Beneficiation of Iron Ore and the treatment of magnetic iron taconites, stage grinding and wet magnetic separation is standard practice. This also applies to iron ores of the non-magnetic type which after a reducing roast are amenable to magnetic separation. All such plants are large tonnage operations treating up to 50,000 tons per day and ultimately requiring grinding as fine as minus 500-mesh for liberation of the iron minerals from the siliceous gangue.
Magnetic separation methods are very efficient in making high recovery of the iron minerals, but production of iron concentrates with less than 8 to 10% silica in the magnetic cleaning stages becomes inefficient. It is here that flotation has proven most efficient. Wet magnetic finishers producing 63 to 64% Fe concentrates at 50-55% solids can go directly to the flotation section for silica removal down to 4 to 6% or even less. Low water requirements and positive silica removal with low iron losses makes flotation particularly attractive. Multistage cleaning steps generally are not necessary. Often roughing off the silica froth without further cleaning is adequate.
The iron ore beneficiation flowsheet presented is typical of the large tonnage magnetic taconite operations. Multi-parallel circuits are necessary, but for purposes of illustration and description a single circuit is shown and described.
The primary rod mill discharge at about minus 10- mesh is treated over wet magnetic cobbers where, on average magnetic taconite ore, about 1/3of the total tonnage is rejected as a non-magnetic tailing requiring no further treatment. The magnetic product removed by the cobbers may go direct to the ball mill or alternately may be pumped through a cyclone classifier. Cyclone underflows usually all plus 100 or 150 mesh, goes to the ball mill for further grinding. The mill discharge passes through a wet magnetic separator for further upgrading and also rejection of additional non-magnetic tailing. The ball mill and magnetic cleaner and cyclone all in closed circuit produce an iron enriched magnetic product 85 to 90% minus 325 mesh which is usually the case on finely disseminated taconites.
The finely ground enriched product from the initial stages of grinding and magnetic separation passes to a hydroclassifier to eliminate the large volume of water in the overflow. Some finely divided silica slime is also eliminated in this circuit. The hydroclassifier underflow is generally subjected to at least 3 stages of magnetic separation for further upgrading and production of additional final non-magnetic tailing. Magnetic concentrate at this point will usually contain 63 to 64% iron with 8 to 10% silica. Further silica removal at this point by magnetic separation becomes rather inefficient due to low magnetic separator capacity and their inability to reject middling particles.
The iron concentrate as it comes off the magnetic finishers is well flocculated due to magnetic action and usually contains 50-55% solids. This is ideal dilution for conditioning ahead of flotation. For best results it is necessary to pass the pulp through a demagnetizing coil to disperse the magnetic floes and thus render the pulp more amenable to flotation.
Feed to flotation for silica removal is diluted with fresh clean water to 35 to 40% solids. Being able to effectively float the silica and iron silicates at this relatively high solid content makes flotation particularly attractive.
For this separation Sub-A Flotation Machines of the open or free-flow type for rougher flotation are particularly desirable. Intense aeration of the deflocculated and dispersed pulp is necessary for removal of the finely divided silica and iron silicates in the froth product. A 6-cell No. 24 Free-FlowFlotation Machine will effectively treat 35 to 40 LTPH of iron concentrates down to the desired limit, usually 4 to 6% SiO2. Loss of iron in the froth is low. The rough froth may be cleaned and reflotated or reground and reprocessed if necessary.
A cationic reagent is usually all that is necessary to effectively activate and float the silica from the iron. Since no prior reagents have come in contact with thethoroughly washed and relatively slime free magnetic iron concentrates, the cationic reagent is fast acting and in somecases no prior conditioning ahead of the flotation cells is necessary.
A frother such as Methyl Isobutyl Carbinol or Heptinol is usually necessary to give a good froth condition in the flotation circuit. In some cases a dispersant such as Corn Products gum (sometimes causticized) is also helpful in depressing the iron. Typical requirements may be as follows:
One operation is presently using Aerosurf MG-98 Amine at the rate of .06 lbs/ton and 0.05 lbs/ton of MIBC (methyl isobutyl carbinol). Total reagent cost in this case is approximately 5 cents per ton of flotation product.
The high grade iron product, low in silica, discharging from the flotation circuit is remagnetized, thickened and filtered in the conventional manner with a disc filter down to 8 to 10% moisture prior to treatment in the pelletizing plant. Both the thickener and filter must be heavy duty units. Generally, in the large tonnage concentrators the thickener underflow at 70 to 72% solids is stored in large Turbine Type Agitators. Tanks up to 50 ft. in diameter x 40 ft. deep with 12 ft. diameter propellers are used to keep the pulp uniform. Such large units require on the order of 100 to 125 HP for thorough mixing the high solids ahead of filtration.
In addition to effective removal of silica with low water requirements flotation is a low cost separation, power-wise and also reagent wise. Maintenance is low since the finely divided magnetic taconite concentrate has proven to be rather non-abrasive. Even after a years operation very little wear is noticed on propellers and impellers.
A further advantage offered by flotation is the possibility of initially grinding coarser and producing a middling in the flotation section for retreatment. In place of initially grinding 85 to 90% minus 325, the grind if coarsened to 80-85% minus 325-mesh will result in greater initial tonnage treated per mill section. Considerable advantage is to be gained by this approach.
Free-Flow Sub-A Flotation is a solution to the effective removal of silica from magnetic taconite concentrates. Present plants are using this method to advantage and future installations will resort more and more to production of low silica iron concentrate for conversion into pellets.
All available copper-bearing natural mineral aggregates are called copper mines. The high-grade copper concentrate can be obtained by the coarse grinding, roughing, scavenging of copper ore, then grinding and concentrating of coarse concentrate.
Due to the different types of ore, the nature of the ore is also different, so the beneficiation process needs to be customized. The specific process for selecting copper ore depends mainly on the material composition, structure and copper occurrence state of the original copper ore.
Before the beneficiation of copper ores, crushing and grinding are required. The bulk ores are crushed to about 12cm by a jaw crusher or a cone crusher. Then the crushed materials are sent to the grinding equipment, and the final particle size of the copper ore is reduced to 0.15-0.2mm.
Copper sulfide can be divided into single copper ore, copper sulfur ore, copper-molybdenum deposit, copper nickel, carrollite and so on. Basically, only flotation can be considered in its separation.
Almost all copper sulphide ores contain iron-bearing sulfides, so in a sense, the flotation of copper sulfide is essentially the separation of copper sulfide from iron sulfide. The common iron sulfide minerals in copper ore are pyrite and pyrrhotite.
1 Disseminated grain size and symbiotic relationship of copper and iron sulfide. Generally, pyrite has a coarse grain size, while copper ore, especially secondary copper sulfide, is closely associated with pyrite. Only when the copper ore is finely ground can it be dissociated from pyrite. This characteristic can be used to select copper-sulphur mixed concentrates, discard the tailings, and then grind and separate the mixed concentrate.
2 The influence of secondary copper sulfide minerals. When the secondary copper sulfide mineral content is high, the copper ions in the slurry will increase, which will activate the pyrite and increase the difficulty of Cu-S separation.
3 The influence of pyrrhotite. The high content of pyrrhotite will affect the flotation of copper sulfide. Pyrrhotite oxidation will consume the consumption of oxygen in the pulp. In severe cases, the copper minerals do not float at the beginning of flotation. This can be improved by increasing inflation.
Generally, copper is floated firstly and then sulfur. The content of pyrite in dense massive copper-bearing pyrite is quite high and high alkalinity (free CaO content> 600800g/m3) and high dosage of xanthine are often used to suppress the pyrite. There is mainly pyrite in its tailings with few gangues, so the tailings are sulfur concentrates.
For the disseminated copper-sulfur ore, the preferential flotation process is adopted, and the sulphur in the tailings must be re-floated. To reduce the consumption of sulfuric acid during the floatation and ensure safe operation, the process condition of low alkalinity should be adopted as far as possible.
It is more advantageous for copper sulfur ore containing less sulfur with copper easy to be floated. Carry out the bulk flotation firstly in the weakly alkaline pulp and then add lime to the mixed concentrate to separate the copper and sulfur in the highly alkaline pulp.
In semi-preferential bulk-separation flotation, Z-200, OSN-43 or ester-105 with good selectivity are used as collectors to float copper minerals firstly. The copper concentrate is then subjected to copper-sulfur bulk flotation and the obtained copper-sulfur mixed concentrate is subjected to separation flotation of floating copper and suppressing sulfur.
It avoids the inhibition of the easily floating copper under high lime consumption and does not require a large amount of sulfuric acid-activated pyrite. It has the characteristics of reasonable structure, stable operation, a good index and early recovery of target minerals.
3 The xanthate collector mainly plays the role of chemisorption together with the cation Cu (2 +), so minerals whose surface contains more Cu (2 +) minerals have a strong effect with the xanthate. The order of the effect is: chalcocite > covellite > porphyrite> chalcopyrite.
4 The floatability of copper sulfide minerals is also affected by factors such as crystal size, mosaic size, being original or secondary. The minerals with fine crystal and mosaic size are difficult to float. Secondary copper sulfide ore is easy to oxidize and more difficult to float than original copper ore.
As for the grinding and floating process, it is more advantageous to adopt the stage grinding and floating process for refractory copper ore, such as the re-grinding and re-separation of coarse concentrate, re-grinding and re-separation of bulk concentrate, and separate treatment of medium ore.
Copper oxide (CuO) is insoluble in water, ethanol, soluble acid, ammonium chloride and potassium cyanide solutions. It can react with alkali when slowly dissolving in ammonia solution. The beneficiation methods of oxidized copper ore mainly include gravity separatio, magnetic separation (see details on copper ore processing plant), flotation and chemical beneficiation.
Flotation is one of the commonly used mineral processing techniques for copper oxide ores. According to the different properties of copper oxide ores, there are sulphidizing flotation, fatty acid flotation, amine flotation, emulsion flotation and chelating agent-neutral oil flotation method.
Process flow: The dosage of sodium sulfide can reach 1~2kg/t during vulcanization. Because the film produced by vulcanization is not stable and is easy to fall off after vigorous stirring, and sodium sulfide itself is easily oxidized, sodium sulfide should be added in batches.
Besides, the vulcanization speed of malachite and azurite is relatively fast, so the vulcanizing agent can be directly added to the first flotation cell with no need to stir in advance during vulcanization and adjust the amount of vulcanizing agent according to the foam state.
Fatty acids and their soaps are mainly used as collectors of fatty acid floatation, also known as direct flotation. During flotation, water glass (gangue inhibitor), phosphate, and sodium carbonate (slurry regulator) are also usually added.
There is a practice of mixing vulcanization and fatty acid methods. Firstly float the copper sulfide and part of the copper oxide with sodium sulfide and xanthate, and then float the residual copper oxide with fatty acid.
For example, the ore in the Nchanga processing plant in Zambia contains 4.7% copper. The copper content achieved to 50% ~ 55% through flotation by adding 500g/t of lime (pH 9 ~ 9.5), 10g/t of cresol (foaming agent), 60g/t of ethylxanthate, 35g/t of amyl xanthate, 1kg/t of sodium sulfide, 40g/t of palmitic acid and 75g/t of fuel oil.
It is mainly to sulfurize the copper oxide mineral firstly and then add the copper accessory ingredient to create a stable oil-wet surface. Then, the neutral oil emulsion is used to cover the mineral surface, resulting in a strong hydrophobic floating state. In this way, the mineral can be attached to the foams firmly to complete the separation.
Many problems should be paid attention to in the flotation of copper ore, such as the length of the vulcanization time, whether to add sodium sulphide in batches and the proportion of chemicals. Here is a brief introduction.
1 The vulcanization time. Different ores require different vulcanization times. Generally speaking, it should be short rather than longer. The suitable vulcanization time is 1 to 3 minutes. After 6 minutes, the recovery rate and concentrate grade will decrease.
2 Add sodium sulfide in batches. The roughing time for processing the ore in the concentrator is about ten minutes, while the ore contains a large amount of carbonaceous gangue and the divalent sulfur ions disappear quickly in the slurry. So the effect of adding sodium sulfide in batches is better than that of adding it once.
3 Add sodium sulfide proportionally. Generally, copper oxide floats in the liquid at a slower speed, and reduce the number of cycles of the mineral in the flotation process can obtain a higher recovery rate. It is of great significance to study the distribution ratio of sodium sulfide among different operations to catch the mineral at the right time.
The chemical beneficiation method is often used for refractory copper oxide and mixed copper. For some copper oxide minerals with high copper content, fine mosaic size and rich sludge, the chemical beneficiation method will be used to obtain good indicators because the flotation method is difficult to realize the separation.
The solution of ammonia and ammonium carbonate in a concentration of 12.5% was used as the solvent to leach for 2.5h at a temperature of 150, a pressure of 1925175~2026500Pa. The mother liquor can be distilled by steam at 90 to separate ammonia and carbon dioxide. Copper, on the other hand, is precipitated from the solution as black copper oxide powder.
Because some copper oxide minerals are not tightly combined with iron, manganese, etc., it is difficult to separate them by using the magnetic separation method alone, and flotation has a good separation effect.
Therefore, the flotation method is used to obtain high-grade concentrates, the magnetic separation is for tailings and wet smelting is carried out finally. This process combines flotation, magnetic and wet smelting very well, which greatly increases the recovery rate and reduces the beneficiation cost.
The above are several common beneficiation methods for copper oxide minerals. For the selection of copper oxide minerals, it is best to conduct a professional beneficiation test and customize the process according to the report.
Flotation is the most widely used method in copper mine production. The copper ore pulp is stirred and aerated, and the ore particles adhere to the foams under the action of various flotation agents. The foams rise to form a mineralized foam layer, which is scraped or overflowed by the scraper. This series of flotation processes are all completed in the flotation machine. (Contact Manufacturer)
The internal magnetic system of the barrel adopts a short circuit design to ensure that the barrel skin has no magnetic resistance at high speeds, and the stainless-steel barrel skin does not generate high temperatures, extending the life of the magnetic block.
Since it adopts a dynamic magnetic system design, the roller does not stick to the material, which is conducive to material sorting. The selected grade can be increased by 3-6 times to more than 65%.
Copper mines are generally purified by flotation, but for the beneficiation of copper minerals with coarser grain size and higher density, the pre-selection by the gravity separation method will greatly reduce the cost and achieve flotation indicators.
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Wen-li Jiang, Hai-feng Xu, "Treatment and Recycling of the Process Water in Iron Ore Flotation of Yuanjiacun Iron Mine", Journal of Chemistry, vol. 2017, Article ID 9187436, 8 pages, 2017. https://doi.org/10.1155/2017/9187436
Coagulating sedimentation and oxidation treatment of process water in iron ore flotation of Yuanjiacun iron mine had been studied. The process water of this mine carried residual polyacrylamide (PAM), poly(diallyldimethylammonium chloride) (PDADMAC), and Ca2+ from the flotation and caused decrease of the iron flotation recovery or grade of the concentrate. The studies on high-intensity magnetic separation (HIMS) tailings for coagulating sedimentation showed that the settling performance of coagulant (named CYH) was better than that of PDADMAC. The analyses of FTIR spectra and zeta potential demonstrated that CYH is adsorbed mainly through electrostatic attraction onto HIMS tailings. Sodium hypochlorite was adopted to oxidize the residual organics in tailings wastewater. When sodium hypochlorite is at the dosage of 1.0g/L, reaction temperature is of 20C, and reaction time is of 30 minutes, the removal rates of PAM, COD, and Ca2+ were 90.48%, 83.97%, and 85.00%, respectively. Bench-scale flotation studies on the treated tailings wastewater indicated that the iron recovery and grade of concentrate were close to those of freshwater.
Iron ore resources are extremely rich in China, but most of them belong to complex ultrafine iron ore with high content of impurities . Reverse flotation has been proved to be an efficient process for economic reasons . In order to ensure iron concentrate grade and iron recovery, a large number of processing reagents are selected and applied in the iron ore beneficiation. The process water carried plenty of residual processing reagents, and such wastewater with color depth and strong smell could seriously affect the environment and the local people. Prevailing campaign for a cleaner and safer environment with clean surface water and ground water has led to increased recycling of process water within the production cycle of mineral flotation . Since the chemistry property of process water is entirely different from fresh water, there is a concern about the possible effects of the contained components on the efficiency of the flotation process . In iron ore flotation process, a large amount of NaOH is taken to adjust the pH. The wastewater pH of iron ore is more than 9, so physicochemical treatment of iron wastewater from a certain degree is rather difficult . At present, the common methods of treatment of wastewater from flotation are acid-alkali neutralization [10, 11], precipitation, coagulation, and sedimentation , chemical oxidation degradation , constructed wetland , ion exchange method , adsorption method [21, 22], and biological method [23, 24]. However, the single method above cannot completely clear wastewater pollution with harsh operating conditions. In addition, they may also produce secondary pollution.
Yuanjiacun iron mine of Shanxi province in China is the largest iron mine of micrograined hematite combined with magnetite in Asia. The annual processing capacity of iron ore reaches 22 million tons, and the wastewater from mineral processing is about 400 thousand m3 every day. If the dressing wastewater is directly discharged into the nature, it will not only pollute the surrounding environment but also has a great potential safety risk. Recycling of process water is a good way to solve the problem. In this study, the influence of tailings wastewater components were investigated on flotation of hematite in order to assess the practicability of recycling process water in flotation practice first. A novel coagulant CYH was introduced to coagulate HIMS tailings, and the coagulating sedimentation mechanism of CYH was investigated by FTIR spectra and zeta potential. Moreover, sodium hypochlorite was adopted to oxidize the organic pharmaceutical residues in wastewater. Besides, the effect of wastewater treatment was evaluated by recycling of tailings wastewater.
All test samples including tailings wastewater, high-intensity magnetic separation (HIMS) tailings, and flotation samples were procured from the Yuanjiacun iron mine concentrator of Shanxi province in China. The tailings wastewater was collected from flotation tailings water combined with magnetic separation tailings water, stirred evenly, and stored in plastic bottle. The wastewater quality was shown in Table 1. Test results showed that the tailings wastewater presented yellowish brown, suspended solids (SS) content and the concentration of Ca2+, Cl, and COD exceeded the standard.
The concentration of the HIMS tailings pulp was 11.93% by weight, and the main minerals were quartz (62.54%), amphibole (17.44%), chlorite (4.16%), hematite (4.75%), montmorillonite (3.82%), calcite (2.93%), and feldspar (4.36%). The XRD analysis of the sample was shown as Figure 1. The sample was classified into different size fractions as shown in Table 2. As shown in Table 2, most particle size of the sample was fine and more, even if the sample was on quiescent standing in a month, it also would not naturally subside.
The samples for flotation were taken from on-site samples of flotation, which were ground to 95% passing 0.045mm, and contained 34.90% specularite and hematite, 8.70% magnetite, 0.20% limonite, 39.40% quartz, 11.40% chlorite and hornblende, and 4.90% calcite and dolomite.
Coagulant (CYH) of the molecular weight of 80 thousands was synthesized in our lab and it was of technical grade. Amphoteric polyacrylamide (PAM) of the molecular weight of 12 million was of technical grade. Collector (named RA-715) and coagulant poly(diallyldimethylammonium chloride) (PDADMAC) were from the Yuanjiacun iron mine concentrator of Shanxi province in China and they were of all technical grade. CaCl2 and NaCl acted as sources of Ca2+ and Cl ions, respectively. Other chemical materials were bought from commercial companies and their purity was above chemical purity grade.
Use 300mL HIMS tailings pulp mentioned above at every turn to conduct coagulation contrast test. First, add the required amount of CYH into the testing pulp, and mix them up evenly; then add PAM into the testing pulp, and after a period of stirring, leave the samples standing and record the height of supernatant as a function of standing time.
Use 1.0L iron tailings wastewater mentioned above at every turn to conduct oxidation-sedimentation experiment. When the wastewater was heated to the desired temperature, sodium hypochlorite (10% of available chlorine) was added. After stirring the mixture for the desired time, FeCl2 were added to the mixture, and the KI-starch paper was used to detect the reaction progress until it did not change blue. Afterwards, the pH was adjusted to 9~10 with NaOH solution. After several hours of standing, the supernatant was separated, and then the PAM concentration, COD value, and Ca2+ content were measured.
The bench-scale flotation tests were conducted in a XFD-63 flotation cell (self-aeration) whose volume for rougher flotation and cleaning flotation was 0.5L, using 200g ore at every turn to obtain a Fe concentrate. Fatty acids (RA-715) were used as collector, NaOH was used as pH regulators, starch was used as depressant, and CaO was used as activator. The flotation flow sheet was illustrated in Figure 2.
The infrared spectra of samples were recorded by Nicolet AVATAR370 FTIR spectrometer (USA) using the KBr disk technique. The quartz samples used for this purpose were ground in an agate mortar and pestle to pass 5m. 50mg of samples was mixed with 30mL distilled water in the absence or presence of 100mg/L coagulant at pH 9 and 25C. After stirring for 3min, still standing for 4h, the solid product in the mixture was filtrated and rinsed three times and then dried in a vacuum oven and recorded infrared adsorption spectra from 400cm1 to 4000cm1.
Zeta potentials of HIMS tailings samples were measured by using a Brookhaven ZetaPlus zeta potential analyzer (USA). The samples used for this purpose were ground to less than 5m in an agate mortar and pestle. 50mg of the samples was added to 30mL aqueous solution with or without 100mg/L coagulant. After stirring for 10min, then the pH values were adjusted with HNO3 or NaOH solutions and measured. All measurements were conducted in a 0.1mol/L KNO3 background electrolyte solution. The agitated suspension was sampled to record the zeta potential. The results presented were the average of five independent measurements with a typical variation of 5mV.
Bench-scale flotation studies on the process water showed that tailings wastewater reduced the flotation iron concentrate grade and iron recovery . Large doses of coagulants such as PAM and PDADMAC were used in wastewater treatment in Yuanjiacun concentrator previously. Only in the part of HIMS tailings concentration, doses of PAM and PDADMAC were several times of other similar mines . In order to assess the practicability of recycling of process water in flotation practice, the influence of tailings wastewater components of PAM, PDADMAC, Ca2+, and Cl ions was investigated independently. The results were shown in Figure 3.
Figures 3(a) and 3(b) showed that an increasing PAM or PDADMAC concentration reduced the iron flotation recovery. However there was very little reduction in iron grade. When the concentration increased from 0mg/L to 3mg/L for PAM and 0mg/L to 30mg/L for PDADMAC, the recoveries of iron decreased from 74.66% to 66.74% and 74.66% to 64.63%, respectively. Figure 3(c) demonstrated that an increasing Ca2+ ions concentration reduced the iron grade and increased the iron flotation recovery obviously. When the concentration of Ca2+ ions increased from 0mg/L to 180mg/L, the iron grade of concentration decreased from 66.27% to 64.42%; the iron grade decreased sharply on addition of 360mg/L. However there was a slight increase in recovery in process water in presence of a number of Ca2+ ions. Figure 3(d) showed that, with increasing the dose of Cl, there was no change in iron grade and recovery decreased only slightly.
Single factor tests of components contained in tailings wastewater showed that PAM, PDADMAC, and Ca2+ ions reduced the flotation iron recovery or grade of the concentrate. The origins of PAM, PDADMAC, and other organic species in tailings wastewater were the coagulants, flocculants, and flotation reagents such as PAM and PDADMAC for coagulating sedimentation of concentration and wastewater treatment, starch for depressing hematite flotation, and RA-715 for flotation of activated silicate minerals. The origin of calcium species in tailings wastewater was the ore and flotation reagents such as lime for activating SiO2 and silicate minerals. The method of coagulating sedimentation was used in the process of concentration of HIMS tailings, flotation concentrate and flotation tailings, and treatment of wastewater of tailings pond in Yuanjiacun concentrator. In order to lower the content of suspended solids, PAM, and PDADMAC, coagulating sedimentation experiments were conducted with a novel coagulant CYH. Given the maximum amount of coagulants and flocculants used in HIMS tailings concentration in Yuanjiacun concentrator at present, we chose coagulating sedimentation testes of HIMS tailings concentration as a representative for detailed investigations to study the performance of CHY.
Figure 4 exhibited the effect of PDADMAC or CYH dosage on coagulation efficiency by using PAM as a flocculant at 26.88g/(t undressed ore, the same below) initial concentration. The results in Figure 4 showed that, at the coagulant dosage of 8.96g/t, the settle rate of CYH was faster than that of PDADMAC, even faster than that of PDADMAC at the dosage of 17.92g/t. The turbidity of liquid supernatant by using CYH as a coagulant at 8.96g/t initial concentration was lower than that of PDADMAC same as that of initial concentration and was rough equal to that of PDADMAC as a coagulant at 17.92g/t initial concentration. When the initial concentration of CYH increased from 8.96g/t to 17.92g/t, the settle rate and turbidity changed little. Compared with PDADMAC, CYH exhibited superior coagulating ability.
The flocculation response of PAM as a function of initial concentration by using CYH as a coagulant at 8.96g/t initial concentration was presented in Figure 5. As it could be observed from Figure 5, the settle rate rapidly increased with increasing flocculant concentration. However, increasing PAM concentration had minor influence on turbidity of liquid supernatant.
As shown in Figure 1, quartz was the highest content of mineral in HIMS tailings. So, we chose quartz as a representative for FTIR spectrum investigation to study the adsorption mechanism of quartz before and after interaction with CYH. The FTIR spectra were presented in Figure 6.
Figure 6 showed that, after interaction with CYH, the stretching and bending vibrations of saturated C-H bonds in CYH molecules appeared at around 2962.17, 2933.24, 2871.53, and 1432.15cm1 on quartz surfaces, respectively. The results of FTIR spectra exhibited that, after CYH treatment, new adsorption peaks on quartz surfaces did not appear except for CYHs adsorption bands, which inferred that CYH might adsorb onto quartz surface without the formation of new complexes.
Zeta potentials of HIMS tailings particles as a function of pH values in the absence and presence of CYH were shown in Figure 7. It indicated that the potential of HIMS tailings was high. So coagulating sedimentation treatment of HIMS tailings was rather difficult from a certain degree because of the electrostatic repulse force among particles. As it could be observed from Figure 7, CYH could lower -potential values within the scope of pH 5~12, inferring that cationic CYH might adsorb onto particles surfaces. The results of zeta potential indicated that CYH adsorbed onto HIMS tailings mainly through electrostatic attraction, which agreed with the FTIR spectra results.
CYH that is a good coagulant has strong binding force by using hydroxy. After the mixture of HIMS tailings and CYH, the stable silicate mineral groups in the wastewater were exposed; meanwhile, CYH entered into the wastewater and hydrolyzed strongly to be a polyhydroxy polymer compound. CYH adsorbed characteristically onto particles surfaces through electrostatic force, hydrogen bond, hydrophobic association, and van der Waals force, bridging the various silicate minerals in the long CYH chain, forming large flocs to reach rapid subsidence through trapping and rolling.
Table 3 showed that, by using CYH as a coagulant, the content of SS of tailings wastewater decreased obviously, and the wastewater became clear. However, the COD value and concentration of PAM in tailings water were still much higher than freshwater. The tailings water still could not meet the requirements of reuse for flotation. Thus, oxidation-sedimentation treatment was still needed in the process.
We chose removal rates of PAM and COD as representatives for detail investigations to study the effect of reaction parameters. Figure 8(a) showed the effect of dosage of sodium hypochlorite (NaClO) on the removal rates of PAM and COD. As shown in Figure 8(a), the removal rates of PAM and COD rapidly increased with increasing NaClO concentration when it was less than 1.0g/L and then slowly increased and reached 90.48% and 83.97% at 1.0g/L dosage, respectively. As it could be observed from Figure 8(b), the removal rates of PAM and COD increased with increasing reaction temperature. When the reaction temperature increased from 15C to 30C, the removal rates increased from 87.62% to 94.61% and 77.86% to 87.64% for PAM and COD, respectively. The reaction temperature is higher, the reaction rate is faster, but the decomposition of NaClO is faster. The choice of 20C as the reaction temperature can ensure NaClO has good oxidation ability, but also conform to the most of natural temperature of tailings wastewater in Yuanjiacun concentrator. Figure 8(c) demonstrated that the removal rates of PAM and COD rapidly increased with prolonging reaction time when it was less than 30min and then maintained roughly constant.
As shown in Table 4, after treatment by using oxidation-sedimentation method, pH of tailings wastewater increased from 9.12 to 9.35, suspended solids content decreased from 178mg/L to 48mg/L, and concentration of Ca2+ ions, TFe, PAM, and COD decreased from 240mg/L to 36mg/L, 4.97mg/L to 0.36mg/L, 3.15mg/L to 0.30mg/L, and 131mg/L to 21mg/L, respectively. And the removal rates of PAM, COD, and Ca2+ was 90.48%, 83.97%, and 85.00%, respectively. The water quality of treated tailings wastewater was close to the quality of freshwater.
Bench-scale flotation studies were conducted according to the procedure as shown in Figure 2. These studies would show whether oxidation-sedimentation treatment of the tailings wastewater occurring in the pilot plant had any effect on selectivity and/or recovery of minerals during flotation. Results were given in Table 5. The results showed that the concentrate yield, grade, and recovery increased by 0.95%, 1.25%, and 2.75%, respectively, by using treated tailings wastewater as flotation water compared to that of tailings wastewater without treatment. Compared to freshwater, in the case of the equivalent yield, treated tailings wastewater achieved an excellent concentration containing 65.45% Fe with 73.49% Fe recovery, and the Fe grade increased by 0.31%.
(1) Single factor tests of components contained in tailings wastewater in Yuanjiacun concentrator showed that PAM, PDADMAC, and Ca2+ ions reduced the flotation iron recovery or grade of the concentrate.
(2) When CYH was used to coagulate HIMS tailings, -potential decreased, and silicate minerals formed flocculent mass by bridging; thus suspended matters decreased effectively. This was subsequently confirmed by FTIR spectrum and zeta potential analysis. But the tailings wastewater could not reach recycling standards.
(3) Oxidation experiment showed that a 90.48% reduction in PAM, 83.97% reduction in COD, and 85.00% reduction in Ca2+ were achieved at the sodium hypochlorite dosage of 1.0g/L, reaction temperature of 20C, and reaction time of 30 minutes. Bench-scale flotation tests on the treated tailings wastewater indicated that the Fe recovery and grade of concentrate were close to those of freshwater.
The authors gratefully acknowledge the financial supports of the National Natural Science Foundation of China (no. 5160041305) and the China Postdoctoral Science Foundation (no. 2016M591382). This project is also supported by Lanxian County Mine Co., Ltd., TISCO, China.
Copyright 2017 Wen-li Jiang and Hai-feng Xu. This is an open access article distributed under the Creative Commons Attribution License, which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.
Process water collected from the iron ore beneficiation plant was treated by electrocoagulation (EC) process to make it suitable for reuse or safe for discharge. Experimental studies were carried out towards efficient removal of the various metal ions such as iron (Fe), chromium (Cr), copper (Cu), zinc (Zn), manganese (Mn), lead (Pb), aluminum (Al) from process water along with monitoring of total dissolved solids (TDS), turbidity, conductivity, salinity. The influence of various operating parameters of EC, such as electrode material, electrode configuration, current density, inter-electrode distance, and solution conductivity, were explored for effective treatment of process water. Experimental results showed that aluminum electrodes at monopolar mode with a current density of 68.50A/m2, an inter-electrode distance of 1cm, and solution conductivity of 1033S/cm were the ideal operating conditions to get treated water with a removal efficiency of 99.95%, 99.46%, 99.33%, 97.99%, 73.44% for Fe, Cr, Pb, Mn, and Cu ions respectively after 60min of EC with electric energy consumption of 3.93kWh/m3 and operating cost of 0.6115 US$/m3. X Ray Diffraction analysis of EC generated sludge confirmed the exclusion of major elements like Fe, Cr, Cu, Zn, Mn, Pb, Al as well as trace elements such as Nickel, Cobalt, Magnesium, Titanium, Vanadium, Zinc, Neodymium, Samarium, Yttrium, etc. from process water.
Mineral processing comprises two principal steps: size reduction to liberate the grains of valuable mineral (or paymineral) from gangue minerals, and physical separation of the particles of valuable minerals from the gangue, to produce an enriched portion, or concentrate, containing most of the valuable minerals, and a discard, or tailing (tailings or tails), containing predominantly the gangue minerals.
Most mineral processing plants are represented by the flow sheet shown in Fig. 11. Simpler operations, such as a quarry producing aggregate, would involve only the initial stages of size reduction. Conversely, a more complex plant, producing a number of concentrates, requires a series of concentrating circuits.
Rod and ball mills have larger reduction ratios. Because rod mills provide some internal sizing action, they are commonly operated in open circuit. There is an increasing tendency to use an extra stage of crushing instead of a rod mill because crushers are more energy efficient.
Wet ball mills are usually operated in closed circuit to control the product size distribution. At these fine sizes, only classifiers have sufficiently high capacities. Because of their low cost, hydrocyclone classifiers are virtually always used. Unfortunately, they can be very inefficient; not only do they have the inherently spread performance curve of a sedimentation classifier, but at the high pulp Figure 11 Representation of mineral processing circuit. densities used, bypassing markedly lowers the efficiency. Two-stage classification is gradually being adopted, with recently developed polyurethane and sandwich screen surfaces being potential alternatives.
In new operations, semiautogenous grinding is now as common as the conventional size reduction shown in Fig. 11. In such plants, the crushing and grinding equipment shown in Fig. 11 is replaced by a single stage of (primary) crushing, followed by a SAG mill operating in closed circuit with hydrocyclone classifiers.
The purpose of the size reduction circuit is to achieve liberation to allow concentration. Because some ores contain massive mineralization or because the gangue mineral liberates at a coarser particle size, some plants employ preconcentration before the valuable is fully liberated. Heavy media separations are usually used between crushing stages, but a few instances of conventional gravity preconcentration after crushing have been reported.
Because reduced performance curves represent separator efficiency, these curves can be used to compare separators. In general, however, efficiency (as indicated by the steepness of the reduced performance curve) and throughput decrease with decreasing particle size. Although each separator tends to have its own optimum particle size range, comparisons between separators must be made at the same particle size.
With finer particle sizes (particularly those encountered in flotation), efficiencies become so low that single stage treatment is inadequate. Instead, roughing, cleaning, and scavenging stages are needed (Fig. 11).
Roughers aim to recover liberated valuable and reject middlings and liberated gangue to the scavengers, which recover a concentrate containing the middlings, and a tailings of liberated waste. This scavenger concentrate is then subjected to regrinding to increase liberation. Traditional practice was to return the ground material to the rougher; modern practice is to treat it in a secondary circuit more appropriate to the reduced particle size, and also to prevent massive swings in the circulating load through the rougher.
In reality, at each stage of separation, not all particles report to their correct outlet stream. Instead, there are misplaced particles that go to the wrong outlet. Consequently, the rougher concentrate is sent to a cleaning stage to recycle the misplaced particles (a procedure that of course also generates its own misplaced particles). When selectivities are very low, as is typical of nonsulfide flotation, additional recleaner stages may be needed. Although extra complexity may improve the separation, the improvements achieved must justify the cost of the additional equipment.
Because any given item of equipment is designed for a specific feed rate and particle size, variations in the feed rate and the size distribution of particles being treated should be minimized. Thus, circuits should be as simple as possible, with recycle minimized, since fluctuations in a recycle stream can become magnified and cause marked variations in the flow rate through, and an expansion of the particle size range being treated in, a given item of equipment. In turn, this reduces the separation efficiency.
Because output depends on grade and recovery, determination of optimum plant capacity (before or after startup) is a very complex issue. Some plants have combined concentrate grade and recovery with the smelter schedule and use this to maximize the profit per tonne of input. Others have suggested that the profit from the plant should be maximized via high throughput and consequential high tailings losses. In reality, the subject should be viewed in terms of the economics of the whole mining, concentrating, and smelting operation, under which the analysis becomes exceedingly complex, because of the effect of outside parameters and the difficulties in defining optimum.
Mining and mineral processing have significantly contributed to the advancement of human civilization and national economies, but they also have the potential to cause serious environmental degradation. As a result, the industry, with oversight by governmental agencies, is increasingly moving toward sustainable and environmentally friendly practices. The examination of mining and mineral processing trends reveals that production is increasing due to the demand from population growth, urbanization, and industrialization. The continual increase in demand is driving new mining developments throughout the world, as mineral commodities play increasingly larger roles in the economies of select countries. Developing mining and mineral processing projects, while minimizing adverse environmental impacts, poses a significant number of challenges. This book focuses on such challenges.
Mineral processing techniques have been suggested for the recovery of nickel-based alloys from spent batteries. The process involved hammer milling, magnetic separation, knife milling, a second and a final magnetic separation, and size separation. The AB5 alloy type from NiMH batteries can be used as an alloy in stainless steel.
D. A. Bertuol and colleagues have examined a method of mechanical processing of nickel batteries labeled NiMH. A sample of the spent batteries was milled, and the polymer and metal fractions were magnetically separated. The batteries, five types in total, selected for the characterization were of the AB5 type, because titanium was not found. Titanium is one of the most important elements present in batteries of the type AB2.
In the batteries characterization, the components were manually separated and classified as accumulators (active battery components), polymers, and metals (electronic circuits and contacts). The average results showed that of the total weight, the accumulators represented 81.5%, polymers 17.3%, and metals 1.2%. Among the five selected batteries, three were prismatic accumulators (denoted batteries 13) and two were cylindrical accumulators (denoted batteries 4 and 5). Overall, the electrodes represented >50wt% of the accumulators for all batteries and for batteries 4 and 5, 70wt%.
For all batteries, it was possible to separate the perforated plates from the paste in the negative electrodes. The positive electrode paste of the batteries with cylindrical accumulators was easily removed from the separator, whereas it was not possible to separate the components of the positive electrode for the prismatic accumulators, these being a very fine metallic screen impregnated with the paste.
The battery cases are made essentially of a NiFe alloy that is also contaminated by the presence of substances from the electrolyte, glues, and other contaminants that adhered to the case surface. The perforated plates from the negative electrode were also basically a NiFe alloy with elements such as cerium, lanthanum, manganese, and sulfur, related to the contamination of the plates by the electrolyte and the paste. The metals content of the pastes from negative electrodes, and of the paste and screen from positive electrodes, are presented in Table 2. The negative electrodes of batteries 14 had a high concentration of REEs, nickel, and cobalt. In the negative electrodes, nickel was not present in batteries 1 and 3. In battery 2, there was a high concentration of lanthanum but no cerium and neodymium. In this battery, the concentration of praseodymium was higher and the manganese concentration lower than in the other batteries. The positive electrodes from batteries 1 to 4 contained mainly nickel, cobalt, and zinc.
A significant outcome of the battery composition analysis was that battery 5, despite being labeled NiMH, contained a large amount of cadmium and in reality was NiCd. In general, this feature complicates the recycling of NiMH batteries as it is not possible to characterize all batteries before sending them to the recycling process. Therefore, an efficient recycling process should foresee the possible presence of NiCd batteries together with NiMH batteries.
Mineral processing comprises many unit operations, such as gravity concentration, magnetic concentration and flotation, which are all aimed at extracting valuable material from ores. Usually, the process operation conditions are defined to control the balance between a high recovery of the desired metal and a high grade value of the metal in the product outflow (Mndez et al., 2009a). These processes usually include multiple stages that are interconnected (forming circuits) to maximize recovery and concentrate grade. The design and analysis of these circuits, including the design and analysis of each stage, continues to be a challenging task (Ghodadi et al., 2011).
A designer initially solves a synthesis problem (for any process) by trial and error. There are many arrangements of a concentration circuit that correspond to an acceptable trial and error solution: many of these arrangements can be incorrect, ineffective or highly expensive, which is shown when feedback on an existing process becomes available. Concentration circuits commonly evolve over time, solving some existing problems, while creating new ones (Schena and Casali, 1994).
Several methods for the design of these circuits have been presented in the literature: these methods aim to obtain a systematic procedure to replace the trial-and-error method, which is time-consuming and requires much experimentation. Among the methods developed are those that use heuristics to develop a feasible design or improve an existing design (Connolly and Prince, 2000). However, these procedures use rules that are not always satisfied or that contradict each other and therefore do not guarantee an optimal design. Other methods use optimization or mathematical programming procedures (Mndez et al., 2009b, Ghobadi et al., 2011), using a superstructure to create a set of alternatives from which the optimum design can be selected. However, the use of these methods requires training in optimization techniques because the problems are usually formulated as MINLP models, for which there are no commercial codes that ensure optimality. For the aforementioned reasons, none of the developed methodologies are widely used in industry.
Floating is difficult to model and ore characteristics change with mine operation. Currently, there is no theoretical model that can predict the floatability of different species of a mineral, so that experimentation is necessary to develop models that can be used to design these systems. However, these experimentally-based models have a limited range of application depending on the experimental conditions and the number of experiments. The compositions and mineralogical species change with mine operation, which in turn affects the floatability behavior and undermines the model validity. Thus, there are at least two sources of uncertainty: the model and the ore characteristics.
Sensitivity analysis (SA) can be employed to address uncertainties in the model and the application scenarios, thereby facilitating the evaluation of process structures and operational behavior. Lucay et al. (2012) applied local SA to analyze and design separation circuits. The authors studied the effect of each stage on the general circuit by identifying the relationships between the recovery of each stage and the global recovery of the circuit. However, local SA only considers the neighborhood of the input variation and the effect of each input parameter is measured by keeping all the other input parameters at their nominal values. Global SA can overcome these limitations and has other advantages (Saltelli et al. 2000).
This work aims to show how a global SA can be used in the analysis, design and retrofit of concentration circuits. This work is expected to complement current design techniques, whether trial-and-error methods, heuristics or optimization.
Mineral processing operations involve a number of process variables that change randomly with uncertain frequencies. The control strategies developed with the use of PID controllers have been found to be inadequate, especially in non-linear systems and systems with large lag times. The present development to solve these problems falls under two categories:
The self-tuning control algorithm has been developed and applied on crusher circuits and flotation circuits  where PID controllers seem to be less effective due to immeasurable change in parameters such as the hardness of the ore and wear in crusher linings. STC is applicable to non-linear time-varying systems. It however permits the inclusion of feed forward compensation when a disturbance can be measured at different times. The STC control system is therefore attractive. The basis of the system is
The disadvantage of the set-up is that it is not very stable and therefore in the control model a balance has to be selected between stability and performance. A control law is adopted. It includes a cost function CF, and penalty on control action. The control law has been defined as
A block diagram showing the self-tuning set-up is illustrated in Figure20.26. The disadvantage of STC controllers is that they are less stable and therefore in its application, a balance has to be derived between stability and performance.
The empirical model predicts the process output for a certain predicted time. The error is not fixed as in a PID system, but extends over a time period and minimized. The concept is therefore time based and known as an extended horizontal control system. The algorithm is known as multivariable optimal constrained control algorithm (MOCCA). The MOCCA system can be considered an improvement on the level concept described earlier. It is based on the fact that the prediction of output equals the sum of the future actions plus past control action. It is developed around a step response under steady-state conditions by combining
To derive the model, Sripada and Fisher  considered a steady-state condition for a single inputsingle output system (SISO). The predicted output for horizon 1 to P is obtained in N number of step responses. The future and past control actions were written as
The predicted horizon P is the number of predicted outputs that the control objective has been optimized. The control horizon H is the number of future control actions which minimize the cost function against the predicted horizon.
Optimisation of the control system is achieved from performance criteria including any constraints. It is necessary to know the set point and predicted output trajectories for future control effort. The errors and control efforts have to be minimized. For the error trajectory the square of the difference of set point trajectory and the predicted output trajectory is taken. Taking these into consideration, Vien etal.  describe the cost function, Cf, in terms of minimising the error trajectory plus control effort. Taking the weighted least square performance, the cost function Cf is given as
Based on the process model, the control block calculates the predictions for future control actions and the supervisory block generates the desired set point trajectory. The feedback loop with filter and disturbance predictor corrects incongruity between the model and unaccounted (therefore unmeasured) disturbances. It also reduces the noise levels. The predictor in the feed back control loop intimates the future effects of disturbances. Combination of the feed back corrections and the predictions from the model provides the necessary estimate of output.
MOCCA has been found to be far superior to the conventional PID or PI controllers and is being increasingly used. It is particularly useful where long time delays are involved. Its advantage is that it uses discrete step response data and can be used to model processes with unusual dynamic behaviour. Its added advantage over the PID system of control is that it rises faster and has no overshoot. This system has been used successfully in control of grinding circuits. Circuit designs are continuously upgraded and the interested reader needs to consult appropriate text books and literature.
In mineral processing, ores are concentrated using flotation a technology that requires several stages (forming a flotation circuit) to separate the valuable minerals from the gangue. The flotation circuit design is a complex task, which is why several authors have proposed design procedures based on optimization. However, the modeling of the recovery of each flotation stage is complex, requires experimentation, and depends on many variables, including the selected circuit and mineral feed (which changes over time). These difficulties indicate that there are uncertainties in the actual values of the recovery of each stage and the information necessary to create the circuit design. This manuscript analyzes the effect of uncertainty on the recovery during each stage to facilitate the selection of an optimal circuit. We conclude that approximate values for stage recoveries are sufficient for the selection of flotation circuits.
Iron ore processing by Rio Tinto in the Pilbara region of Western Australia does not involve any chemical treatment. Flowsheets for the Brockman 2 and Paraburdoo processing plants are given in Figures 8.4 and 8.5 (Kinnel, 2013). The flowsheets are relatively simple. Dry processing involves up to three crushing circuits to produce lump and fines. Wet processing is primarily aimed at the removal of clays. Key processing variables include the following:
Crusher amps: crusher current (amps) is an indication of how hard the crusher is working. Alow bowl level, small gap setting, and high amps are a sign that the crusher could be working too hard. However, a high bowl level and low amps indicate the crusher is working too little. The crusher amps can also be affected by the hardness of the feed type.
Crusher product tons per hour: the crusher product tons per hour can be affected by the gap setting. A smaller gap setting lowers the tons per hour treated by the crusher but overall could increase the amount of material produced in the desired product size fraction. This could also minimize recirculation in the system, reducing the overall power consumption.
In mineral processing, an excessively large mean particle size results in inefficient extraction of the wanted materials whilst over-grinding is time-consuming and expensive. For efficient operation, therefore, ores should be ground to an optimum particle-size distribution, and hence there is a need for an instrument which indicates the distribution in order to enable grinding operations to be controlled effectively. Such an instrument has recently been developed.(15961)
The principle of this instrument is based on centrifuging the particles, which are in an aqueous suspension (slurry), by pumping the slurry through a pipe shaped into a single-turn helix. The pipe is of rectangular cross-section and the depth of the pipe is small compared with the width. The concentration of particles in the uniformly dispersed suspension entering the helix is measured by a -particle absorption gauge over a region where the pipe is fitted with thin plastic windows, a 300 Ci 90Sr/90Y source and scintillation detector being mounted on the arms of a small C-frame.
As the slurry flows through the helix, the particles are centrifuged away from the inner wall towards the outer wall of the pipe, thus creating a concentration gradient across the pipe. The concentration near the inner and outer walls is also measured downstream of the helix, by moving the source and detector in turn to two positions where the pipe is also fitted with thin plastic windows. The measured concentrations are correlated empirically with available particle-size information so as to provide a basis for calibration.
The measuring head is moved automatically by an electro-pneumatic device and a nomogram, derived using the calibration data, is solved by using a small electro-mechanical computer. An accuracy within 2% (1 ) is claimed.
Most iron ore processing in the fine particle ranges is carried out on a wet basis. The alternative is dry magnetic separation, but many laboratory tests have shown that it is not as efficient as wet magnetic separation in the fine particle ranges. However, some iron ore resources are located in dry areas where adequate water supply is very difficult to obtain and hence dry processing is the only option.
Dry permanent magnetic drum separators have been developed to separate magnetite and have been used in industry. And a SLon-1000 vertical ring and vibrating high-gradient magnetic separator has been developed for dry magnetic separation. It can treat weakly magnetic minerals in the size range 02.0mm and has been used in industry for upgrading roasted kaolin of 00.2mm by removing the impurities of oxidized iron ore and other weakly magnetic minerals. It is a potential dry magnetic separator for oxidized iron ore beneficiation.
In mining and ore processing XRF is used for quality and process control. The spectrometers used differ very much depending on their application. Laboratory spectrometers for quality control may be WD or ED systems with tube excitation. The on-stream spectrometers are located at in-plant locations, can be WD and ED systems, with radioisotope excitation or X-ray tube excitation and equipped with flow cells. In-stream instruments can be installed in slurry streams, mainly equipped with radioisotope excitation and scintillation counters for single-element determination. Field instruments must be portable and battery operated. The big advantage of XRF over other analytical tools for that application is its simplicity and speed. The usefulness of X-ray instruments to selective mining is well established, the information being used for orewaste sorting.