TBP-NX混合萃取体系从盐湖卤水提锂试验研究

本文主要采用30%TBP+20%NX混合萃取体系对青海某盐湖卤水提锂过程涉及的萃取、洗涤、反萃等过程进行了详细研究。萃取过程：运用正交试验法对萃取相比、铁锂摩尔比及水相酸度进行显著性分析,得出相比对30%TBP+20%NX混合萃取体系影响显著,水相pH影响不明显;在O/A=1.5、n(Fe³⁺/Li⁺)=2和水相酸度 0.05 mol/L的最佳条件下,Li⁺、Mg²⁺萃取率分别为89.19%和1.69%;采用饱和容量法测量30%TBP+20%NX混合萃取体系单位负载锂量为2284 mg/L,所需分相时间为15 min。洗涤过程：在c(H⁺)=0.25 mol/L和O/A=20:1的最佳条件下,Li⁺、Mg²⁺的洗涤率分别为5.12%、89.46%。反萃过程：在c(H⁺)=3 mol/L和O/A=15:1的最佳条件下,Li⁺、Mg²⁺反萃率分别为80.08%、67.56%。TBP-NX混合萃取体系Li⁺负载量及反萃性能较好,但分相性能相对较弱。     

澳洲某锂辉石选矿试验研究及生产实践

国内多家锂辉石选矿厂处理澳洲某锂辉石矿,由于生产工艺和药剂制度的限制,锂精矿的品位和回收率均不高,造成极大的资源浪费。因此,对该进口的澳洲锂辉石矿进行选矿试验研究具有十分重要的现实意义。通过对澳洲进口的锂辉石矿进行X荧光半定量分析、偏光显微分析,结果表明:该锂辉石矿原矿中Li_2O的含量为1.22%,锂主要以锂辉石和锂云母形式赋存,脉石矿物主要为石英、长石等。经过选矿试验研究,确定了预先脱泥、HP作捕收剂、正交试验得到Na OH、Na_2CO_3、CaCl_2用量的优水平组合,经一次粗选两次精选两次扫选、扫选Ⅰ中矿返回精选Ⅰ,最终得到了锂精矿品位为5.86%,回收率为75.27%。同时,对进口该锂辉石矿的某企业选矿厂进行现场调试与工艺改造,使得锂精矿Li_2O的品位在4.5%以上,回收率在70%以上,年增经济效益为4 500万元。     

某复杂难选低品位锂辉石矿综合回收工艺

根据矿石性质研究某锂辉石矿产资源高效综合利用的工艺流程。采用硫酸作pH调整剂,十二胺作捕收剂优先浮选云母,以NaOH作调整剂、CaCl2作活化剂,油酸作捕收剂浮选锂辉石矿物,浮选粗精矿经再磨后以Na2CO3为调整剂精选,获得含Li2O 6.04%、回收率76.77%的锂辉石精矿和纯度较高的云母精矿。     

Paste backfill of high-sulphide mill tailings using alkali-activated blast furnace slag: Effect of activator nature,concentration and slag properties

The effect of activator type, concentration and slag composition on the strength and stability properties of paste backfill (CPB) of high-sulphide tailings using alkali-activated slag (AAS) as binder (7 wt.%) were investigated in this study. Acidic and neutral (AS–NS) slags were activated with liquid sodium silicate (LSS) and sodium hydroxide (SH) at 6–10 wt.% concentrations. Ordinary Portland cement (OPC) results were used for comparison. The strength development was found to remarkably improve with increasing the concentration from 6 to 8 wt.%. Further increase in concentration did not enhance the strength. SH was determined to produce higher early-age strength whilst LSS produced higher long-term strengths as an indication of slag selectivity for activators. More extensive gypsum formation was observed at lower concentrations in SEM/EDS studies. An increase in Na₂O concentration raised the activator consumption. High concentrations also led to poorly crystallized C–S–H gel, loose structure and drying shrinkage cracks especially in NS–SH samples. A reduction in total porosity up to 20% was obtained in AAS samples compared to OPC. Amorphous structure, chemical modulus ratio and/or basicity index (BI) values were seen to control the pozzolanic reactivity, and therefore, the alkali-activation and hardening process.     

Mineral carbonation of PGM mine tailings for CO2 storage in South Africa: A case study

The ultra-fine milled tailings generated during the processing of PGM ores in South Africa have a theoretical potential to sequester significant amounts of CO₂ (∼14 Mt per annum) through mineral carbonation. Mg-bearing orthopyroxene is the major sequestrable mineral in these tailings, which also contains significant quantities of Ca-bearing plagioclase, as well as minor quantities of clinopyroxene, olivine, serpentine and hornblende. In this study, the feasibility of using PGM tailings to sequester CO₂ has been investigated empirically using the two-step, pH swing method. The rates and extents of cation (Ca, Mg and Fe) extraction and subsequent carbonation were determined and compared. Both organic (oxalic and EDTA) and HCl solutions were utilised in the cation extraction step, which was conducted at time periods up to 8 h and at a temperature of 70 °C. The extents of cation dissolution were relatively low under all experimental conditions investigated, particularly for the case of Mg (between 3.3% and 5.0% extraction). A comparison of the extents of leaching with the mineralogical composition of the tailings indicated that the extracted Mg originated primarily from clinopyroxene, with the orthopyroxene remaining relatively inert under the experimental conditions. Subsequent carbonation of the acid leach solution after pH adjustment with NaOH resulted in the rapid formation of a number of carbonate minerals, including gaylussite (Na₂Ca(CO₃)₂·5(H₂O)), magnesite (MgCO₃), hydromagnesite (Mg₅(CO₃)₄(OH)₂·4H₂O), dolomite (CaMg(CO₃)₂), ankerite (Ca(Fe,Mg)(CO₃)₂), and siderite (FeCO₃). On the basis of these findings, further studies will be focused on developing a better understanding of the factors affecting the dissolution of Mg-bearing orthopyroxene minerals, and on exploring alternative leach reagents and conditions, with a view to developing a more effective process for the accelerated carbonation of PGM tailings.     

内蒙某钽铌尾矿回收锂云母工艺研究

内蒙某钽铌尾矿含有大量的锂云母矿物,尾矿中的脉石矿物主要为长石、石英类硅酸盐矿物,矿石中的细泥(含原生细泥和磨矿产生的次生细泥)矿物制约锂云母浮选精矿品质的提高。对含Li2O 1.02%的钽铌尾矿,采用尾矿脱泥-锂云母浮选(一次粗选、一次选扫)的工艺流程,锂云母浮选采用碳酸钠作调整剂,椰油胺+MC-2作组合捕收剂,获得锂精矿含Li2O 5.02%,达到优质锂盐级标准;锂精矿对钽铌尾矿回收率为74.82%,有效实现了尾矿中锂资源的综合回收利用。     

从选钼尾矿中回收制钾肥长石选矿试验研究

黑龙江某大型钼矿选钼尾矿K2O品位为6.91%,Na2O品位为1.79%,经过试验研究采用脱泥-浮选除杂-长石浮选-强磁选除杂的工艺流程,选钼尾矿脱泥后采用油酸钠浮选除杂,然后添加硫酸调整pH值至3.6,采用BK440作长石捕收剂浮选分离长石与石英,长石浮选精矿在15000kA/m场强下脱除磁性矿物,获得长石精矿K2O品位为11.54%, Na2O品位为2.51%,K2O回收率为47.73%;Na2O回收率为40.30%。长石精矿达到制钾肥钾长石质量标准。     

Mechanical enrichment of converted spodumene by selective sieving

Spodumene ore can be concentrated in large particle size of 1–6 mm by a dense media separation (DMS) process. The concentrate typically contains relatively large amounts of impurities due to the incomplete liberation of spodumene and the quite narrow density difference between spodumene and gangue minerals. The extraction of lithium from spodumene requires the phase transformation of spodumene from α- to β-form which takes place at 1100 °C. After heat treatment the Li₂O content of spodumene concentrate can be improved by particle size separation. In this work the technical and economical possibilities of producing first low grade concentrates by DMS method and increasing the grade afterwards by particle size separation were studied. Various DMS spodumene concentrates were mechanically enriched after heat treatment by selective sieving. Autogenous and ball mill grinding were used to increase the separation efficiency between the spodumene and the gangue minerals. The Li₂O contents of the concentrates were increased from 5.1%, 3.6%, 3.1% and 1.4% to 6.8%, 5.8%, 5.9% and 5.0%, respectively. The energy economy of the sieving method was evaluated from the perspective of increased energy consumption in the heat treatment process.     

BK414在宜春钽铌矿浮选锂云母的工业试验

BK414为北京矿冶研究总院自主开发的新型锂云母浮选药剂。针对宜春钽铌矿钽铌重选尾矿,以BK414作为捕收剂进行锂云母浮选工业试验,获得锂云母浮选指标:浮选给矿Li_2O品位1.08%,锂云母精矿Li_2O品位3.79%,Li_2O回收率63.09%。与宜春钽铌矿目前所使用的椰油胺相比,采用BK414所获得的锂云母精矿回收率提高了8.07%。BK414具有选别指标好、凝固点低、腐蚀性小等优点。     

TBP-D2EHPA协萃体系分离盐湖卤水中锂镁

针对磷酸三丁酯(TBP)-FeCl3萃取体系从镁锂比高的盐湖卤水分离锂镁的工艺中存在的反萃酸度高、再生有机相试剂消耗大等问题,筛选出添加二(2-乙基己基)磷酸(D2EHPA)的协萃体系。探究了有机相组成、Fe/Li摩尔比、萃取相比等因素对盐湖卤水中锂镁萃取分离效果的影响。在有机相组成为40%TBP-20%D2EHPA-40% 260#溶剂油、Fe/Li摩尔比为1.5、相比O/A=2的优化萃取条件下,单级Li⁺萃取率达78.56%,锂镁分离系数达46.26。使用水作为洗涤剂及反萃剂,优选了洗涤相比O/A=60及反萃相比O/A=20,单级洗涤及反萃阶段锂镁分离系数分别为46.26和22.97,取得了较好的锂镁分离效果,可为后续扩大试验与工业设计提供依据。     

A novel approach for recovery of rare earths and niobium from Bayan Obo tailings

It is difficult to economically recover rare earths (RE) and niobium (Nb) from Bayan Obo tailings by the existing metallurgical processes. In this study, a novel hydrometallurgy process was employed for separating and recovering RE and Nb from Bayan Obo tailings. Firstly, by sulfating roasting at 250 °C and subsequent leaching at 60 °C, the RE and Nb present in the polymetallic minerals can be efficiently extracted into the leach solution. Secondly, after the reduction of Ti⁴⁺ and Fe³⁺ ions (to Ti³⁺ and Fe²⁺ ions) with iron powders followed by hydrolysis at pH 2.01, the Nb can be efficiently precipitated from the leach solution. The impurities present in the precipitated product can then be removed by treating with NH₃⋅H₂O–H₂C₂O₄ system at pH 4.50. Thirdly, the RE can be efficiently precipitated at pH 7.15 from the filtrate of above hydrolysis reaction mixture. Finally, the impurities present in the crude RE can be removed by oxalate co-precipitation method. The yield of RE and Nb in this novel process reaches up to 90% and 78%, respectively. Both the Nb (60.67 wt% Nb₂O₅) and RE products (>88.65 wt% RE_xO_y) have high application value.     

Mechanochemical processing of chrysocolla with sodium sulphide

Autogenous solid state processing of chrysocolla with sodium sulphide has been carried out at different CuSiO₃–Na₂S-ratios, particle sizes and temperatures. Water soluble sodium silicate (Na₂SiO₃) and covellite (CuS) were determined by means of chemical analysis and X-ray examination as reaction products. More than 84% of chrysocolla and 50% of associated malachite were converted to covellite at 80 °C, atmospheric pressure, 10 min contact time, stochiometric CuSiO₃–Na₂S-ratio and particle sizes of <100 μm. The mechanochemical treatment using a vibrating mill increases the conversion rate of chrysocolla and malachite up to 90% and 70%, respectively. After washing of the products, the sodium silicate solution was converted into sodium sulphate (Na₂SO₄) by adding sulphuric acid, with precipitation of silica as by-product. In addition, the carbothermic reduction of crystallised sodium sulphate led to sodium sulphide necessary for solid state processing.     

Some aspects of nickel extraction from chromitiferous overburden by sulphatization roasting

Nickel bearing chromitiferous overburden of Sukinda region, Orissa, India, was pugged with sulphuric acid and subjected to roasting in two stages. Various operating parameters, such as roasting temperature, time, acid concentration and role of additives were studied. Addition of moisture, during pugging, helped in the formation of insoluble ferric hydroxy sulphate that maintains the equilibrium of SO₂/SO₃ partial pressure. The double stage roasting improved the nickel extraction over the single stage isothermal roasting of the material at 700°C. An extraction of 85% nickel with 2–3% iron could be achieved at a temperature of 700°C when roasted for 15 mins, 25 wt% sulphuric acid and 20 wt% moisture in two stages. The nickel extraction was improved a further 8–10% by the addition of alkali and alkaline earth metal salts. The role of additives during roasting was also investigated.     

铁尾矿硫酸焙烧法提取铁制备α-Fe2O3光催化剂

铁尾矿中富含铁、硅等有价元素，可作为制备功能性材料的原料。为考察硫酸焙烧法提取铁制备α-Fe2O3光催化剂的可能性，研究了焙烧过程中酸矿比、焙烧温度、焙烧时间对铁提取率的影响，得到适宜的焙烧条件为：酸矿比2∶1、焙烧温度280 ℃、焙烧时间2 h，此时铁的提取率为89.80%。焙烧熟料经浸出、过滤制得含铁的硫酸盐溶液，采用中和沉淀法制备含铁的前驱体，再经400 ℃煅烧2 h制得粒径为40~50 nm、分散性较好的α-Fe2O3光催化剂。α-Fe2O3光催化降解甲基橙时，暗反应20 min吸附率为56%，光催化120 min时的降解率可达99%，表明α-Fe2O3具有良好的光催化性能。该研究结果实现了铁尾矿中铁的综合利用，促进光催化剂的实用化。     

Oxygen-18 and Deuterium Fingerprinting of Tailings Seepage at the Sullivan Mine

Stable isotopes of water, ¹⁸O and ²H, were used to fingerprint and quantify the origin and percentage of mixing between tailings pond effluent, seepage collection waters, and background waters at the Sullivan Mill Site, which is located in Kimberley, British Columbia, Canada. The old iron and siliceous ponds are covered by a layer of permeable float rock, which allows meteoric waters to directly infiltrate without undergoing evaporation. In contrast, the east gypsum pond is covered with a layer of till and vegetation, which causes meteoric waters to pond and subsequently be subjected to evaporative processes. The different evaporative environments result in distinctly different isotopic compositions for the waters collected from these sources. The difference in surface cover was used advantageously to identify particular isotopic ratios for the different tailings sources. Water collected from the east gypsum pond is relatively enriched in the heavier isotopes (δ¹⁸O and δ²H), whereas water collected from various locations in the siliceous and old iron pond are relatively depleted in δ¹⁸O and δ²H. The various seepage collection waters were observed to plot at intermediate positions, corresponding either to the old iron pond mixing line, the gypsum pond evaporation line, or somewhere between the two.     

Purification of rare earth elements from monazite sulphuric acid leach liquor and the production of high-purity ceric oxide

The present work describes the development of an efficient and relatively simple process to obtain high grade CeO₂ from sulphuric acid leach liquor. The liquor was obtained through acid digestion of monazite. The steps investigated in the process for obtaining ceric oxide were: (i) purification of the RE elements through their precipitation as rare earth and sodium double sulphate (NaRE(SO₄)₂·xH₂O), (ii) NaRE(SO₄)₂·xH₂O conversion into RE-hydroxide (RE(OH)₃) through metathetic reaction and (iii) recovery of cerium and (iv) purification of cerium from the mixture of ceric hydroxide and manganese dioxide precipitate through dissolution of the solid with HCl and precipitation of the cerium through the addition of oxalic acid (H₂C₂O₄) or ammonium hydroxide (NH₄OH) solution. The X-ray diffraction spectra of the double sulphate obtained indicated the presence of monohydrated double sulphate. X-ray diffraction and chemical analysis indicated that the precipitation should be carried out at 70 °C and at 1.1 times the stoichiometric ratio of NaOH. An excess of 30% of KMnO₄ was necessary to separate cerium from the other RE elements. Both oxalic acid and ammonium hydroxide proved efficient in the precipitation of cerium from the mixture of Ce/Mn obtained in the cerium separation. Following purification, calcinated products were obtained, assaying between 99% and 99.5% CeO₂. The cerium recovery yield was greater than 98%.     

混盐型高锂卤水萃取分离锂

以阿根廷某高锂混盐型盐湖卤水萃硼后得到的萃余液为研究对象,采用磷酸三丁酯(TBP)- FeCl3体系萃取分离锂,考察了铁锂比、TBP浓度、稀释剂种类、萃取相比O/A、有机添加剂种类及浓度等因素的影响,优化了萃取工艺参数,分析了锂在萃取两相体系中的分配行为以及锂与镁、钠和钾的分离情况。在有机相组成为65% TBP - 5%MIBK - 30% 260#溶剂油、铁锂比1.5:1、相比O/A=2:1、萃取混合时间5 min、室温的条件下,锂的单级萃取率大于79%,镁、钠、钾的萃取率分别仅有5.24%、8.57%和0.88%,Li与Mg、Na、K的分离系数分别达到62、37和388,TBP显示出对锂的良好选择性,且MIBK的加入改善了分相和锂萃取性能。夹带与共萃进入有机相中的镁、钠和钾可用氯化锂的稀盐酸溶液洗涤去除。负载锂的有机相可用7 mol/L盐酸进行反萃,得到富锂溶液。     

Selective flotation separation of spodumene from feldspar using new mixed anionic/cationic collectors

The selectivity in flotation separation of spodumene from other pegmatic aluminosilicates such as feldspar and mica, using a single fatty acid anionic collector, is found to be very difficult. It is attributed to the same surface Al site for interaction with fatty acid. In this study, a collector mixture of sodium oleate (NaOL) and dodecyl trimethyl ammonium chloride (DTAC) demonstrates a high selectivity for the flotation of spodumene from feldspar. The influences of important factors such as molar ratio of the mixed collectors, dosages of calcium chloride, sodium hydroxide and sodium carbonate, and pulp temperature on spodumene and feldspar flotation using the mixed collectors have been investigated through micro-flotation tests. The optimum molar ratio of NaOL and DTAC is found to be 9:1. Sodium hydroxide has to be used as a pH regulator. Without the use of depressants, no selectivity is observed as the flotation of spodumene and feldspar are activated by Ca²⁺. Sodium carbonate must be used as depressant of feldspar. A comparison of the flotation for a lithium pegmatite ore using mixed fatty acid soaps and mixed anionic/cationic collectors (NaOL/DTAC) was carried out by the batch flotation tests. The results indicated that NaOL/DTAC decrease collector consumption by two-thirds. The recovery and grade of Li₂O concentrates increase by 4.93% and 0.31%, respectively.     

石膏/轻烧菱镁矿复合滤床法对铅污染水体修复研究

基于吸附共沉淀机理,将石膏和球磨菱镁矿粉复合加入处理模拟铅污染水体（含过量硫酸铅微粒的微混浊饱和硫酸铅溶液,Pb²⁺约2 mg/L,总铅10 mg/L）,研究了石膏（CaSO₄·2H₂O）、球磨菱镁矿（Mg(OH)₂和MgCO₃混合物）用量以及过滤时间对去除低浓度铅的影响。发现在最佳条件下可将水体的总铅含量降低到0.02 mg/L。在此基础上提出石膏轻烧菱镁矿复合滤床工艺,在对铅锌尾矿淋滤水和蓄电池厂排出含微量铅水进行试验中发现,可将水体中总铅降低到0.04 mg/L以下,总镉降低到低于0.1 mg/L,适合用于自来水供水及矿山尾矿库治理。     

The effect of sodium sulphate on the hydrogen reduction process of nickel laterite ore

A series of nickel laterite ores with different calculated amounts of anhydrous sodium sulphate were prepared by physical blending or sodium sulphate solution impregnation. The reduction of the prepared nickel laterite ore by H₂ was carried out in a fluidised-bed reactor with provisions for temperature and agitation control, and the magnetic separation of the reduced ore was performed using a Davis tube magnetic separator. The mineralogical properties of the raw laterite ore, reduced ore and magnetic concentrate were characterised using ICP, TG–DSC, N₂ adsorption, X-ray diffraction and optical microscopy. The catalytic activity of sodium sulphate was also studied by using Hydrogen temperature-programed reduction. The experimental results indicate that Na₂SO₄ could overcome the kinetic problems faced by the laterite ore and that it exhibited noticeable catalytic activity only if the temperature reached at least 750 °C. This high temperature accelerated the crystal phase transition of the silicate minerals and increased the utilisation of H₂. In comparing the results from the two different methods for adding Na₂SO₄, the nickel content and recovery of the magnetic concentrate were increased by using the impregnation method rather than the physical blending method and the increasing amount of sodium sulphate assisted in the further beneficiation of nickel. The partial pressure of H₂ and the reducing time also affected the reduction process of the iron oxides. The results of the microscopic study indicated that the formation of a Fe–S solid solution, which was derived from the SO₂ sulphide reduction of FeO, was conducive to mass transfer and accelerated the coalescence of metallic ferronickel particles. For the nickel laterite ore, under the synergistic effect of sodium sulphate and hydrogen, a nickel content and nickel recovery of 6.38% and 91.07% were obtained, respectively, with high product selectivity.