The efficient recovery of fine argentite from polymetallic lead–zinc (Pb–Zn) sulfide ore is challenging. This study investigated nanobubble (NB) adsorption on the argentite surface and its role in enhancing fine argentite flotation using various analytical techniques, including contact angle measurements, adsorption capacity analysis, infrared spectroscopy, zeta potential measurements, turbidity tests, microscopic imaging, scanning electron microscopy, and flotation experiments. Results indicated that the NBs exhibited long-term stability and were adsorbed onto the argentite surface, thereby enhancing surface hydrophobicity, reducing electrostatic repulsion between fine argentite particles, and promoting particle agglomeration. Furthermore, the NBs formed a thin film on the argentite surface, which decreased the adsorption of sodium diethyldithiocarbamate. Microflotation tests confirmed that the introduction of NBs considerably enhanced the recovery of argentite using flotation technology.

Efficient treatment of low C/N wastewater from ionic rare earth mining areas is an industrial challenge. This study employed a membrane bioreactor to acclimated a partial nitrification-denitrification microbial community within 30 days by controlling free ammonia (FA) concentration, the operating parameters were optimized using the response surface method, achieving a total nitrogen removal rate as high as 97.12 %, and saving 33 % of the carbon source during the denitrification stage. Three-dimensional fluorescence analysis indicates that the main components of the soluble microbial products (SMP) and extracellular polymeric substances (EPS) are tryptophan protein-like and humic acid substances, and the fluorescence intensity increased significantly, its protein and polysaccharide contents were significantly increased. High-throughput sequencing results showed that ammonia-oxidizing bacteria were successfully enriched, with Nitrosomonas becoming the dominant species, and the FA control strategy enhanced the abundance of functional nitrogen metabolism genes hao, nirK, and norB. This study provides reliable data support and theoretical guidance for the use of membrane bioreactor to treat low C/N wastewater from rare earth mining areas.

The recovery of fine silver (Ag)-bearing minerals from polymetallic lead–zinc (Pb–Zn) sulfide ore is crucial. This study investigated the effect and mechanisms of nanobubbles (NBs) on fine argentite flotation through contact angle tests, adsorption capacity tests, zeta potential tests, turbidity tests, microscopic observation, and flotation tests. The results indicated that NBs exhibited long-time stability and adsorbed onto the argentite surface, thereby enhancing surface hydrophobicity, reducing electrostatic repulsion between fine argentite particles, and strengthening their agglomeration. Additionally, NBs reduced the specific surface area of the fine argentite particles, leading to reduced adsorption of sodium diethyldithiocarbamate (DDTC) on the argentite surface. Micro-flotation tests indicated that the introduction of NBs enhanced argentite flotation recovery. Furthermore, flotation tests on actual ore confirmed that NBs enhanced Ag-bearing minerals in Pb concentrate.

Raw ores are mixed with filling materials during mining, which makes it difficult to sort copper‑sulfur resources. Inductively coupled plasma analysis revealed that the primary dissolved metal ions from the filling materials were Ca2+ and Fe3+. Flotation solution chemistry, flotation tests, and X-ray photoelectron spectroscopy measurements confirmed that the dissolved calcium and iron materials were adsorbed onto the chalcopyrite surface in different component forms, which reduced chalcopyrite recovery. Moreover, zeta potential measurements, turbidity analysis, microscope observation, scanning electron microscopy detection, adsorption capacity measurements, and contact angle measurements revealed that the dissolved anions and cations from filling materials can change the surface zeta potential of chalcopyrite and filling materials, weaken the mutual repulsion energy, and strengthen the aggregation effect, thus reducing the adsorption of butyl xanthate on the chalcopyrite surface and its surface hydrophobics. The chalcopyrite recovery eventually decreased from 96.41% to 86.63%.

The flotation of sulfide minerals has been always a problem due to the slime coating of serpentine. In this study, phytic acid (PA) was introduced as a selective depressant to eliminate the adverse effect of serpentine on the flotation of pyrite. Micro-flotation tests showed that the flotation performance of pyrite depressed by serpentine could be regained with the addition of PA. In addition to micro-flotation tests, the action mechanism of PA was systemically investigated based on a series of analyses including zeta potential tests, adsorption measurements, turbidity studies and X-ray photoelectron spectroscopy (XPS) analyses. It was found that PA could chemically react with the active Mg sites and then selectively adsorb onto serpentine surface. In this case, the surface potential of serpentine was shifted from positive to negative at pH 9, thus resulting in the transformation of electrostatic interaction between pyrite and serpentine from electrostatic attraction to electrostatic repulsion. Hence, the slime coating of serpentine on pyrite surface was removed and the adsorption amount of potassium butyl xanthate (PBX) onto pyrite surface was regained. In this way, the adverse effect of serpentine on the flotation of pyrite was eliminated.

为了应对工业废水迅速发展对环境带来的危害,探讨了电絮凝(EC)和电Fenton(EF)工艺对稀土冶炼含铅废水中Pb2+去除效果的研究.通过调整电压、pH值和停留时间等因素分析电絮凝反应中Pb2+去除效果.同时H2 O2投加量是调控电Fenton反应的关键.结果表明,对于铅含量为0.35 mg/L的实际废水其最佳操作条件为:停留时间180 min,电压20 V,进水pH=3,质量分数30％的H2 O2投加量为330 mg/L,对Pb2+达到了较好的去除效果,处理后的出水Pb2+浓度为0.06 mg/L,达到了国标(GB 26451—2011)要求.

Due to the optimum dissolution of the hydroxyl ion, serpentine is positively charged and tends to cover the sulfide mineral surface as a slime coating through electrostatic attraction, which intensively worsens sulfide flotation. To handle this problem, the sodium of polyaspartic acid (PASP) was employed as the selective depressant for the flotation of pyrite from serpentine in this work. Micro-flotation results showed that the fine serpentine of −10 µm could intensively hinder pyrite flotation, with a maximum decrease of about 75.8% in pyrite recovery at pH 9. However, at this pH, pyrite recovery remarkably increased from 20.17% to 92.68% when 15 mg/L PASP was introduced. Zeta potential results depicted that the addition of PASP overcompensated the positive charge on the serpentine surface and rendered it negative, while it had little impact on that of pyrite. Hence, the hetero-coagulation between serpentine and pyrite was broken due to the electrostatic repulsion, which was further confirmed by turbidity results. After that, the adsorption of PAX on the pyrite surface was restored, and the selective flotation of pyrite from serpentine was obtained. XPS analyses revealed that the chelation interaction between the carboxylate groups in PASP and the magnesium cations that remained on the serpentine surface were the main driving forces for the adsorption of PASP on the serpentine surface.

Rare earth carbonate precipitation is mainly amorphous, of large volume and difficult to filter. To prepare crystalline rare earth carbonate, mother liquor of heavy rare earth was taken as research object, and the experimental scheme was designed based on the response surface central composite design (CCD) method. The concentration of mother liquor, aging time and seed crystal dosage were taken as independent variables, and the particle size of rare earth carbonate was taken as the response value to establish a quadratic polynomial numerical model to optimize the reactive-crystallization process of rare earth carbonate. The results show that these three factors have significant effect on the particle size of rare earth carbonate, and the influence order is mother liquid concentration > aging time > seed crystal dosage. Moreover, the interaction between mother liquor concentration and seed crystal dosage has a significant effect on the size of rare earth carbonate particles. The optimal parameters predicted by the model are as follows: the concentration of mother liquid is 1.75 g/L, seed crystal dosage is 13.56 wt%, and aging time is 8 h. Under these conditions, the predicted particle size is 28.74 μm, and the experiment particle size is 28.23 μm, between both, the relative error is 0.73%, which indicates that the established response surface model has a good prediction effect and a certain practical significance to guide the reactive-crystallization process of rare earth carbonate. The obtained rare earth carbonate has a crystallinity of 97.82%, uniform particles size, and low-hydrated crystals with a tengerite structure.

Molecular dynamics simulation was adopted to study the interaction between sulfosalicylic acid and aluminum, lanthanum and yttrium, and adsorption on kaolinite surfaces. A complexation reaction occurs between sulfosalicylic acid and aluminum, with an interaction energy of -10472.05 kJ/mol. O-Al covalent bonds are formed with a peak value of 7.93, while there is only weak adsorption between sulfosalicylic acid and rare earth ions. A hydrogen bonding reaction with 13605.82 kJ/mol energy occurs between sulfosalicylic acid and the surface of kaolinite (100). Thus, sulfosalicylic acid can form a complex with free aluminum ions, and can also be adsorbed on kaolinite by hydrogen bonding with aluminum in kaolinite (100) surfaces. Leaching of ion-adsorbed type rare earth ore was performed with aluminum inhibited, results show that when sulfosalicylic acid dosage increases from 0 to 0.15 wt%, aluminum ion concentration in the leaching solution decreases from 273.23 to 47.19 mg/L. And the effect of leaching pH value on the effect of sulfosalicylic acid on aluminum inhibition was studied, the result shows that, when the leaching pH value is 4.0-5.0, the rare earth leaching rate and the aluminum ion concentration basically remain unchanged. The molecular dynamics simulation results were verified by detection and analysis of XPS and SEM. (C) 2019 Published by Elsevier B.V. on behalf of Chinese Society of Rare Earths.

为提高稀土浸出过程的选择性,以酒石酸为抑杂剂考察其对硫酸铵浸出稀土时浸出行为的影响.结果表明,以硫酸铵为浸出剂、酒石酸为抑杂剂浸出不同种类稀土矿时,与单独以硫酸铵作浸出剂相比,浸出液中铝、铁去除率可达90%以上,且不影响稀土浸出率.对酒石酸与Al3+、Fe3+的溶液化学计算和分析结果表明,3<pH<6时,酒石酸可解离出与金属离子稳定络合的配位离子C4H5O6-、C4H4O62-,与离子型稀土矿浸出时矿石表面解离出的Al3+、Fe3+生成难溶络合物而提高稀土浸出的选择性,浸出过程酒石酸络合铁离子与铝离子的适宜pH范围为4~5.

针对Wemco自吸式浮选机,采用CFD方法对叶片倾角β分别为前倾(-22.5.、-37.5.)、径向(0.)和后倾(22.5.、37.5°)浮选机的流场进行数值模拟,探讨了叶片倾角对浮选机内流场特征、湍动能和压力分布的影响,分析了不同转速下叶片倾角对Wemco浮选机循环量和功耗的影响.结果表明,叶片倾角对浮选机内流场结构影响较小,但对叶轮旋转区域的轴向速度、湍动能和压差等影响较大.叶片倾角为0.时,流体在叶轮旋转区域的轴向速度、湍动能和压差最大;叶轮循环量和功耗均随着转速的增加而增加,径向叶片比前倾和后倾叶片的循环量和功耗更大.

There is large amount of waste water discharged by concentrator,and the solid suspended substance,flotation reagent,heavy metal ions and other substances in the wastewater are high,and the cost of effluent treatment is high and the sec-ondary pollution is easy to be caused when direct discharge,and the direct recycling affects the floatation index.Therefore,it is of great significance to realize recycle utilization of mineral processing wastewater to save limited water resources and reduce environmental pollution.The results of simulated wastewater flotation test showed that the Al3+and Fe3+in the waste water had a significant inhibitory effect on the flotation of sulfide copper,and the effect of Pb2+,Zn2+and Fe2+on the flotation of sulfide copper was relatively lower.With the concentration increase of butyl xanthate and Z-200,selectivity of collector on copper de-creased.Coagulation,ozone oxidation technology on Anhui a copper ore dressing wastewater treatment results showed that the polyacrylamide for waste water of heavy metal ions and the settlement of suspended solids has significant effects; During the whole pH,ozone on butyl xanthate removal effect is remarkable,pH=8,ozone on Z-200 removal effect is good.Using floccula-tion sedimentation-ozonation process,the waste water is back to the flotation test,and its flotation index is litter lower than that of clear water flotation index,which is far better than the direct return flotation index of wastewater.Therefore,application of this process to treat total copper sulfide flotation tailings water,the bad component content can effectively reduce,and high effi-ciently utilize water resources.

研究了某浮选银精矿在“[Bmim] HSO4—硫脲”和“H2SO4—硫脲”两种浸取体系中银的浸出行为.研究了矿浆pH值、Fe3浓度、硫脲浓度、浸出时间等工艺参数对浸出的影响.结果表明,[Bmim] HSO4代替H2SO4作为pH调整剂,可在酸性较弱的矿浆pH值,较少的Fe3用量、较低的硫脲浓度条件下提高银浸出率9.80个百分点.

Separation of ion-absorbed rare earth elements (REE) by oxalate used as a precipitator is the principal extraction process employed in the mining industry. The oxalic acid precipitator plays the key role in the precipitation process. In this paper, the enhanced effect of a static magnetic treatment applied to the precipitation process of REs was investigated. Compared to the non-magnetization, the results of magnetic processes appeared to be better and its optimal precipitation conditions was obtained at pH value of the precipitation solution at 2,RE2O3: H2C2O4=2.5:1, a magnetic strength 400kAm−1, a 20min magnetization time in a magnetized oxalate solution. The dosage of oxalate precipitator is less 5% to that employed in conventional processes conditions with non-magnetization while the purity of the rare earth still keeps stable. The electrical conductivity of the oxalate precipitator was increased by magnetization. A possible change in the electron spin of the magnetized precipitator ultimately affected the pace of rare earth precipitation crystallization and growth direction of the rare earth-bearing precipitated crystals.

To judge the possibility of rapid flotation separation on chalcopyrite and pyrite, the dynamics parameter K (flotation rate coefficient) was used as the basis. At first, through investigating the change of K value of chalcopyrite and pyrite pure minerals under the conditions of pulp density, flotation particle size and flotation machine speed, respectively, the dynamics mechanism and possibility of rapid flotation separation on chalcopyrite and pyrite was interpreted to some extent. Meanwhile, to make the technology of rapid flotation separation feasible on chalcopyrite and pyrite applicable in the industry, the dynamics characteristic of two mixed ore in different size fraction that composed by chalcopyrite and pyrite was further investigated. lt was proved that flotation rate difference between chalcopyrice and pyriec could be expanded through controling appropriate flotation conditions and then furfiued the rapid prioriey flotation of chalcopyrite.

The kinetics of the forward extraction of NbO(C2O4)33− by N235 was investigated using a Lewis cell. Different parameters such as the niobium concentration, oxalic acid concentration, hydrogen ion concentration, and R3N concentration, as well as extracting temperature were studied separately to examine their effect on the extraction rate of NbO(C2O4)33−, and a rate equation was deduced based on the analysis. The experimental results demonstrate that the rate of NbO(C2O4)33− extraction is controlled by a chemical reaction at the interface rather than in the bulk phase. Further analysis suggests that the NbO(C2O4)33− extraction rate is first order with respect to hydrogen ion concentration, niobium concentration, and R3N concentration, and zero order with respect to oxalic acid concentration. The mechanism of NbO(C2O4)33− extraction is controlled by the slow distribution of the R3NH+ at the interface to form NbO(C2O4)3⋅R3NH2− between the liquid phase and the interface.

利用MLA自动矿物分析仪(Mineral Liberation Analyzer)对福建某银多金属矿进行工艺矿物学研究,并根据银及铜铅锌矿物嵌布特征和赋存状态,选择合理的浮选原则流程,实现回收铜铅锌主金属的同时高效回收共伴生银.工艺矿物学研究表明,该银多金属矿黄铜矿与方铅矿共生关系密切,铁闪锌矿结晶粒度粗大,与硫铁矿物和脉石矿物连生.矿石中55.93％的银以辉银矿单体形式存在,其他银与铋、铅、硫元素形成硫银矿和辉铋银铅矿等以微细包裹体或类质同象形式存在于黄铜矿、铜蓝及部分方铅矿中,黄铜矿、硫银铋矿、方铅矿中的银分布率分别为29.96％、8.18％、4.53％,因此提高铜铅回收率是银高效回收的关键.浮选试验表明,对铜捕收能力强的Z-200与对铅矿物捕收能力强的SN-9#组合使用可以获得较高的铜铅回收率,银回收率与铜铅回收率具有相关性,铜铅回收率高时银回收率也高.当磨矿细度为-74 μm占70％,以SN-9#和Z-200组合捕收剂浮选铜铅矿物,以石灰调节矿浆pH值在10 ～11,可以获得含银3 704.1 g/t,银回收率90.26％的铜铅混合精矿.

The leaching experiments are performed on samples collected from a silver concentrate with cyanide, thiosulfate and thiourea solutions respectively to compare their extraction effects. The result indicates that, under the optimal process parameters, (Cyanide: ammonium salt added on the pretreated alkali, sodium cyanide concentration being 0.15 mol/L, pH value being 11, liquid-solid rate 2∶1, mixing strength 500 r/min, leaching time 48 h; Thiosulfate: oxidizing roasting pretreatment, ammonia concentration of 1.0 mol/L, copper sulfate concentration of 0.01 mol/L, sodium thiosulfate concentration of 0.2 mol/L, pH of 10, liquid-solid rate of 4∶1, mixing strength of 500 r/min, leaching time of 12 h;Thiourea:oxidizing roasting and then dilute sulphuric pretreatment acid pre-leaching, Fe3+ concentration of 7.5 ×10-3 mol/L, thiourea concentration of 0.18 mol/L, pH of 1.0, liquid-solid rate of 3∶1, mixing strength of 250 r/min, leaching time of 24 h], the leaching rate of Ag with the method of thiosulfate is 78.58 %, while the rate in the thiourea system is about 62.43%, 3.33%lower than that in the cyanide system.

研究低品位铌精矿硫酸焙烧及草酸浸出过程，考察影响铌提取效果的因素。结果表明，在硫酸焙烧阶段，在精矿粒级为-104+74μm、焙烧温度为200℃和硫酸浓度为80%的条件下，焙烧40 min，可使铌精矿中97%以上的铌和90%以上的钙留在渣相，93%以上的铁、锰及钠进入液相。在草酸浸出阶段，在草酸浓度为85%及浸出温度为80℃的条件下，含铌滤渣浸出60 min，可使铌浸出率达到98%以上，钙浸出率低于0.6%。