Assessment of pH-responsive nanoparticles performance on laboratory column flotation cell applying a real ore feed

Nanoparticles(NPs) can promote the column flotation process in mining industry. Nanoparticles' effects on column flotation process(copper recovery, grade and flotation rate constant) are assessed in Sarcheshmeh Copper Complex, Iran, through response surface methodology(RSM) optimization technique. The c-Al_2O_3, a-Fe_2O_3, SiO_2, and TiO_2 nanoparticles are selected for these experiments. A flotation rate constant is chosen as a response to assess the effect of nanoparticles on flotation in its kinetic sense.The process p H and nanoparticle dosage are selected as the influential parameters. Results obtained from RSM indicated that the maximum percentage of Cu recovery and grade is obtained at p H of 12 and nanoparticle dosage of 6 kg/t, through a-Fe_2O_3 and c-Al_2O_3 nanoparticles, respectively. Applying nanoparticles in particular c-Al_2O_3 and a-Fe_2O_3 increases the Cu recovery by 8–10% together with the grade by 3–6% in a significant manner. It is revealed that nanoparticles could effectively be applied in enhancing the flotation performance.     

Growth of covellite crystal onto azurite surface during sulfurization and its response to flotation behavior

In this work, the growth of copper sulfide crystal onto azurite surfaces during sulfurization and its response to flotation are investigated. Filed emission scanning electron microscopy-energy dispersive X-ray spectroscopy (FESEM) and X-ray diffraction (XRD) studies confirmed that the sulfurization of azurite is not limited to the mineral surface, but rather penetrates into the bulk to form covellite crystal (syn-CuS), creating favorable conditions for the stable adsorption of xanthate and greatly promoting the azurite flotation. Additionally, as demonstrated by X-ray photoelectron spectroscopy (XPS) and time of flight secondary ion mass spectrometry (TOF-SIMS) analyses, a redox reaction occurred during this process, and Cu(Ⅱ) onto the mineral surface was reduced to Cu(I). Correspondingly, reduced S^2- was oxidized to (S₂)^2-, (S_n)^2-, and even to deeper oxidation state S⁰, (S_xO_y)^n- and SO₄^2-. Excess sodium sulfide strengthens copper sulfide to form onto the azurite surface, and provides enough raw material for crystal copper sulfide to grow, resulting in the formation of “flake-like” covellite with a better crystallinity. However, the floatability of azurite decreased dramatically under this condition, because the generated massive colloidal copper sulfide in flotation pulp deteriorates the flotation environment, resulting in a decreased effective adsorption of collector onto azurite surfaces.     

Effect of organic depressant lignosulfonate calcium on separation of chalcopyrite from pyrite

In order to selectively separate chalcopyrite from pyrite, the effect of organic depressant lignosulfonate calcium (LSC) on the flotation separation of chalcopyrite from pyrite was investigated by flotation tests. The depression mechanism was studied by Fourier-transform-infrared (FTIR) analysis. The flotation tests of single mineral show that LSC can depress the flotation of pyrite in a certain pH range, but it has little effect on chalcopyrite flotation. Flotation separation of a mixture of chalcopyrite and pyrite can be completed to obtain a copper concentrate grade up to 24.73% with a recovery of 80.36%. IR analysis shows that LSC and butyl xanthate compete in absorption on pyrite surface, and there exists an LSC characteristic peak on pyrite surface. There is little adsorption of LSC on chalcopyrite.     

Influences of silver sulfide on the bioleaching of chalcopyrite, pyrite and chalcopyrite-containing ore

The effects of silver sulfide (Ag2S) on the bioleaching of chalcopyrite and pyrite were investigated in this paper. It has been shown that Ag2S enhanced the yields of bioleaching of chalcopyrite but inhibited the bio-oxidation of pyrite. The addition of Ag:S selectively increased the copper dissolution from the chalcopyritecontaining ores in shake flasks with a recovery of 85.3 % compared with 24.3 % without Ag2S, while slightly decreased the iron yields from 51% to 41.8%. The copper extraction of the chalcoopyrite-containing waste rock in column leaching charged with 18 kg mass increased up to 21.7% in the presence of Ag2S, while only 3.4% in the absence of the catalyst. The mechanism of Ag2S catalysis could be explained well by the “Mixedpotential model”.     

Fabrication of W@Cu composite powders by direct electroless plating using a dripping method

A direct electroless copper (Cu) coating on tungsten powders method requiring no surface treatment or stabilizing agent and using glyoxylic acid (C₂H₂O₃) as a reducing agent was reported. The effects of copper sulfate concentration and the pH of the plating solution on the properties of the prepared W@Cu composite powders were assessed. The content of Cu in the composite powders was controlled by adjusting the concentration of copper sulfate in the electroless plating solution. A uniform, dense, and consistent Cu coating was obtained under the established optimum conditions (flow rate of C₂H₂O₃ = 5.01 mL/min, solution pH = 12.25 and reaction temperature 45.35 °C) by using central composite design method. In addition, the crystalline Cu coating was evenly dispersed within the W@Cu composite powders and Cu element in the coating existed as Cu0. The formation mechanism for the W@Cu composite powders by electroless plating in the absence of surface treatment and stabilizing agent was also proposed.     

Recovery of copper sulfate after treating As-containing wastewater by precipitation method

The solid sodium hydroxide neutralized acidic As-containing wastewater till pH value was 6. Green copper arsenite was prepared after copper sulfate was added into the neutralized wastewater when the molar ratio of Cu to As was 2:1 and pH value of the neutralized wastewater was adjusted to 8.0 by sodium hydroxide. The arsenious acid solution and red residue were produced after copper arsenite mixed with water according to the ratio of liquid to solid of 4:1 and copper arsenite was reduced by SO₂ at 60 °C for 1 h. The white powder was gained after the arsenious acid solution was evaporated and cooled. Copper sulfate solution was obtained after the red residue was leached by H₂SO₄ solution under the action of air. The results show that red residue is Cu₃(SO₃)₂·2H₂O and the white powder is As₂O₃. The leaching rate of Cu reaches 99.00% when the leaching time is 1.5 h, molar ratio of H₂SO₄ to Cu is 1.70, H₂SO₄ concentration is 24% and the leaching temperature is 80 °C. The direct recovery rate of copper sulfate is 79.11% and the content of CuSO₄·5H₂O is up to 98.33% in the product after evaporating and cooling the copper sulfate solution.     

Adsorption behavior and mechanism of copper ions in the sulfidization flotation of malachite

Malachite is one of the main minerals used for the industrial enrichment and recovery of copper oxide resources, and copper ions are unavoidable metal ions in the flotation pulp. The microflotation, contact angle, and adsorption experiments indicated that pretreatment with an appropriate concentration of copper ions could improve the malachite recovery, and the addition of excess copper ions reduced the hydrophobicity of the malachite surface. The results of zeta potential tests indicated that sodium sulfide and butyl xanthate were also adsorbed on the surface of malachite pretreated with copper ions. X-ray photoelectron spectroscopy (XPS) results indicated that —Cu—O and —Cu—OH bonds were formed on the surface of the samples. After pretreatment with an appropriate concentration of copper ions, the number of —OH groups on the mineral surface decreased, whereas the number of Cu—S groups on the mineral surface increased, which was conducive to the sulfidization of malachite. After adding a high concentration of copper ions, the —OH groups on the mineral surface increased, whereas the number of Cu—S groups decreased, which had an adverse effect on the sulfidization flotation of malachite. Time-of-flight secondary ion mass spectrometry showed that pretreatment with copper ions resulted in a thicker sulfidization layer on the mineral surface.     

从会理锌矿铅锌矿石中分选铜的试验研究

为从会理锌矿含铜的铅锌硫化矿中分选铜,对会理锌矿含铜的铅锌硫化矿进行了系统的工艺矿物学研究和选矿工艺研究．工艺矿物学研究结果表明会理锌矿含铜的铅锌硫化矿中铜矿物以黄铜矿、银黝铜矿-砷黝铜矿、硫锑铜银矿、车轮矿等矿物形式存在,并以黄铜矿为主,其次为银黝铜矿-砷黝铜矿系列矿物．铜矿物嵌布特征复杂,与闪锌矿互相包裹及呈固溶体分离结构较为普遍．铜矿物嵌布粒级较均匀,多集中于＋0．08mm以上粒级中,铜矿物单体解离度相对较好．基于会理锌矿舍铜铅锌硫化矿的矿石特性,采用“铜铅混浮-铜铅分离-再浮锌”浮选工艺,在原矿含铜0．94％、铅0．92％、锌10．60％的情况下,可获得含铜17．76％、铅4．35％、锌12．87％,铜回收率62．89％的铜精矿,含铜2．99％、铅54．52％、锌10．48％,铅回收率50．34％的铅精矿,含铜0．75％、铅1．20％、锌55．86％,锌回收率88．56％的锌精矿,银在铜、铅、锌精矿中的回收率分别为50．93％、2．32％、29．51％,实现了从会理锌矿含铜铅锌硫化矿中分选铜的目标．     

Assessment model for tribological property of ceramic/stainless steel rubbing pairs in H2O2 solution

The tribological properties of ZrO2 ceramic and 1Cr18Ni9Ti stainless steel rubbing pairs were investigated using a special tribo-tester under different concentrations of hydrogen peroxide (H2O2) solution.The comparison analyses of the friction coefficients,worn volume rates,worn particles and surface roughness were conducted under the tested conditions.There were significant differences of the tribological properties of the rubbing pairs in the different concentrations H2O2 solution because of oxidation and corrosion.This research has revealed that the main wear mechanisms between the rubbing pairs are severe adhesive wear,abrasive wear and corrosive wear in the H2O2 solution with different concentrations.A model has been established to assess the most suitable concentration of the H2O2 solution for the tribological properties of the rubbing pairs.The result shows that 50% concentration of the H2O2 solution is the most suitable.The assessment result is consistent with the experimental result.It is believed that the knowledge gained in this study is useful for the optimization of the friction pairs in the extreme condition.     

Effect of pyrite content of feed and configuration of locked particles on rougher flotation of copper in low and high pyritic ore types

Misreported pyrite into copper concentrates dramatically declines copper grade and recovery. Copper flotation can be also more complicated if flotation feed comes from an elevated-pyritic copper ore. In this investigation, the effect of two different ore types(high pyritic and low pyritic feeds) was studied on rougher stage of industrial copper flotation circuit. Samples were taken from different streams and the structure of chalcopyrite within the pyrite and non-sulfide gangue minerals was examined in various size fractions for mentioned ore types. Results indicated that 72% and 56% of the total floated pyrite was transferred to concentrate in first four cells in the low and high pyritic feeds, respectively. Whereas, this proportion for floated SiO_2 in last ten cells was detected as 72% and 71%, respectively. A detailed interpretation of the effect of locked particles in different size fractions on rougher flotation cells is studied from industrial point of view.     

Potential control flotation of galena in strong alkaline media

The electrochemical oxidation of galena in collectorless and collector flotation systems, particularly in strong alkaline media, was studied. The results show that, with pH value higher than 12.5 and potentials below 0.17 V, the oxidation products of galena are elemental sulfur and HPbO2^- . Elemental sulfur was present on the mineral surface in excess of oxidized lead species due to dissolution of HPbO2^- , which is beneficial to the flotation of galena. Under the same conditions, sphalerite and pyrite were depressed as a result of significant surface oxidation. Diethyldithioearbamate (DDTC) was found to be the most suitable collector for galena flotation in strongly alkaline media. The very potential produced hydrophobic PbD2-the surface reaction produet of DDTC with galena, is 0 to 0.2 V. Meantime DDTC can depress the surface over-oxidation of galena.Investigations also indicate that, in the range of - 0.9 V to 0.6 V, hydrophobic PbD2 can be firmly adsorbed on galena.     

脂类捕收剂DLZ对黄铜矿和黄铁矿浮选的选择性作用

通过浮选实验、吸附量和红外光谱测定,考察了捕收剂DLZ对黄铜矿和黄铁矿浮选性能的影响及作用机理.结果表明:DLZ在pH=2.7～12.05时对黄铜矿的捕收能力强,最大回收率95.7%,而对黄铁矿的捕收能力弱,且PH=6.9～12.05时其回收率小于10%.用CaO作pH调整剂时,在pH=7～11时黄铜矿的回收率与用NaOH调PH相差不大,但黄铁矿可浮性被抑制,回收率低于5%.DLZ在黄铜矿上的吸附量比在黄铁矿上的大,特别是强碱条件下;其吸附量随用量的增加而增大.DLZ与矿物作用的红外光谱分析表明:黄铜矿与DLZ作用后出现了DLZ的相关特征吸收峰,而黄铁矿与DLZ以及Cu~(2+)作用前后的红外光谱曲线基本没有变化,可见DLZ在黄铜矿表面的吸附属于化学吸附,在黄铁矿表面的吸附属于物理吸附.     

A new technology for processing niobite ore found in Jiangxi province

A new process is proposed for refining niobite ore that is found in Jiangxi province of China. Niobite, also known as columbite or niobite–tantalite, is a mineral that contains tantalum and niobium. The separation process includes two-stage grinding, gravity concentration, magnetic separation, and flotation. The tantalum/niobium concentrate obtained had a grade of Ta₂O₅ 18.28%, Nb₂O₅ 41.62% at a recovery rate of Ta₂O₅ 49.08%, Nb₂O₅ 70.77%. Other minerals occurring along with the niobite, such as zinnwaldite, feldspar, and quartz, were also recovered to comprehensively utilize this ore.     

Comparative studies on phosphate ore flotation collectors prepared by hogwash oil from different regions

In order to compare the differences between variations of phosphate ore flotation collectors prepared by hogwash oils, JZQ-F collectors were prepared by the hydrolysis method using four types of hogwash oils from different regions in China. The components of the hogwash oils were determined using Fourier transform infrared spectroscopy(FIIR) spectra analysis, while the components of the JZQ-F were determined through gas chromatography–mass spectrometry(GC–MS). The flotation effects of the JZQ-F collectors were investigated by flotation experiments and the adsorption characteristics were determined through the quartz crystal microbalance(QCM-D). Results show that the JZQ-F collector from the Ordos area has 27.43% unsaturated fatty acids. JZQ-F collectors from Beijing, Qingdao, and Dongguan areas contain over 62% of unsaturated fatty acids, which can acquire phosphorous concentrates with a the grade of P_2O_5 above 31.96% and the recovery higher than 91.52%. These three collectors have a larger adsorption capacity, faster adsorption rate, and stronger adsorption structure on apatite surface, which contributed to a good adsorption performance and a favorable flotation effect.     

Water-oxygen interaction on marcasite(101) surface: DFT calculationOA

Marcasite (Fe S₂) is widespread in nature,its oxidation plays a vital role in acid mine drainage,mineral resource recovery,and photoelectric material applications.In this paper,the oxidation mechanism of marcasite has been studied for the first time using density functional theory (DFT).It is found that,unlike the oxidation of pyrite,the oxidation of marcasite merely occurs at surface S atoms.Under the coexistence of water and oxygen,S atoms around surface Fe atoms are replaced by O a...     

Influences of collector DLZ on chalcopyrite and pyrite flotation

The interaction mechanism of collector DLZ in the flotation process of chalcopyrite and pyrite was investigated through flotation experiments, zeta potential measurements and infrared spectrum analysis. Flotation test results indicate that DLZ is the selective collector of chalcopyrite. Especially, the recovery of chalcopyrite is higher than 90% in neutral and weak alkaline systems, while the recovery of pyrite is less than 10%. When using CaO as pH regulator, at pH=7-11, the floatability of pyrite is depressed and the recovery is less than 5%. Zeta potential analysis shows that the zeta potential of chalcopyrite decreases more obviously than that of pyrite after interaction with DLZ, confirming that collector DLZ shows selectivity to chalcopyrite and pyrite. And FT1R results reveal that the flotation selectivity of collector DLZ is due to chemical absorption onto chalcopyrite surface and only physical absorption onto pyrite surface.     

Effect of galvanic contact on the flotability of galena in the presence and absence of a collector

In this paper, the effect of various electrochemical environments in the galena flotation is investigated. The electrochemical environments consisted of a ball mill for grinding of galena ore and a Denver flotation cell for flotation of galena in the laboratory scale. In order to achieve the maximum recovery with sodium hyposulfite, the concentrations of 0.06 and 0.12 M of air and nitrogen gases have been used to control the pulp potential in the Denver flotation cell. The galena sample was from the “Era mine” which is located in the Kiyasar area, north of Iran. This mine contains: Galena (PbS) at least 22%, Fluorite (CaF₂) 73.37%, Quartz (SiO₂) 2.54% and other minerals such as Cerussite (PbCO₃) and Kaolinite (Al₂Si₂O₅(OH)₄). Flotation of Galena was conducted in a 0.12 M of sodium hyposulfite solution. It was found that the amount of recovery by this process was 85% and 63% in the presence and absence of potassium amyl xanthate (KAX) and at an estimated potential of 280 to 350 mV respectively while the recovery was 70% and 52% at an approximate potential of 175 to 210 mV in water in the presence and absence of collector (KAX).     

Technological conditions and kinetics of leaching copper from complex copper oxide ore

The kinetic behavior of leaching copper from low grade copper oxide ore was investigated. The effects of leaching temperature, H₂SO₄ concentration, particle size of crude ore and agitation rate on the leaching efficiency of copper were also evaluated. And the kinetic equations of the leaching process were obtained. The results show that the leaching process can be described with a reaction model of shrinking core. The reaction can be divided into three stages. The first stage is the dissolution of free copper oxide and copper oxide wrapped by hematite-limonite ore. At this stage, the leaching efficiency is very fast (leaching efficiency is larger than 60%). The second stage is the leaching of diffluent copper oxides, whose apparent activation energy is 43.26 kJ/mol. During this process, the chemical reaction is the control step, and the reaction order of H₂SO₄ is 0.433 84. The third stage is the leaching of copper oxide wrapped by hematite-limonite and silicate ore with apparent activation energy of 16.08 kJ/mol, which belongs to the mixed control.     

Fenton法预处理颜料企业生产废水

针对颜料企业生产废水pH波动大,有机污染物成分复杂与可生化性差的问题,采用Fenton法对颜料废水进行了预处理实验研究.通过研究在不同实验条件下,Fenton试剂预处理该类废水的COD去除率及处理成本,验证了Fenton法处理颜料企业生产废水的可行性并得到Fenton法预处理此类废水的最优实验条件.实验结果表明,Fenton试剂处理该废水的COD去除率最高可达81%;在最优实验条件下,COD去除率为52%,处理成本为4.3元/t;Fenton法处理颜料企业生产废水的影响因素依次为H2O2∶Fe2+(mol∶mol)H2O2∶COD(g∶g)出水pH.     

Role of ammonium sulfate in sulfurization flotation of azurite: Inhibiting the formation of copper sulfide colloid and its mechanismOA

In this study, the role of(NH₄)₂SO₄ during the sulfurization of azurite and its response to flotation were investigated. The flotation results showed that adding(NH₄)₂SO₄ prior to sulfurization decreased the formation of colloid in flotation pulp, and the floatability of the suppressed azurite caused by excess sodium sulfide was restored. After adding(NH₄)₂SO₄ prior to sulfurization, the formation ...