cyclone experiment was less than 30% and multi stages of hydro-cyclones would have to be applied for higher weight recovery and better yield. All of the flotation tests were carried out in the conventional Denver flotation cell model D-12. Reverse floatation tests for calcite removal were carried out under alkali pH based on full factorial design of variables in 2 levels including one center point. The three parameters tested were collector dosage (gr/ton), pH and depressant dosage (gr/ton). The collected data from flotation tests were analyzed and optimized collector and depressant dosage were 1000 gr/ton and optimized pH was 8.5. Under the above mentioned conditions, calcite removal was about 75% and CaO content in the final product was reduced from 4.35% to 1.33%. However calcite removal will increase to more than 83% if the feed particle size in flotation tests is reduced from 210µm to finer than 53µm. Finally after optimization, the best working condition and possible flow sheet were developed.

Jan 1, 2014

By John Starkey

"The development of the SAG Power Index (SPI) Test in Canada has created a new and accurate way to measure ore hardness in a SAG environment. Coupled with the standard Bond Ball Mill Work Index Test it is now possible to determine from these two relatively inexpensive laboratory tests, the amount of energy required to grind an ore from the primary crushed size to the desired liberation size. By testing multiple samples to discover where the hard and soft ore zones are located, the representative hardness distribution for the ore body can be determined and grinding equipment selected that is capable to meet production targets for all mining areas of the deposit.Previously, it was deemed prudent to select up to 50 t samples of ore for Semi Autogenous Grinding (SAG) pilot plant tests and a second 5 t sample for metallurgical pilot plant work. The methods outlined in this paper show how the 50 t sample is no longer required for SAG design because it is better to measure the representative hardness distribution of the ore body than the hardness of a “chosen” representative sample. This is demonstrated to be true by the fact that the lower horizons of an ore zone cannot initially be accessed to recover 150mm pieces, and the bulk samples taken, rarely represent the desired design hardness of the ore. Median hardness will not yield a good result in most cases and a clear understanding of the location and extent of the hardest zones is imperative to achieve a truly successful grinding circuit design.This paper discusses two plants (Kubaka and Agnico Eagle) that have been designed using SPI and Bond with no pilot plant test backup. Both have been very successful in that they can grind design tonnage on the hardest ores in the mine. This guarantees that the mine’s profitability is protected in any mining sequence. It also allows the extra power to be converted to higher throughput on softer ores if additional cash flow is needed. The advantages from the owner’s point of view are dramatic. This procedure will be attractive as well when considering the costs for test work, and the samples for testing. Existing diamond drill core is normally used and this allows the inclusion of data from the deepest regions of the mine. In addition, it is also clear that the decision to use a crusher can now be made at the design stage of a project based on measured parameters and the estimated capital costs of alternative designs."

Jan 1, 2003

By Lev O. Filippov, Inna V. Filippova

"Flotation of metal-organic precipitate in column is sensible to the hydrodynamic conditions determined by the diameter and dispersion of rising bubbles. Thus, the dissipation energy in the system resulting of ratio between superficial fluid velocities (gas and feed) in the column is main parameter for precipitate flotation in pilot or industrial scales. The knowledge of aggregate structure can lead to better understanding and control of many solid-liquid separation processes (sedimentation, filtration, thickening and flotation). Many authors highlight the importance of structural properties (i.e. floc density) for different industrial applications.This approach was applied to achieve the flotation of molybdenum precipitate obtained with organic collector a- benzoin oxime for molar ratio 1:2. Flotation tests were carried out at lab and pilot scale with rigorous control of the size distribution of bubbles and the values of dissipation energy in the systems.Structural evolution of Mo precipitate and the destruction phenomenon were described by using of the various fractal dimensions, which were deduced from experimental data of laser sizer devices and optic microscopy pictures of precipitate. The boundary fractal dimensions Df of large aggregates allows the detection of a structural scaled organisation of the precipitate flocs (elementary germs-microflocs-flocs). The fractal dimension of the particle mass distribution Dm=1.94+/-0.15 suggests, for the floc structure, a cluster-cluster aggregation process either with structuration or breakage phases.The destruction process characterised by a fractal dimension Dg controls the final size distribution. The floc destruction occurs for the lowest values of the dissipation energy (0.02 – 0.05 W/kg). The relation between Dg and the system dissipation energy Et (for various energy E dissipated in the agitation cell) involves the different breakage mechanisms of aggregates in function of the aggregation bonds (Fig. 1).Besides, the destruction of floc produces very fine particles (elementary germs or/and microfloc according to E), and thus induces the difficulties of bubble-particle attachment during the precipitate flotation in column. Yet, very small bubbles, which increase the collision probability, would make the flotation feed rate decrease. The optima ratio between superficial gas and feed velocities are determined for bubble size and flotation results control at pilot scale. Thus, it is possible to providing the catastrophically changes of precipitate column flotation by using the following established experimental relation: the mean diameter d of the flocs under the action of dissipated energy Et follows a power like law (d~Et -0.5). This allows estimating the size distribution of flocs for the E values which can appear in column."

Jan 1, 2003

By Hiroshi Sasaki, Herman D. Mendoza

"The surface-chemical property of nickel slag fiber was modified in aqueous media in order to act as an active and efficient collector of colloidal amorphous silica from geothermal effluent. Colloidal amorphous silica (0.05 µm) from geothermal wastewater causes operational and ecological problems. It clogs geothermal pipes and waterways, and when disposed to the environment poses hazardous effects on both the flora and fauna ecology. Because colloidal amorphous silica is extremely fine and remains in suspension in aqueous media, conventional solid/liquid processes to remove colloidal silica from its bulk liquid are rendered ineffective.A rapid separation technique is then developed to remove effectively amorphous colloidal silica from its bulk liquid by attaching the extremely fine particles of silica onto a nickel slag fiber coated with magnesium hydroxide, Mg(OH)2. The nickel slag fiber has the following characteristics: a high surface area and predominantly possesses negative charge in aqueous media with an iso-electric point at pH ~2.5. When soaked and saturated with Mg(OH)2 solution, the surface of the nickel slag fiber becomes coated with fine and compacted Mg(OH)2 precipitates. This modifies the surface- chemical property of nickel slag fiber transforming the negatively charged surface into a positively charged surface in aqueous media with an iso-electric point of around pH 11.8, which corresponds to the iso-electric point of Mg(OH)2. When geothermal effluent containing negatively charged colloidal amorphous silica particles (i.e.p. of silica is pH 2) is passed through a column-bed packed with surface-chemically modified nickel slag fibers under natural pH conditions, the colloidal amorphous silica particles are recovered by attachment onto the surface of the surface- modified nickel slag fiber. The recovery of colloidal amorphous silica by attachment is 97-99%, resulting in the production of a relatively clean and silica free output. Aluminum hydroxide, Al(OH)3, was also used to modify the surface –chemical property of the nickel slag fiber. In the same way, it also exhibited a positive response with the collection of colloidal amorphous silica from its bulk liquid, but yielding a lower recovery of 90%. This may be attributed to the lower surface potential and iso-electric point of pH 9.3 for Al(OH)3. The separation process was evaluated and analysed using existing potential energy theories. Figure 1 shows a micrograph of the surface of an unmodified and the Al(OH)3 –modified surface of nickel slag fiber."

Jan 1, 2003

By H. Breuer

Mineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sinteringMineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sintering

Sep 1, 2012

By Chenyang Zhang, Zhigang Yin, Wei Sun, Jingxiang Zou, Jianyong He, Yuehua Hu, Zhijie Xu

"In this work, a novel compound 4-Amino-5-mercapto-1,2,4-triazole was first synthesized and its selective adsorption mechanism on the surface of chalcopyrite was comprehensively investigated using UV-vis spectra, zeta potential, Fourier Transform Infrared Spectroscopy(FTIR), X-ray Photoelectron Spectrscpy Measurements(XPS), and Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS) and first principle calculations. The experimental and computational results consistently demonstrated that AMT would chemisorb onto the chalcopyrite surface by the formation of a five-membered chelate ring. The first principle periodic calculations further indicated that AMT would prefer to adsorb onto Cu rather than Fe due to the more negative adsorption energy of AMT on Cu on the chalcopyrite (001) surface, which was further confirmed by the coordination bonding energies of AMT-Cu and AMT-Fe based on the simplified cluster models at a higher accuracy level (UB3LYP/Def2TZVP). The bench scale results indicated that the selective index improved significantly when using AMT as a potential chalcopyrite depressant in Cu-Mo flotation separation.INTRODUCTIONFroth flotation is one of the most important processes to separate valuable minerals from gangue by taking advantage of differences in the physical chemistry of their surface properties(Laitinen et al., 2014), particularly in their surface hydrophobicity. Hydrophobicity differences between valuable and gangue minerals could be regulated by adding different chemical additives (surfactants)(Kor, Korczyk, Addai-Mensah, Krasowska, & Beattie, 2014), such as collector and depressant. It could be conducted economically to selectively separate and enrich the complicated and/or poor ores, by adopting appropriate surfactants. For example, the beneficiation of molybdenite usually relies on a two-stage process to realize flotation separation of molybdenite from other unwanted sulphide ores(Wie & Fuerstenau, 1974). In the first step, a bulk molybdenite concentrate is produced by utilizing kerosene/ diesel oil and pine oil/alcohols as collector and frother, respectively. Second, sodium hydrosulphide, sodium cyanide, ammonium sulphide, or Nokes reagent are adopted as depressants in the separation stage. However, these depressants are usually toxicant and cannot meet the requirements of environmental protection. A large number of efforts have been taken to develop alternative environmentally-friendly depressants (Ansari & Pawlik, 2007; J. H. Chen, Lan, & Liao, 2013; Li et al., 2015; Yin, Sun, Hu, Zhai, & Guan, 2017; Yin, Sun, Hu, Zhang, Guan, Liu, et al., 2017; Yin et al., 2016). Although some novel and environmental-friendly depressants have been reported to achieve good performance in Cu-Mo flotation separation at a laboratory scale, there are still some challenging problems, including toxicity, high price or poor performance in practice, which severely limit their practical applications in Cu-Mo flotation separation on a commercial scale."

Jan 1, 2018

By A. Mainza, H. Hill, L. Bbosa

"The main objective of ore breakage characterization is to quantify the relationship between the energy supplied for breakage and the size of the resultant progeny. In this study, the ore breakage characteristics of drill core particles and crushed particles were compared. Single impact breakage tests were conducted on a basalt ore sample using the JK rotary breakage tester. The size-dependent model by Shi and Kojovic was used to fit the breakage data because the effect of initial particle size is accounted for in the model. It was found that crushed particles consistently gave a higher degree of brokenness than drill core particles and the differences were attributed to the internal structure of the particle. Drill core particles contain a lower crack density (paths of weakness) than crushed particles which have been pre-weakened during the crushing process. An increase in crack density is directly related to a decrease in the minimum threshold energy. Upon application of a stressing force above the stress threshold, particles with a higher crack density will fracture more easily because of a higher rate of crack propagation due the numerous paths of weakness available. It was therefore concluded that under the same conditions, drill core particles are more competent than crushed particles. INTRODUCTION Comminution devices have commonly been reported as being the largest energy consumers in mineral processing circuits (Napier-Munn, 2015). Cleary (1998) reported that the inefficiency of comminution devices is a result of less than 5% of the supplied energy being utilized for breakage, with the majority being expended as heat and noise. With steadily declining fossil fuel supplies and a looming global energy crisis, the mineral processing industry has increased its focus on energy reduction measures. Ore breakage characterization plays a critical role in the design and optimization of comminution circuits (Tavares and King, 1998). The main objective of ore breakage characterization is to empirically quantify the relationship between the energy supplied for breakage (specific input energy) and the size of the resultant progeny through laboratory particle breakage tests (Napier-Munn, 1996). Commonly used laboratory breakage devices include the Drop Weight Tester (DWT), Split Hopkinson Pressure Bar (SHPB) and more recently, the JK Rotary Breakage Tester (JKRBT). Typically, a standard single impact breakage test procedure consists of a matrix of experiments on particles in four to five narrow size fractions each tested at low, medium and high energy levels. The resultant progeny of each test is sized to obtain a particle size distribution (PSD) from which the degree of breakage, usually in the form of a t10 value (percentage of the progeny material that passes 1/10 of the feed size) can be obtained for each size fraction and energy level (Napier-Munn, 1996)."

Jan 1, 2018

By Aline Tavares

This paper describes the investigative studies carried out at Vale´s Center of Mineral Development (CDM) for a nickel sulfide deposit aiming to develop an adequate flotation flowsheet. The CDM facilities comprises of very well equipped laboratories to develop hydrometallurgical, comminution, flotation, physical separation processes, besides chemical and mineralogical laboratories, that are required to accomplish pre-feasibility studies, new process routes developments and, eventually, brownfield projects. Close association of the flotation group with the geometallurgy and the mineralogy teams play a key role during greenfield project development. Most of the information required by process engineers responsible for the flowsheet development can be provided by the mineralogy team and its mineralogical laboratory at CDM. During the flowsheet development of this nickel sulfide deposit, two ore types were identified as the most abundant in the deposit, named biotite-chlorite hydrothermalite (HID) and magnetite hydrothermalite (MAG HID). Six samples were selected for the investigative studies based on this initial information. Nickel grade varied from 0.39% to 1.95%. The main nickel-bearing minerals were millerite and pentlandite whereas the main non-sulphide gangue minerals were quartz, plagioclase, biotite, chlorite, actinolite, and/or magnetite. Electron microprobe (EMP) identified the presence of nickel in solid solution in gangue minerals such as biotite, chlorite, and actinolite. Nickel deportment in silicates varied from 3% to almost 40%. Magnesium was distributed in biotite and chlorite, except in sample MAG HID, in which talc was responsible for almost 50% of the Mg content. Liberation studies indicated that the optimum feed grind size was around 150µm and that pyrite occurred highly associated to the nickel sulfides. The flotation tests were planned based on the information provided by the mineralogical data. Rougher flotation tests indicated that the optimum P80 of the primary grind was 212µm for the HID type and 150µm for the MAG HID. From the reagents tested, amyl xanthate proved to be the most efficient in the recovery of the sulfides. In the MAG HID sample talc was well depressed by dosing 600g/t of CMC. Nickel recoveries in the rougher stage were lower than 90% due to the nickel distributed in silicates. However, sulfur recovery was around 95%, indicating high recovery of the total sulfides. Open circuit and locked cycle tests were conducted to evaluate the effect of the regrind size in the nickel grade and recovery (P80 of 44µm and 20µm). Nickel recovery decreased from 70% to 65% when finer regrind size was tested. Although nickel recovery decreased by 5%, the finer regrind size lead to a higher Ni grade (from 11% to 23%Ni) and lower fluorine grade (from 1138ppm to 700ppm) in the concentrate. Keywords: nickel sulfide, flotation, mineralogical characterization, flowsheet development

Sep 1, 2012

By Y. Y. Li, T. T. Yang, R. X. Zhou, Z. H. Shen

"Wetting theory is one of the basic theories of flotation. The effects of concentrations, carbon chain length and pH values of saturated fatty acid collectors (lauric acid, myristic acid, palmitic acid and stearic acid) on the surface wettability of collophanite and dolomite were investigated by measuring the contact angle with dynamic captive-bubble technique and calculating adhesion work of bubble-mineral. The results show that the contact angles both of collophanite and dolomite are higher with enhancement of hydrophobicity when pH value of 5~9. The contact angles both of collophanite and dolomite are reduced with improvement of surface hydrophilicity. When strong acid pH value of 2~4 or strong alkaline pH value of 10~13. The reasons for this are that the fatty acids mainly exist with insoluble RCOOH molecules in strong acid condition, and competitive adsorption of OH- and RCOO- in the mineral surface is formed in strong alkaline condition. With the increase of the collector concentration, the contact angles of the collophanite and the dolomite all first increase then decrease. The formation of hydrophilic micelles results in the decreasing of contact angles and enhancement of surface hydrophilicity opposite when the collector concentration exceeds CMC. The contact angles of the collophanite and dolomite and the WSG of the bubble adhesion increase, the hydrophobicity enhance with the increasing of carbon chains of saturated fatty acid from 12 to 18. The variation tendency for the adsorption quantity of collectors and floatability of collophanite and dolomite are in line with the change trend of contact angle basically. It shows that the wettability of solid-liquid-gas interface can be well characterized by the dynamic captive-bubble technique.INTRODUCTIONPhosphorite resources are the main raw material for producing phosphate fertilizer and producing phosphoric acid (Cao et al.,2005), which is an important strategic mineral resource. The useful minerals in calcareous-magnesium phosphate ore are collophanite, and the main gangue minerals are dolomite (SISH et al.,2003). The lattice ions of the phosphorite and dolomite are both Ca2+, their similar surface properties and similar floatability are the main reason for the flotation of calcareous-magnesium phosphate ore difficultly (Fang et al.,2011). Usually, fatty acid collectors are used for the flotation of calcareous-magnesium phosphate ore under acidic conditions by reversing flotation method to remove dolomite (ABOUZEID et al.,2009), or direct flotation is used to gather collophanite in the alkaline conditions (Gong et al.,2017). The flotation behavior of collophanite and dolomite affects the indexes of concentrate and flotation efficiency directly. Therefore, it is necessary to study the action mechanism and the flotation method of the collector with the phosphate rock and dolomite respectively (Moudgil et al.,1988; Hanna, et al.,1990)."

Jan 1, 2018