Sulfuric acid leaching of high iron-bearing zinc calcine

Sulfuric acid leaching of high iron-bearing zinc calcine was investigated to assess the effects of sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, and the stirring speed on the leaching rates of zinc and iron. The results showed that the sulfuric acid concentration, liquid-to-solid ratio, leaching time, and leaching temperature strongly influenced the leaching of zinc and iron, whereas stirring speed had little influence. Zinc was mainly leached and the leaching rate of iron was low when the sulfuric acid concentration was less than 100 g/L. At sulfuric acid concentrations higher than 100 g/L, the leaching rate of iron increased quickly with increasing sulfuric acid concentration. This behavior is attributed to iron-bearing minerals such as zinc ferrite in zinc calcine dissolving at high temperatures and high sulfuric acid concentrations but not at low temperatures and low sulfuric acid concentrations.

Recovery of scandium from silicate minerals by high-pressure leaching in sulfuric acid

This study investigated the leaching of scandium from a Sc-bearing silicate ore.The Sc concentrate obtained by beneficiation pre-enrichment has 1.03×10^(−4) wt%Sc and 14.66 wt%Fe.The main Sc-bearing minerals are limonite,chlorite,sericite,and hornblende,in the form of isomorphism.High-pressure leaching(HPL)with sulfuric acid(98 wt%)and NaF was used to dispose this samples.The Sc leaching efficiency of 91.82%was obtained under the optimum conditions,which are pressure of 5.0×10^(5) Pa,A/O value of 1.1,L/S value of 1.8,NaF dosage of 2 wt%,at 130℃ for 6 h.The X-ray diffraction(XRD)and scanning electron microscopy(SEM)results show that the residue is mainly quartz phase in the multihole shape.Compared with ordinary leaching for silicate,high-pressure leaching has significant advantages in saving acid and energy consumption.

Systemic treatment for severe concentrated sulfuric acid burns in an adult male at high altitude:A case report

BACKGROUND The aim of this study was to report a case of a patient who suffered from severe concentrated sulfuric acid burns while working at high altitude.This patient recovered after systemic treatment.We also provide a literature review for a better understanding of the disease.CASE SUMMARY A 30-year-old male,who was working in a local chemical plant in Xining at an altitude of 2261 m,suffered from 85%burns(Ⅲ°70%,deep Ⅱ°15%)after a tank containing 80%concentration of sulfuric acid exploded.The patient immediately received a series of first aid treatments,as well as rigorous wound managements after admission,which included protection for the whole body and organs,prevention and treatment of eye burns,and the appropriate oxygen therapy.After 65 d of treatment,the burn wounds had completely healed,and the patient was transferred to another specialized hospital for further eye treatment.The first aid before admission and the emergency treatment of wounds following admission were appropriate.No severe complications of sepsis,severe renal insufficiency,septic costal chondritis,corneal perforation or other burns occurred during the treatment.CONCLUSION The main causes of concentrated sulfuric acid burns consisted of accidental burns at work,accidents in the outside,factitious injuries and improper laboratory operations.The clinical manifestations were mostly deep Ⅱ°and Ⅲ°burns,with a formation of brown-black,leather-like eschar on the wound surface and locally embolized dendrite-like vessels.The clear cause of the injury and typical clinical manifestations in this case made it easy to diagnosis.However,adult cases with severe concentrated sulfuric acid burns in high altitude areas are rare,so the successful treatment of this case is of great significance.

Pressure acid leaching of zinc sulfide concentrate

Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.

熟化银精矿氧压强化酸浸行为研究

考虑到浮选银精矿极具回收价值及现有处理工艺存在不足,本文提出了浓硫酸熟化-氧压酸浸的工艺路线。首先,以湿法炼锌过程产出的浮选银精矿为对象,利用浓硫酸的强氧化性破坏浮选残留有机药剂矿物包裹体;其次,利用加压湿法冶金独特的技术优势,氧化溶解低价金属硫化物;最后,重点研究了工艺过程原理及氧压酸浸过程中Zn、Cu、In的浸出行为及浸出渣中物相演变规律。研究结果表明:银精矿颗粒表面残留浮选药剂的P—O、—OH极性基团被浓硫酸氧化破坏,避免产生水化膜及矿物包裹体,促进有价元素的高效浸出;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,银精矿物相组成由闪锌矿(Zn S)、锌铁尖晶石(Zn Fe_(2)O_(4))、雌黄铁矿(FeS)、氯化亚铜(Cu Cl)为主,转变为以单质硫(S8)、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主,并残留少量的闪锌矿(Zn S);适当提高反应温度和初始质量浓度可促进Zn、Cu、In的浸出,但温度过高会加剧铟铁共沉淀问题,降低锌、铟的浸出率,而质量浓度过高会降低氧气对Zn S及Fe^(2+)的氧化能力,抑制银精矿的溶解;延长反应时间、提高液固比均有利于银精矿的解离浸出;增大氧分压可以加强Fe^(3+)的氧化能力,促进对Zn S及ZnxFe(1-x)S的氧化作用,有利于银精矿的溶解。在反应温度为155℃、初始质量浓度为160 g/L、反应时间为180 min、氧分压为0.8 MPa、液固比为7 m L/g的氧压酸浸条件下,银精矿中Zn、Cu、In的浸出率分别为94.84%、98.36%和91.13%,浸出渣中铅、银的富集率分别达2.88%和7754 g/t。

锌精矿氧压浸出技术研究进展

传统湿法炼锌工艺存在烟气处理成本高、综合回收率低等问题,近年来,氧压浸出技术因其原料适应性强、锌回收率高、环境污染小等优点被应用于锌冶炼工业。介绍了锌精矿氧压浸出的生产工艺,论述了浸出过程反应的基本原理,以及离子催化、温度、酸度、分散剂等条件对浸出过程离子行为的影响与机理。讨论了锌精矿中伴生元素铁、镓、锗、铟、银等的浸出行为及在不同浸出条件下浸出行为的变化。

液相高精度吸附脱除SO_(2)新工艺在混合脱氢装置的工业应用

目的脱除作为混合脱氢装置原料的烷基化碳四(C_(4))中的SO_(2)杂质,控制SO_(2)体积分数<1×10^(-6)。方法采用液相高精度吸附脱硫工艺,通过吸附再生的方式实现其连续运行。当原料通过吸附脱除SO_(2)达到饱和后,流程上可通过氮气循环置换、升温汽提及空气再生的手段,重新恢复SO_(2)吸附剂的吸附能力,并再次投入生产使用,以确保装置的安全平稳长周期运行。结果采用液相高精度吸附脱硫工艺可将烷基化C_(4)中SO_(2)体积分数从30×10^(-6)~400×10^(-6)脱除至1×10^(-6)以下。结论对吸附脱硫工艺的吸附剂、工艺流程进行优化调整,缩短了吸附剂的再生时间,保障了装置工业化连续运行。同时,液相高精度吸附脱硫工艺流程操作简便,能耗较低,安全风险较低。

湿法消解-火焰原子吸收光谱法测定铋精矿中银含量

由于铋精矿中含有大量的硫元素和硅元素,测定铋精矿中银的含量时,王水无法将其完全消解,导致样品溶解不完全,使得测量结果偏低。本研究采取在样品前处理加入高氯酸与氟化氢铵的方法来消解硫元素和硅酸盐化合物,然后采用火焰原子吸收光谱仪(AAS)测定铋精矿中银的含量,有效解决了铋精矿中较高的硫含量和硅含量对样品溶解及测定过程中的影响。试验过程中,采用2.0 mL高氯酸、0.2 g氟化氢铵、王水体系溶解铋精矿,在体积比20%HCl介质条件下,利用火焰原子吸收光谱仪测得铋精矿样品中的银含量,结果表明,本试验方法检出限为0.0066μg/mL,低于测定下限0.022μg/mL,表明本试验方法可行,能够满足铋精矿中银含量的测定要求;利用该方法测定样品中银含量,相对标准偏差(RSD,n=7)为0.20%~1.95%,加标回收率为97.7%~105.0%;综合元素干扰测定过程中没有产生明显的基体效应干扰。该方法适用于铋精矿中50.0~10000.0 g/t银的测定,为企业和检测机构开展铋精矿中银含量的测定提供了一种更为简便、实用性更强的分析检测方法。

催化反应角度的非衡态高浓度SO_(2)转化工艺分析

非衡态高浓度SO_(2)转化工艺为了满足工艺要求的转化率,对催化剂的选型和装填量设计提出了较高要求。从催化反应和转化角度,研究了转化一段的催化剂装填量、进气量、进口烟气温度对出口烟气温度与转化率的影响,认为进口烟气温度对出口烟气温度的影响非常显著,并对催化剂选型和装填方案、进口烟气条件和温度分布提出了相应的建议。

Selective leaching of vanadium from V-Ti magnetite concentrates by pellet calcification roasting-H_(2)SO_(4) leaching process

A novel method of pellet calcification roasting-H_(2)SO_(4) leaching was proposed to efficiently separate and extract vanadium(V)from vanadium-titanium(V-Ti)magnetite concentrates.The leaching rate of V is as high as 88.98%,while the leaching rate of impurity iron is only 1.79%.Moreover,the leached pellets can be used as raw materials for blast furnace ironmaking after secondary roasting.X-ray photoelectron spectroscopy(XPS)and scanning electron microscopy with energy dispersive X-ray spectrometry(SEMEDS)analyses showed that V^(3+)was oxidized to V^(5+)after roasting at 1200℃,and V^(5+)was then leached by H_(2)SO_(4).X-ray diffraction(XRD)analyses and single factor experiment revealed a minimal amount of dissolved Fe_(2)O_(3) during H_(2)SO_(4) leaching.Therefore,a high separation degree of V and iron(Fe)from V-Ti magnetite concentrate was achieved through H_(2)SO_(4) leaching.Compared with the traditional roastingleaching process,this process can achieve a high selectivity of V and Fe,and has excellent prospects for industrial production.

H_(2)SO_(4)分解铁-FeCl_(3)浸出焙烧金精矿中金

采用硫酸分解焙烧金精矿,金从黄铁矿中解离的同时金得到了富集,可采用氯化铁溶液非氰浸出金。研究了硫酸浓度及过量系数、分解温度对铁分解率的影响,优化工艺条件为,焙烧温度180℃,反应时间90 min,硫酸过量系数1.2,在此条件下,铁分解率为92.14%,金含量从原来的51.7 g/t提高到106.1 g/t;研究了反应温度、液固比对氯化铁溶液浸出硫酸浸出渣中金的影响,优化浸出条件为,液固比1.5,80℃浸出90 min,在此条件下,金浸出率96.8%。

银精矿预氧化及有价元素深度强化浸出行为研究

针对浮选银精矿极具回收价值及现有处理工艺不足,提出浓硫酸预氧化—氧压酸浸的工艺。以湿法炼锌过程产出的浮选银精矿为研究对象,重点研究浓硫酸预氧化作用机理、预氧化银精矿微观形貌变化规律及氧压酸浸过程中Zn、Cu、In的浸出行为。研究结果表明:在浓硫酸预氧化温度为60℃、时间为15 h、酸矿质量比为0.6的条件下,浮选药剂中P—O、—OH极性集团被破坏,避免了水化作用及由其引起的矿物包裹体形成,提升了矿物溶解浸出能力;升高反应温度、延长时间和增大酸矿比均可促进浓硫酸预氧化反应的进行,但当酸矿比超过1.07时,预氧化物料呈半流体状态,黏度增大,不利于预氧化过程;在反应温度为155℃、初始酸度为160 g/L、反应时间为180 min、氧分压为0.8 MPa、液固比为7 mL/g的氧压酸浸条件下,预氧化银精矿中Zn、Cu、In的浸出率分别为94.84%、98.36%和91.13%,浸出渣中铅、银的富集率分别达2.88%和7 754 g/t,实现了Zn、Cu、In的深度浸出和铅、银的高效富集。

Recovery of titania from high titanium slag by roasting method using concentrated sulfuric acid

By means of energy-dispersive X-ray spectroscopy (EDX) and scanning electron microscope (SEM) analysis, the phase structure characteristics of high titanium slag were analyzed. Through the single factor and the orthogonal experiment methods, the effects of material particle size, mass ratio of acid to ore, roasting temperature, and roasting time on the acidolysis ratio of TiO2 during the process of roasting high titanium slag with concentrated sulfuric acid were systematically investigated. The results show that the sequence of each factor affecting the acidolysis ratio of TiO2 is: mass ratio of acid to ore, roasting time, and roasting temperature. The optimum technological conditions are obtained as mass ratio of acid to ore of 2.1, roasting temperature of 310°C, roasting time of 75min, and material particle size of 45–53μm. The acidolysis ratio of TiO2 is over 96% under the optimum conditions. The roasting process is proved to be significant in the exploitation and utilization of high titanium slag. The advantages of the proposed roasting process are of high efficiency, low power consumption, and minimum pollution.

乙炔清净废硫酸处理工艺应用

介绍了乙炔清净废硫酸经高温裂解生产硫酸的工艺流程、技术特点和运行情况。该工艺解决了浓硫酸清净乙炔工艺中废硫酸消化处理难的问题。用1t质量分数为75%的乙炔净化废硫酸可以生产0.72t质量分数为98%的成品硫酸。生产1t成品硫酸的原料及动力消耗为680元,且尾气排放中SO,质量浓度低于50 mg/m,从根本上解决了环境污染问题。

锌精矿高压釜硫酸浸出条件下的表面活性剂及其混合物:表面活性剂选择和实验室测试

通过浸出实验和对浸出产物的粒度、相成分等的后续分析研究表面活性剂及其混合物对锌精矿加压浸出效果的影响。所用表面活性剂为木质素磺酸盐(Lignosulfonate,LS)和十二烷基硫酸钠(sodium dodecyl sulfate,SDS)。通过测量表面张力研究表面活性剂对人工模拟溶液表面活性的影响。结果表明,同时添加LS和SDS具有协同效应,导致表面张力降级。在一系列加压浸出实验中,研究表面活性剂及其混合物浓度(СLS/SDS=0.2~0.8 g/L)、温度(T=120~140℃)、氧分压(pO_(2)=0.5~0.7 MPa)和时间(t=20~120 min)等参数对溶液中锌、铁、铜和铟浸出效果的影响。根据研究结果确定的锌精矿浸出的最优参数为:СLS=0.6~0.8 g/L、СSDS=0.2~0.6 g/L、pO_(2)=0.5 MPa、T=140℃和t=100~120 min。在最优条件下溶液中的锌、铟、铜和铁的浸出率分别为93%~94%、65%~66%、64%~68%和48%~49%。

高硅钴白合金的处理工艺研究

为解决高硅钴白合金中有价元素难以浸出的问题,在硫酸体系中,以氯酸钠作为氧化剂,并加入高硅钴白合金中的组成元素之一Fe3+作为氧化传质体,有效浸出了高硅钴白合金中的Co和Cu,并减少了Si的浸出。通过正交试验研究了初始Fe3+浓度、高硅钴白合金粒度和反应温度对浸出效率的影响,得到了优化工艺参数,并根据现实生产需求,确定了浸出成本更低的经济工艺参数。优化工艺参数为:起始Fe3+浓度为15 g/L,高硅钴白合金粒度为<0.075 mm,反应温度为85℃。其中,起始Fe3+浓度对浸出效率的影响最大。在优化工艺参数下浸出5.5 h,Co、Cu浸出率分别达到99.3%、98.7%。经济工艺参数为:起始Fe3+浓度为10 g/L,白合金粒度为0.075~0.150 mm,反应温度为85℃;浸出6.7 h后Co、Cu浸出率分别达到98.9%、98.5%。

某高硫微细粒金矿选冶试验研究

某高硫金矿中金矿物与黄铁矿紧密共生,金矿物和黄铁矿平均粒径分别为3.58和32.80μm,嵌布粒度较细,在实际生产中精矿品位和回收率均不理想。为进一步提高精矿品位和回收率,针对该矿特点,确定黄铁矿为浮选目标矿物,通过系统条件试验和精选试验得到了最优浮选工艺流程和参数,继而开展开路试验和闭路试验,得到的浮选指标并不理想。基于中矿中黄铁矿多是连生体的判断,对浮选中矿进行再磨处理,获得了精矿金品位17.4g/t、金回收率80.66%的满意指标。同时,结合“超细磨”、“碱浸预处理”工艺,对浮选精矿进行浸出试验,得到金浸出率77%的满意指标,为生产现场提供了具有指导意义的技术方案。

国外某渣选硫化铜精矿加温铁氧化酸浸试验研究

针对国外某渣选硫化铜精矿,采用加温铁氧化酸浸工艺回收其中铜,考察了氧化剂用量、浸出温度、浸出时间、初始硫酸浓度、液固比等因素对渣选硫化铜精矿中Cu浸出率的影响。结果表明,适宜的浸出条件为:氧化剂赤铁矿用量0.2 g/g矿、浸出温度85℃、浸出时间6 h、液固比5∶1、初始硫酸浓度200 g/L,此时铜浸出率可达97.96%。不同类型氧化剂验证试验结果表明,赤铁矿和磁铁矿在酸浸体系中均有较好的氧化性,可实现渣选硫化铜精矿中铜在中温条件下浸出,且三价铁化合物纯度越高,铜氧化浸出效果越好。

Influence of Mechanical Activation on Acid Leaching Dephosphorization of High-phosphorus Iron Ore Concentrates

High pressure roll grinding（HPRG）and ball milling were compared to investigate the influence of mechanical activation on the acid leaching dephosphorization of a high-phosphorus iron ore concentrate,which was manufactured through magnetizing roasting-magnetic separation of high-phosphorus oolitic iron ores.The results indicated that when high-phosphorus iron ore concentrates containing 54.92 mass% iron and 0.76 mass% phosphorus were directly processed through acid leaching,iron ore concentrates containing 55.74mass%iron and 0.33mass%phosphorus with an iron recovery of 84.64%and dephosphorization of 63.79% were obtained.When high-phosphorus iron ore concentrates activated by ball milling were processed by acid leaching,iron ore concentrates containing56.03mass%iron and 0.21mass% phosphorus with an iron recovery of 85.65% and dephosphorization of 77.49%were obtained.Meanwhile,when high-phosphorus iron ore concentrates activated by HPRG were processed by acid leaching,iron ore concentrates containing 58.02mass%iron and 0.10mass% phosphorus were obtained,with the iron recovery reaching 88.42% and the dephosphorization rate reaching 88.99%.Mechanistic studies demonstrated that ball milling can reduce the particle size,demonstrating aprominent reunion phenomenon.In contrast,HPRG pretreatment contributes to the formation of more cracks within the particles and selective dissociation of iron and P bearing minerals,which can provide the favorable kinetic conditions to accelerate the solid-liquid reaction rate.As such,the crystal structure is destroyed and the surface energy of mineral particles is strengthened by mechanical activation,further strengthening the dephosphorization.

硫化锌精矿氧压酸浸渣中综合回收硫的工艺研究

氧压酸浸渣是硫化锌精矿冶炼过程的副产物,其中硫仍具有回收利用价值。某硫化锌精矿氧压酸浸渣中,粒度小于0.038 mm的占比为67.65%,硫含量为36.27%。本文采用絮凝-浮选工艺,高效回收硫化锌精矿氧压酸浸渣中的硫。其间以O-异丙基-N-乙基氨基甲酸酯(Z-200)为捕收剂,以聚丙烯酰胺为絮凝剂,进行闭路浮选试验。试验结果表明,经一次空白浮选、两次加药和两次空白清洗,硫的回收率可达94.59%,硫精矿品位可达87.92%。作为浮选尾渣,滤饼含有硫化物,返回浸出工序进行锌回收。