有 色 冶 金 专 业 英 语

（适用于冶金工程专业）

2009年9月

Lesson 3 Ore Dressing

Ore dressing 选矿

Concentrate v. 富积，浓缩，集聚

n. 精矿，浓缩物

Concentration n. 集中，浓缩，浓度

Acid concentration 酸浓度

Bulk n. 正体，主体，团块

Gangue n. 脉石，尾矿，矿脉中的夹杂物

Tailing n. 尾矿

Severance n. 分离，隔离，碎散

Beneficiation n. 分选

Comminution n. 粉碎

Run-of-mine n. 原矿

Middling n. 中矿

Liberation n. 解离

Crush n. v. 粉碎，碾碎，挤压

Grind n. v. 研磨，磨细

Screen n. v. 筛，筛分

Jigging n. 跳选，跳汰选

Hand picking 手选

Luster n. 光泽，光亮 v. 闪光，发光

Specific gravity 比重

Magnetic permeability 磁导率

Inductive charging 感应电荷

Electrostatic separation 静电分离

Fracture n. 断口，裂缝

Automatic sorting of radioactive natures

放射性自动选矿

Magnitude n. 大小，尺寸，量级，强度，等级

Magnetic separation 磁选

Magnetic field 磁场

Gravity concentration 重力选矿

Medium n. 介质，媒介，中间物，培养基

Dilate v. （使）膨胀，扩张，扩大

Dilated bed 松散床层

Dilation n. 膨胀系数，传播，伸缩，蔓延

Lip n. 凸出部分，唇部

Diverse adj. 不同的，互异的，各种各样的

Table n. 摇床，淘汰盘

Tabling 摇床选，淘汰选

Motion n. 运动，输送，行程，机械装置，运动机构

Sink-float separation 重介质分选

Suspension n. 悬浮物，悬浮液

Cone n. 圆锥体，锥形漏斗，圆锥破碎机

Stir n. v. 移动，摇动，搅拌

Stirrer n. 搅拌器，搅拌机

Rotary adj. 旋转的，回转的，转动的

Circumference n. 圆周，周边

Rotating motion 旋转装置，旋转设备

Floatation n. 浮选

Pulp n. 矿浆，浆料 v. 制浆，浆化

Sluice n. 槽，排水道，水槽

Froth floatation 泡沫浮选

Hematite n. 赤铁矿

Pyrolusite n. 软锰矿

Diamond n. 金刚石

Graphite n. 石墨

Ore dressing concerns with the technology of treatment of ores to concentrate their valuable constituents (minerals) into products (concentrate) of smaller bulk, and simultaneously to collect the worthless material (gangue) into discardable waste (tailing). The fundamental operations of ore-dressing processes are the breaking apart of the associated constituents of the ore by mechanical means (severance) and the separation of the severed components (beneficiation) into concentrate and tailing, using mechanical or physical methods which do not effect substantial chemical changes.1

Severance. Comminution is a single, or multistage processes whereby ore is reduced from run-of-mine size to that size needed by the beneficiation processes. The process is intended to detailed control, a class of particles containing both mineral and gangue (middling particles) are also produced. The smaller the percentage of middling the greater the degree of liberation. Comminution is divided into crushing (down to 6-to 14-mush) and grinding (down to micron size). Crushing is usually done in three stages: coarse crushing from run-of-mine size to 4-6 in., or coarser; intermediate crushing down to about 1/2 in.; and fine crushing to 1/4 in. or less. Screen is a method of sizing whereby graded products are produced, the individual particles in each grade being of nearly the same size. In beneficiation, screening is practiced for two reasons: as and integral part of the separate on process, for example, in jigging, and to produce a feed of such size range as is compatible with the applicability of the separation process.

Beneficiation. This step consists of two fundamental operations: the determination that an individual particle is either a mineral or a gangue particle (selection); and the movement of selected particles via different paths (separation) into the concentrate and tailing products.2 When middling particles occur, they will either be selected according to their mineral content and then caused to report as concentrate or tailing, or be separated as a third product (middling).3 In the latter case, the middling is reground to achieve further liberation, and the product is fed back into the stream of material being treated.

Selections based upon some physical or chemical property in which the mineral and gangue particles differ in kind or degree or both. Thus in picking, the old form of beneficiation, color, luster, and shape are used to decide whether a lump of ore is predominantly mineral or gangue. Use is made of differences in other physical or chemical properties, such as specific gravity, magnetic permeability, inductive charging (electrostatic separation), surface chemical properties, bulk chemical properties, weak planes of fracture (separation by screening), and gamma-ray emission (automatic sorting of radioactive nature).

Separation is achieved by subjecting each particle of the mixture to a set of forces that is usually the same irrespective of the nature of the particles excepting for the force based upon the discriminating property. This force may be present for both mineral and gangue particles but differing in magnitude, or it may be present for one type of particle and absent for the other. As a result of this difference separation is possible, and the particles are collected as concentrate or tailing.

Magnetic separation utilizes the force exerted by a magnetic field upon magnetic materials to counteract partially or wholly the effect of gravity. Thus under the action of these two forces different paths are produced for the magnetic and nonmagnetic particles.

Gravity concentration is based on a discriminating force, the magnitude of which varies with specific gravity. The other force that is usually operating in gravity methods is the resistance to relative motion exerted upon the particles by the fluid or semi-fluid medium in which separation takes place

Jigging is a gravity method that separates mineral from gangue particle by utilizing an effective difference in settling rate through a periodically dilated bed. During the dilation heavier particles work their way to the bottom while the lighter particles remain on top and are discharged over the lip. Jigging is practiced on materials that are liberated upon being reduced to sizes ranging from 3/2 in., down to several millimeters. It has been used on such diverse ores as coal, iron ores, gold and lead ores.

Tabling is a gravity method in which the feed, introduced onto an inclined plane and reciprocated deck, moves in the direction of motion while simultaneously being washed by a water film which moves it also at right angles to the motion of the deck.4 The heavier mineral and the lighter gangue are usually collected over the edges of the deck. The boundary between the heavier mineral and lighter gangue particles is roughly a linear diagonal band on the deck of the table. This diagonal band is not stationary; rather it tends to move about a mean position. In practice therefore, a third product, the middling, is collected between the discharge edges of concentrate and gangue. If the feed to the table has been crushed or ground to produce liberation, then the middling is returned to the feed. If liberation has not been achieved, the middling is returned to the crushing-grinding section of the mill. Tables may be used to treat relatively coarse material (sand tables) with sizes ranging from about 2~3 mm down to 0.07 mm.

Sink-float separation is the simplest gravity method and is based on existing differences in specific gravity. The feed particles are introduced into a suspension, the specific gravity of which is between that of the mineral and gangue particles, with the result that particles of higher specific gravity sink while those of lower specific gravity float.5 The separator is a cone equipped with a slowly operated stirrer which serves to impart slow rotary motion to the suspension and prevent the suspension from settling out on the walls. Feed is introduced at one point of the circumference and is slowly moved by the rotating motion of the suspension. By the time this material has reached the discharge point on the circumference, those particles whose specific gravity is greater than that of the suspension have moved down through the suspension so that only float particles are discharged at the top, the sink particles are discharged at the bottom.

Flotation is used to separate valuable minerals from waste rock or gangue, in which the ground ore is suspended in water and, after chemical treatment, subjected to bubbles of air. The minerals that are to be floated attach to the air bubbles, rise through the suspension, and are removed with the froth that forms on top of the pulp. Froth flotation was first used to recover sulfide minerals that were too fine to be recovered by gravity concentrators such as jigs, tables, and sluices. Froth flotation is also used to concentrate oxide minerals such as hematite (Fe2O3) and pyrolusite (MnO2), and native elements such as sulfur, silver, gold, copper and carbon (both graphite and diamond). Froth flotation is also used to separate the silicate minerals.

Lesson 5 Materials Science and Engineering

Embrace 包括

Ceramics 陶瓷

Inanimate 无生命的

Homogeneous 均匀的

Predominate 主导

Rigidity 刚性

Weldability 可焊性

Composite 复合材料

Spectrum 种类

Brass 黄铜

Bronze 青铜

Invar 因钢(NiFe)

Cement 水泥

Ferrite 铁素体

Garnet 石榴石

PVC 聚氯乙烯

Polyethylene 聚乙烯

PTFE 聚四氟乙烯

Terylene 涤纶

nylon 尼龙

leather 皮革

reinforced 增强

dispersion 弥散

supersonic 超声波

optimum 最优

fabrication 人工制作

invariable 不变的

corrosion 腐蚀

fatigue 疲劳

assess 评估

1.    Materials Science

“Materials Science” is a subject for engineers of the modern age. It embraces a study of different materials regarding their structures, properties and uses. The “material” here does not refer to all matter in the Universe. If this were so, it would include all the physical sciences and the life sciences form astronomy to zoology. We can restrict the definition only to matter useful to mankind. Even here, the range is too broad for the purposes of the engineer. For example, we can list a large number of things useful, to man, such as food, medicines, explosives, chemicals, water, steel, plastics and concrete, only a few of which qualify as engineering materials. We have then to be more specific, and define materials as that part of inanimate matter that is useful to the engineer in the practice of his profession.1 Recently the term, materials refer only to solid materials, even though it is possible to quote a number of examples of liquid and gaseous materials such as sulfuric acid and steam, which are useful to the engineer.

The word ‘science’ refers to the physical science, in particular to chemistry and physics. As we confine ourselves mainly to solid in material science, the subject is related to solid state chemistry and solid state physics. The engineering usefulness of the matter under study is always deep in mind. In this respect, material ceramics science comes heavily from the engineering sciences such as metallurgy, and polymer science. These, in their own time, have grown out of their interaction with the basic sciences of chemistry and physics.

Therefore, Material Science refers to that branch of applied science concerned with investigating the relationship existing between the structure of materials and their properties, and it concerns with the interdisciplinary study of materials for entirely practical purposes.2 Material science has developed rapidly during the last ten years. The new approach of material science has paid of handsomely in many ways and they have solved the problems in selection of right materials in complex situations.

2. Classes of Engineering Materials

Within the scope of material science, the engineering materials may be classified in three broad groups according to their mode of occurrence:

(1)    Metals and alloys

(2)    Ceramics

(3)    Organic polymers.

A metal is an elemental substance. An alloy is a homogeneous mixture of two or more metals or a metal and nonmetal. Among the solid materials, metals and alloys predominate because of their useful characteristics of hardness, strength, rigidity, formability, machinability, weldability, conductivity and dimensional stability.

Ceramics are materials consisting of phases. A phase is a physically separable and chemically homogeneous constituent of a material. These are themselves compounds of metallic and non-metallic elements. All metallic compounds, rocks minerals, glass, glass-fiber, abrasives and all fired clays are ceramics.

Organic materials are those materials derived directly from carbon. They usually consist of carbon chemically combined with hydrogen, oxygen or other nonmetallic substances, and their structures are, in many instances, fairly complex. Plastics and synthetic rubbers are common organic engineering materials.

Table 1 shows a broad spectrum of engineering materials which shows not only typical examples from each of these three groups but also gives a number of examples of materials which are composite up of two groups.3 In general, in each and every engineering application we find material from all the three basic types of materials described above.

Table 1. Some important grouping of materials

Since the engineer must specify the materials for TV sets, computers, suspension bridges, oil refineries, rocket motors, nuclear reactors, or supersonic transports he must have sufficient knowledge to select the optimum material for each application. Although experience provides the engineer with a starting point for selection of materials, the skill of the engineer will be limited unless he understands the factors that contribute to the properties of materials.4

3. Selection of Materials

Right type of material is to be selected for a particular type of work. The selections of the right materials for given requirements, the proper use of those materials, development of new ways of using them for greater effectiveness, all are direct responsibility of the engineer.

To fulfill this responsibility, the engineer must have a thorough knowledge of the nature and behavior of materials. The study of the nature of materials has its foundation in chemistry and physics and that of behavior of materials involves the application, of the principles of the nature of materials, under the varied conditions found in engineering practice.3 This behavior of materials is determined by composition, structure, service conditions, and the interactions among them. All materials have limitations within which they perform well but beyond which they cannot be used satisfactorily.

However, the selection of a material for a specific application is invariably a thorough, lengthy, and expensive investigation. Almost always more than one material is suited to the application, and the final selection is a compromise that weights the relative advantages and disadvantages. The varied requirements to three broad demands: (1)  Service requirements; (2)  Fabrication requirements; (3) Economic requirements.

The service requirements have important role in material selection. The material must stand up to service demands. Such demands commonly include dimensional stability, corrosion resistance, adequate strength, hardness, and toughness, heat resistance. In addition to any such basic requirements, other properties may be required such as a low electrical resistance, high or low heat conductivity, fatigue resistance, or others.

Fabrication requirements are also to be considered in material selection. It must be possible to shape the material, and to join it to other material. The assessment of fabrication requirements concerns questions of machinability, hardenability, heat treatability, ductility, castability, and weldability, qualities that are sometime quite difficult to assess.

Along with the above two requirements, the economic requirements give final shape in material selection. Goods must be produced at lower cost. The object is the minimum over all cost of the component to be made, and this objective is sometimes attained only by increasing one or more of the cost components

Lesson 6 Metallurgy

Metallurgy n. 冶金，冶金学

Non-ferrous metallurgy 有色冶金学

Chlorine metallurgy 氯化冶金学

Powder metallurgy 粉末冶金学

Extractive metallurgy 提取冶金学

Meteoric iron 陨铁

Craftsmanship n. 手艺，技能

Craftsman n. 技工，工匠

Ornamental adj. 装饰用的，观赏的

n. 装饰品

Metalworking n. 金属加工

Ceremonial adj. 正式的，礼仪的，仪式的

Decorative adj. 装饰的

Decorative arts 装饰艺术

Cast n. v. 铸造，铸件

Process metallurgy 过程冶金

Production metallurgy 生产冶金

Physical metallurgy 物理冶金

Chemical metallurgy 化学冶金

Mechanical metallurgy 机械冶金，力学冶金

Unit operation 单元操作

Unit process 单元过程

Flux n. 熔剂

Solvent n. 溶剂

Slag n. 渣，炉渣 v. 造渣

Electrolyte n. 电解质，电解液

Depletion n. 用尽，消耗，贫化，提取金属

Deposit n. v. 沉积，沉淀，电积

Blast furnace 鼓风炉，高炉

Crude iron 生铁

crystal structure 晶体结构

neutron n. 中子

diffraction n. 衍射

crystal imperfection 晶体缺陷

plastic deformation 塑性变形

metallography n. 金相学

microscopy n. 显微镜学，显微技术

forging n. 锻造，锻件

blowhole n. 气孔

thermodynamics n. 热力学

kinetics n. 动力学

Steelmaking n. 炼钢

Scrape n. 废料

Leach v. 浸出，溶出

Electrochemical reduction cell 电化学还原电池

Inorganic chemistry 无机化学

Pyro-metallurgy 火法冶金

Hydro-metallurgy 湿法冶金

elevated temperature 高温

reduce v. 还原

reduction n. 还原

charcoal n. 木炭，炭

spontaneous adj. 天然的，自动的，自发的

residue n. 残渣，剩余物，残余物，炉渣

roasting n. 焙烧

pig iron 粗铁，生铁

refine v. n. 精炼，提纯，纯化

uranium n. 铀

tungsten n. 钨

molybdenum n. 钼

isolate v. 隔离，隔绝，切断

recovery n. 回收，回收率，回复，恢复

scope n. 范围，领域，目标

revert n. 返料

metalloid n. 类金属 adj. 类金属的

selenium n. 硒

tellurium n. 碲

amenability n. 可控制性，可处理性

adaptability n. 适应性

hafnium n. 铪

zirconium n. 锆

flexibility n. 适应性，灵活性

Metallurgy is the science of metallic materials. Metallurgy as a branch of engineering is concerned with the production of metals and alloys, their adaptation to use, and their performance in service. As a science, metallurgy is concerned with the chemical reactions involved in the processes by which metals are produced and the chemical, physical, and mechanical behavior of metallic materials.1

Metallurgy has played an important role in the history of civilization. Metals were first produced more than 6000 year age. Because only a few metals, principally gold, silver, copper and meteoric iron, occur in the uncombined state in nature, and then only in small quantities, primitive metallurgists had to discover ways of extracting metals from their ores. Fairly large-scale production of some metals was carried out with technical competence in early Near Eastern and Mediterranean civilizations and in the Middle Ages in central and northern Europe. Basic metallurgical skills were also developed in other parts of the world.

The winning of metals would have been of little value without the ability to work them. Great craftsmanship in metalworking developed in early times; the objects produced included jewelry, large ornamental and ceremonial objects, tools and weapons. It may be noted that almost all early materials and techniques that later had important useful applications were discovered and first used in the decorative arts.2 In the Middle Ages metalworking was in the hands of individual or groups of craftsmen. The scale and capabilities of metalworking developed with the growth of industrial organizations. Today’s metallurgical plants supply metals and alloys to the manufacturing and construction industries in many forms such as beams, plates, sheets, bars, wire, and castings. Rapidly developing technologies such as communications, nuclear power, and space exploration continue to demand new techniques of metal production and processing.

The field of metallurgy may be divided into process metallurgy, (production metallurgy, extractive metallurgy) and physical metallurgy. According to another system of classification, metallurgy comprises chemical metallurgy, mechanical metallurgy (metal processing and mechanical behavior in service), and physical metallurgy. The more common division into process metallurgy and physical metallurgy, which is adopted here, classifies metal processing as a part of process metallurgy and the mechanical behavior of metals as a part of physical metallurgy.

Process metallurgy Process metallurgy, the science and technology used in the productions of metals, employs some of the unit operations and unit processes as chemical engineering. These operations and processes are carried out with ores, concentrates, scrap metals, fuels, fluxes, slag, solvents, and electrolytes. Different metal adopts different combinations of operations and processes, but typically the production of a metal involves two major steps. The first is the production of an impure metal from ore minerals, commonly oxides or sulfides, and the second is the refining of the reduced impure metal, for example, by selective oxidation of impurities or by electrolysis. Process metallurgy is continually challenged by the demand for metals that have not been produced previously or are difficult to produce; by the depletion of the richer and more easily processed ores of the traditional metals; and by the need for metals of greater purity and higher quality. The mining of leaner ores has greatly enhanced the importance of ore dressing methods for enriching raw materials for metal production. Several nonferrous metals are commonly produced from concentrates. Iron ores are also increasingly treated by ore dressing.

Process metallurgy today mainly involves large scale operations. A single blast furnace produces crude iron at the rate of 3,00~11,000 tons per day. A basic oxygen furnace for steelmaking consumes 800 tons of pure oxygen together with required amounts of crude iron and scrap to produce 12,000 tons of steel per day. Advanced methods of process analysis and control are now being applied to such processing system. The application of vacuum to extraction and refining processes, the leaching of low-grade ores for the extraction of metals, the use of electrochemical reduction cells, and the refining of reactive metals by processing through the vapor state are other important developments.

Because the production of metals employs many different chemical reactions, process metallurgy has been closely associated with inorganic chemistry. Techniques for analyzing ores and metallurgical products originated several centuries ago and represented an early stage of analytical chemistry. Application of physical chemistry to equilibrium and kinetics of metallurgical reactions has led to great progress in metallurgical chemistry.

According to temperature at which the process is carried out process metallurgy may be divided into pyrometallurgy and hydrometallurgy. Pyrometallurgy is processes employing chemical reactions at elevated temperatures for the extractions of metals from ores and concentrates. The use of heat to cause reduction of copper ores by charcoal dates from before 3,000 B.C. The techniques of pyrometallurgy have been gradually perfected as knowledge of chemistry has grown and as sources of controlled heating and materials of construction for use at high temperature have become available.3 Pyrometallurgy is the principal means of metal production.

The advantages of high temperature for metallurgical processing are several: chemical reaction rates are rapid, reaction equilibriums change so that processes impossible at low temperature become spontaneous at higher temperature, and production of the metal as liquid or gas facilitates physical separation of metal from residue.4

The processes of pyrometallurgy may be divided into preparation processes which convert the raw material to a form suitable for further processing (for example, roasting to convert sulfides to oxides), reduction processes which reduce metallic compounds to metal (the blast furnace which reduces iron oxide to pig iron), and refining processes which remove impurities from crude metal (fractional distillation to remove iron, lead, and cadmium from crude zinc).

The complete production scheme, from ore to refined metal, may employ pyrometallurgical processes (steel, lead, tin, zinc), or only the primary extraction processes may be pyrometallurgical, with other methods used for refining (copper, nickel). 5 In some case (uranium, tungsten, molybdenum), isolated pyrometallurgical processes are used in a treatment scheme that is predominately nonpyrometallurgical.

Hydrometallurgy is the extraction and recovery of metals from their ores by processes in which aqueous solutions play predominant role. Two distinct processes are involved in hydrometallurgy; putting the metal values in the ore into solution via the operation known as leaching; and recovering the metal values from solution, usually after a suitable solution purification or concentration step, or both. The scope of hydrometallurgy is quite broad and extends beyond the processing of ores to the treatment of metal concentrates, metal scrap and revert materials, and intermediate products in metallurgical processes. Hydrometallurgy enters into the production of practically all nonferrous metals and or metalloids, such as selenium and tellurium.

The advantages of hydrometallurgy are applicability to low-grade ores (copper, uranium, gold, silver), amenability to the treatment of materials of quite different compositions and concentrations, adaptability to separation of highly similar materials (hafnium from zirconium), flexibility in terms of the scale of operations, simplified materials handling as compared with pyrometallurgy, and good operational and environmental control.

Physical metallurgy investigates the effects of composition and treatment on the structure of metal and the relations of the structure to the properties of metals. Physical metallurgy is also concerned with the engineering applications of scientific principles to the fabrication, mechanical treatment, heat treatment, and service behavior of metals.

The structure of metals consists of their crystal structure, which is investigated by x-ray, electron, and neutron diffraction, their microstructure, which is the subject or metallography, and their macrostructure. Crystal imperfections provide mechanisms for processes occurring in solid metals, for example, the movement of dislocations results in plastic deformation. Crystal imperfections are investigated by x-ray diffraction and metallographic methods, especially electron microscopy. The microstructure is determined by the constituent phases and the geometrical arrangement of the microcrystals (grains) formed by those phases. Macrostructure is important in industrial metals. Phase transformations occurring in the solid state underlie many heat-treatment operations. The thermodynamics and kinetics of these transformations are a major concern of physical metallurgy. Physical metallurgy also investigates changes in the structure and properties resulting from mechanical working of metals.

Lesson 12 Calcination and Roasting

Calcination n. 焙烧，煅烧

calcine 焙砂

Decomposition n. 分解，裂解

Metal hydrate 金属氢氧化物

Carbonate n. 碳酸盐

Basic sulphate 碱式硫酸盐

Rotary kiln 回转窑

Shaft furnace 竖炉

Dead roasting 死烧

Sulphating roasting 硫酸化焙烧

Reduction roasting 还原焙烧

equillibrium constant 平衡常数

kellog diagram 凯洛格相图

predominance n. 优势，优越

predominance area 优势区

partial roasting 部分焙烧

selective roasting 选择性焙烧

chloridizing roast 氯化焙烧

smelt n. v. 熔炼

noble adj. 贵重的，惰性的

noble metal 惰性金属，贵金属

hypothetical adj. 假定的，有前提的

fume n. 烟气

halide n. 卤化物

volatilizing roast 挥发焙烧

magnetizing roast 磁化焙烧

magnetite n. 磁铁矿

flash roaster 闪速焙烧炉,飘悬焙烧炉

inject v. 喷射，喷入

fluidise v. 流态化

fluidized bed roaster 流态化焙烧炉

burner n. 喷嘴

suspend v. 悬浮，漂浮

fluo-solids roaster 流化-闪速焙烧炉

matte n. 冰铜，锍

reverberatory furnace 反射炉

1. Calcination

Calcination involves the chemical decomposition of the mineral and is achieved by heating to above the mineral’s decomposition temperature (TD) or by reducing the partial pressure of the gaseous product (PH2O, PCO2) below that of its equilibrium partial pressure for a certain constant temperature.1 For example,

CaCO3 = CaO + CO2

TD = 900℃ (under standard thermodynamic conditions)

Calcination is mainly used to remove water, CO2 and other gases which are chemically bound in metal hydrate and carbonates as these minerals have relatively low decomposition temperatures.2

Calcinations are conducted in rotary kilns, shaft furnaces or fluidized bed furnaces.

2. Roasting of metal concentrates

The most important roasting reactions are those concerning metal sulfide concentrates and involve chemical combination with the roasting atmosphere.

Possible reactions include:

MS + 3O2 = 2MO + 2SO2 (dead roast)

MS + 2O2 = MSO4 (sulfating roast)

MS + O2 = M + SO2 (reduction roast)

Other equilibria which need to be taken into account include:

(1/2)S2 + O2 = SO2 and

SO2 + (1/2)O2 = SO3

Thus, when PSO2 is large and PS2 become large. Also when PO2 and PSO2 become large, PSO3 become large which is the required condition for sulfating roasting; the sequence of reactions being

MS + (3/2)O2 = MO + SO2

SO2 + (1/2)O2 = SO3

MO + SO3 = MSO4

Fig. 12-1 Hypothetical thermodynamic phase diagram for the roasting

of a metal sulfide concentrate at a constant temperature

giving the overall sulfating reaction:

MS + 2O2 = MSO4

If the metal forms several sulfides, oxides, sulfates and basic sulfates, e.g. M2S, M2O3, M2(SO4)3, MSO4, XMO3 further equilibria must be considered. By examination of the equilibrium constant for each of these roasting reactions it is possible to determine the values of PO2 and PSO2 at which each of the roasting products (calcine) is in equilibrium with the metal sulfide at a constant temperature. Increasing or decreasing the PO2 or PSO2 may produce other roasting reactions. Kellog has used this criterion to construct thermodynamic phase diagrams for the roasting reactions at a constant temperature (Fig. 12-1). Reactions which involve both SO2 and O2 are seen to have a diagonal line since the equilibrium phases produced will depend on the partial pressure of both gases. The roasting of a metal sulfide to a sulfate or a basic sulfate will produce a vertical reaction line since only O2 will take part in the reaction. The ‘Kellog diagram’ provides the predominance areas for each phase within which PO2 and PSO2 can be varied without altering the roasting product or calcine. It should be noted that if roasting is carried out in air the sum of the partial pressure of O2 and SO2 is about 0.2 atm, i.e. PO2+ PSO2 = 0.2 atm..

The ore concentrate will undoubtedly contain other metal sulfide each with their own phase predominance areas dependent upon PO2 , PSO2 and temperature. By examination of the appropriate diagrams it is possible to obtain information which will enable the roasting operator to select the most appropriate roasting conditions for each particular concentrate. In this way selective roasting of one metal sulfide to its oxide may be achieved while another metal present in the concentrate remains as the sulfide.

A dead roast is used when the metal oxide is to be reduced by carbon or hydrogen. A sulfating roast is used when the metal sulfate is subsequently leached with a dilute sulfuric acid solution. Metal sulfates decompose at low temperatures, therefore sulfating is normally conducted at about 600~800℃, i.e. below the corresponding decomposition temperature, with a restricted amount of air, while dead roasting is conducted at 800~900℃ with excess air, i.e. a high PO2 / PSO2 ratio.

Differential sulfating roasting is possible by operating at a temperature which will decompose one sulfate but not another.4 Thus, roasting a Ni3S2—FeS concentrate at 850℃ will produce NiSO4 and Fe2O3 since Fe2(SO4)3 will decompose at this temperature, i.e. Fe2(SO4)3 =Fe2O3 + 3SO3. Reduction roasts are generally rare since this reaction usually requires very low PO2 values and high temperatures is demanded by the thermodynamic considerations.

Other roasting reactions include:

⑴ Chloridizing roast which is generally used for the conversion of a reactive metal such as Ti, Zr, U, which form extremely stable oxides, to a less stable chloride or other halide. The halide is relatively easy to reduce with another element which forms more stable halide.

⑵ Volatilizing roasts remove volatile impurity elements and oxides such as Cd, As2O3, Sb2O3, ZnO. These may be recovered from the process fume using bag filters.

⑶ Magnetizing roast using controlled reduction of hematite (Fe2O3) to magnetite (Fe3O4) which can be subsequently magnetically separated from the gangue.

The roasting reactions are gas—solid reactions and therefore rely on the diffusion of oxygen into and sulfur dioxide out of each concentrate particle.

Reducing the particle size (less than 6 mm) improves the gas—solid contact and increases throughput. This principle is used in the modern flash roasters in which preheated ore particles are injected through a burner with air, and fluidized bed roasters in which the fine ore particles are suspended in the roasting gas. A development which incorporates both these principles is the fluo—solids roaster in which air and ore fines are injected into the side of a reactor and fluidized by an upward draught of preheated (500℃) air.

It has been claimed that the fluo—solids roaster offers certain advantages over the multihearth unit for the partial roasting of Cu2O ores. Such advantages include greater control over sulfur elimination, less space and operating labor required, it is more amenable to automation and a higher quality matte from the reverberatory furnace is produced.5

Lesson 13 Mattes

Ionic adj. 离子的

Immiscible adj. 不能混合的，互不溶解的

Straightforward smelting 直接熔炼，连续熔炼

Soda ash 纯碱，碳酸钠

Gypsum n. 石膏

Ternary adj. 三元的 Eutectic structure 共晶结构

Ternary eutectic 三元共晶

Deviation n. 偏差，偏离

positive deviation 正偏差 negative deviation 负偏差

Raoult’s law 拉乌尔定律

Pseudo 假，伪，准，似

Terminal adj. n. 极限（的），终点（的），终端（的）

Solidify v. 固化、凝固，变硬

Feature n. 特征，性能

Partition n. v. 分配，分布，分隔，隔板

Availability n. 可得到，存在，现有

Balance n. 平衡，均衡

Antimony n. 锑

Preferential oxidation 优先氧化

Convert v. 吹炼

1. Mattes

Mattes are solution of metallic sulfides. They have electrical conductivities suggesting that their structures, when molten, are ionic or possibly partly metallic.1 They have rather lower melting—point ranges than slags. They are much denser than slags (specific gravity about 5 for matte and 3 for slag) and they are immiscible in both slag and metal phases, though either can dissolve sulfur. Mattes are used either for the collection of the valuable mineral in a straightforward smelting process (Cu, Ni) or for the collection of impurities in a sulfide phase from which the principal metal (e.g. Sb) has been displaced by Na, charged as soda ash (Na2CO3). Oxide ores can be smelted to matte with pyrites or gypsum as the source of sulfur but this is rare.

The copper mattes, which are by far the most important commercially, have been well discussed by Ruddle. Referring to the Cu—Fe—S ternary diagram (Fig. 13-1) the copper—iron mattes are found in a narrow band of composition running between the FeS and Cu2S compositions. Cu2S is immiscible in Cu but FeS is soluble in iron, the solution having, however, very marked positive deviation from Raoult’s law so that it is not surprising that a wide band of immiscibility separates the Fe—Cu edge of the diagram from the FeS—Cu2S band in which the mattes lie.2

Within this band several workers have detected an eutectic structure, but the actual eutectic point lies off the FeS—Cu2S join to the side of rather higher sulfur content and it is in fact a ternary eutectic.

The activities along the pseudo—binary join between Cu2S and FeS have been determined. Both species show small deviations from Raoult’s law, negative in the case of Cu2S and positive in the case of FeS, but this latter observation has been disputed. Most commercial mattes, however, contain rather less sulfur than that required to provide a mixture of the two terminal sulfides and the true eutectic would be obtained only under a sulfur pressure that is rather higher than atmospheric.3

Mattes usually contain some oxygen and experimental work on mattes should be carried out under controlled oxygen pressure as well as sulphur pressure. Analysis for oxygen in mattes is not easy as it is essentially a determination of the state of combination of iron in a sample which is usually rather difficult to dissolve without using an oxidizing agent.

In low—grade mattes made under reducing conditions in blast furnace iron may be present as the metal with FeS and Cu2S. High—grade commercial matte may contain FeS•Cu and Cu2S if made under reducing conditions. Reverberatory mattes made under an oxidizing atmosphere and slag may solidify with up to 20% of Fe3O4 and a small amount of FeO but the state of the iron in the melt is not apparently know.4 The solubility of Fe3O4 in copper—iron mattes may be about 30% (12% oxygen) in very lean matte (i.e. one which is high in FeS and low in Cu2S) but decreases rapidly as the copper content increase. The precipitation of magnetite in reverberatory hearths and in converters as iron is oxidized out of the matte and slgged (so increasing the Cu content and lowering the solubility of the Fe3O4) is a well—known feature of these processes which gives a great deal of trouble.

The copper content (“grade”) of a copper matte depends on the Cu/Fe ratio and on the O/S ratio of the concentrate of calcine being smelted. All of the copper present goes into the matte but the iron is partitioned between matte and slag in proportions which are determined by the availability of sulfur and oxygen. Sufficient iron and sulfur must be present to provide enough FeS to protect the copper in the matte from oxidation. Low—grade matte can have as little as 20% Cu (25% Cu2S) but normally the grade is higher with about 45% Cu (55% Cu2S), the balance being mainly FeS with iron oxide in solution.5

Mattes also collect a number of other metals besides iron and copper, particularly nickel, cobalt, zinc and lead, gold, silver and platinum metals. Nickel—copper—iron mattes are produced from nickel ores and are converted into copper—nickel mattes during nickel extraction, i.e. FeS is blown out, Cu2S and Ni3S2 form an eutectic which can be produced coarse enough for the constituents to be separated by flotation after crushing.

Mattes are important only in the extraction of copper, nickel and sometimes antimony. In lead smelting copper is collected in a matte in the hearth, a controlled quantity of sulfur being included in the charge for the purpose. Sulfur is partitioned between the matte and and any metallic phase present with a small amount also entering the slag. Oxygen and iron are partitioned between the matte and any slag in the system. Mattes are intermediate products and the distinct lack of knowledge about them is partly due to this fact, partly to the difficulty of investigating them, and partly to the fact that in the next stage of their processing, conversion, the problems encountered are not primarily connected with matte constitution and structure.6

The product of matte smelting provides the metal to be extracted in the form of a concentrated metal sulfide which needs to be converted to the metal.

Converting of the sulfide matte is achieved by blowing air and/or oxygen through the liquid matte effecting preferential oxidation of the more reactive impurity metal sulfides, e.g. MeS to their respective oxides which are collected in an appropriate slag:

2MeS + 3O2 = 2MeO + 2SO2

(impurity sulfide) (impurity oxide)

The is normally conducted in a horizontal converter having tuyeres along one side. The air blowing is controlled to convert the remaining more noble metal (less stable metal oxide) to the required metal, i.e.

MS + O2 = M + SO2

Owing to the larger volume of MS present compared with impurity metal sulfides, oxidation of MS to MO will take place initially followed by a mutual reduction reaction between MS and MO:

MS + 2MO = 3M + SO2

In practice this is achieved by adding more MS to the converter or lower PO2. Thus, oxidation of a metal sulfide to its metal is only possible where the mutual reduction reaction has a negative free energy change.

Fig. 13-1 The Cu—Fe—S phase diagram showing the composition ranges

in which typical copper mattes are found

Lesson 14 Slags

gaseous atmosphere 气氛

thermal barrier 隔热层

interface n. 分界面，接口，边界，界面

fluorspar n. 英石，氟石

sole a. 唯一的，单独的

trap v. 捕获，收集

rate-controlling factor 速度控制因素

orthosilicate n. 原硅酸盐

acid slag 酸性渣

basic slag 碱性渣

neutralization n. 中和

amphoteric a. 两性的

undissociated a. 未离解的

pentoxide n. 五氧化物

appoach n. 方法，途径，处理

basicity n. 碱性，碱度

invalid a. 无效的，不成立的

survey v. 检查，调查，评价

interpretation n. 解释，分析

desulphurization n. 脱硫作用

self-consistent 前后一致的，独立的

Slags, which consist primarily of oxides, are used in most pyrometallurgical processes—that is processes involving elevated temperatures and broadly covering extraction of metals such as iron, zinc and lead using a reducing agent such as carbon, refining by preferential oxidation (fire refining of copper and steelmaking) and matte smelting and conversion in the extraction of copper and nickel from sulfides.1 Slags fulfill two main functions—in extraction processes they take up the gangue minerals which are not reduced to the metallic state and in refining processes they act as the receiver for unwanted constituents of the metal. Iron provides a good example of both these functions the slag in the iron blast furnace contains the gangue minerals such as silica, alumina and calcium oxide and, because it is liquid and separates easily from liquid iron, provides a method of removing these waste materials from the furnace separately from the liquid metal.2

The iron produced by the blast furnace contains up to 10% impurity elements by weight, and to convert this relatively useless material into a purer alloy, steel, oxygen is introduced into the iron. Silicon and manganese are converted into oxides which enter the steelmaking slag, and by suitable adjustment of the slag composition, sulfur and phosphorus may also be removed from the liquid metal into the slag. Extraction slags can play a refining role—as in the iron blast furnace where some sulfur is removed from the iron by the slag. Slags may also control the supply of oxygen, nitrogen, hydrogen and sulfur from the gaseous atmosphere of a reverberatory furnace to the liquid metal and can act as thermal barriers where heat is either leaving a liquid metal bath or entering it from a flame playing on the slag surface (open hearth steelmaking).

To allow these functions to be adequately fulfilled, slags must possess the following properties: they must be sufficiently fluid to allow easy separation from the metal and to increase the rate of mass transfer to and from the slag/metal interface. They must become fluid at a low enough temperature for the process to be worked economically with as little heat input and refractory wear as possible—fluxes such as lime, quartz, fluorspar or iron oxide may be added solely to lower the liquidus temperature and viscosity of slags. Their specific gravity must be sufficiently different from that of the metal to allow easy separation. They must have the correct composition and structure to dissolve impurities and gangue minerals at low activity and to allow any desired slag/metal reactions to occur.

The structures of molten oxides were mentioned briefly, and as silica forms the basis of most slags being the commonest gangue constituent—we can form an initial picture of slags based on two types of oxide, RO and SiO2, where RO can represent any of the oxide CaO, MnO, MgO, FeO, ZnO, PbO, Cu2O, Na2O, K2O. Ward outlined a method of defining the composition of a slag in terms of the relative amounts of the two types of oxide, RO and SiO2. RO is called a basic oxide because it provides oxygen ions when dissolved in a slag.

RO = R2+ + O2-

Silica is an acidic oxide which will absorb oxygen ions provided by a basic oxide.

SiO2 + 2O2- = SiO44-

An acid slag is one which contains more acidic oxide than the orthosilicate composition 2RO·SiO2 at which each silicon atom exists as a separate SiO44- anion.3 A basic slag contains more basic oxide than the orthosilicate composition and must therefore contain excess oxygen ions which are not part of the silicate anion structure.

In slags, other acidic oxides than silica are present and their oxygen requirements must also be satisfied by the addition of basic oxides before the slag becomes basic. For example, the neutralization of alumina and phosphorus pentoxide can be represented by the equations:

Al2O3 + 3O2- = 2AlO33-

P2O5 + 3O2- = 2PO43-

And each molecule of alumina and phosphorus pentoxide would require three molecules of basic oxides to neutralize them. This approach is useful in certain cases but it must be remembered that the picture is complicated by the tendency of some oxides (Al2O3, Fe2O3, SnO2, ZnO and PbO) to behave amphoterically, that is as acidic oxides in basic slags and as basic oxide in acid slags.

The extent to which aluminum is present as polymer anions in acid slags is not known, and this presents a considerable barrier to the understanding of the behavior of iron blast furnace slags which tend to have significant alumina contents. For the sake of comparison, the basicity of certain extraction and refining slags has been calculated. The basicity is calculated as the ratio

[(Moles RO) – 3 (Moles Al2O3 + Moles P2O5)] / [2(Moles SiO2)]

and although the assumption that amphoteric oxides do not exist is invalid in the statement that a basicity of 1 indicates a neutral slag, it is felt that the comparison is still of some value.

In order that the activity of slag components can be calculated from slag compositions—for the purpose of predicting equilibria in slag/metal systems—some model of slag behavior must be proposed. Ward gave an excellent survey of slag models which have been used. The models are for ferrous extraction and refining because it is in steelmaking that the control of slag/metal reactions is most important and has received most attention, but there seems no reason why the treatment should not be extended to other metals than iron.4 The earliest models were based on the assumption that undissociated molecules formed the structure of liquid slag; for example in basic slags, the silica existed entirely as orthosilicate molecules 2RO·SiO2 and the remaining basic oxides were present as “free” oxide molecules CaO, MnO, etc. This approach allowed a reasonable interpretation of the behavior of phosphorus in basic steelmaking processes.

The “molecular theories” are extended by Schenck to include the possibility that these molecules might be partially dissociated.

2CaO·SiO2 = 2CaO + SiO2

For example “Dissociation constants” were calculated being the apparent equilibrium constants for the dissociation reactions under the conditions existing in the slags, and hence the proportion of “free” CaO available for desulphurization of liquid iron in the blast furnace by the reaction could be calculated for acid slags.

(CaO) + S = (CaS) + O

The molecular theories eventually became so complex when they tried to account for every eventuality that it became imperative that the ionic nature of slags be recognized in any slag models5. The attraction of ionic slag models, in addition to the knowledge that slags are of ionic structure, is the fact that the number of ionic species may be limited and therefore the treatment may be intrinsically less complicated.

This is certainly true of basic slags, but the range of possible polymer anions in silica—rich slags has so far prevented the establishment of a self—consistent theory for slags. Temkin considered that a basic steelmaking slag consisted of SiO44-, PO43-, AlO33-, and, O2-, anions, and the cations R2+ from the basic oxide, and that the slags were ideal ionic solutions. The anion and cation functions were separated assuming that the anions and cations existed separately on their own lattices. The Temkin model is satisfactory so long as it can be assumed that all the cations are equally important, but this is not the case—for Ca2+ is more important in phosphorus removal than Mg2+. The theory of Flood, Forland and Grjotheim uses Temkin’s model but takes this difference into account, shows the importance of lime in phosphorus removal in steelmaking and is the most successful attempt to provide a model of liquid slags based on their ionic structure.6 They consider cation equilibrium, which is safer than anions because of the whole cation do not change their extent of polymerization or complex—formation as much as anions.

Lesson 15 Reduction of Metal Oxides

Cope v. 对抗，克服 cope with

Contamination n. 污染

Evolution n. (气体等的放出)释放，进展

Partial pressure 分压

Procedure n. 过程，步骤，方法

Tuyere n. (冶金炉)风口，风眼

Favorable adj. 有利的，顺利的，合适的

Onset n. 开始，发动

Penetrate v. 穿透，穿过，掺入

Reasoning n. 推论，推理，论证，论据

Diffusivity n. 扩散性，扩散系数

Superior adj. 在上的，优越的，优良的

Metallothermic reduction 金属热还原

Electric arc furnace 电弧炉

Inert atmosphere 惰性气氛

Bomb n. 还原钢弹

Assess v. 估计，估定，评定

Metal oxides may be reduced to the metal by carbon, carbon monoxide, hydrogen or other metals which form more stable oxides (more negative △G for oxide formation) than the metal oxide to be reduced.1 If the oxides and the reduced phases are assumed to be pure solids and MO is the hypothetical metal oxide required to be reduced, the reaction lines on the △Go—T diagram can be used to assess the various reducing agents.

2M(s) + O2(g) = 2MO(s) (15-1)

C(s) + O2(g) = CO2(g) △Ho298 = -397KJ (15-2)

2C(s) + O2(g) = 2CO(g), △Ho298 = -222KJ; (15-3)

2CO(g) + O2(g) = 2CO2(g), △Ho298 = -572KJ; (15-4)

2H2(g) + O2(g) = 2H2O(g), △Ho298 = -486KJ; (15-5)

From △Go—T diagram it can be seen that if a sufficiently high temperature is achieved any metal oxide may be reduced with carbon according to the reaction:

2MO(s) + C(s) = 2M(s) + CO2(g) below 650℃ (15-6)

MO(s) + C(s) = M(s) + CO(g) above 650℃ (15-7)

Reaction (15-6) is therefore only applicable to the reduction of the less stable, low melting point oxides such as PbO.

As the minimum reduction temperature rises above 1,800℃ the cost of reduction with carbon increases substantially. This is due to the problem of providing refractory materials that can cope with these high temperatures and the increases reactivity of the metal with its environment resulting in increased contamination.2 For these reasons iron, manganese, chromium, tin, lead and zinc, are the main metal oxides reduced with carbon. Owing to the slow rates of reaction for (15-6) and (15-7) it is likely that CO is first formed by reaction (15-3) and which then reduces the metal oxide according to the reaction:

MO(s) + CO(g) = M(s) + CO2(g) (15-8)

The CO2 produced is simultaneously reduced with the carbon，i.e.

CO2(g) + C(s) = 2CO (g) △Ho298 = +175KJ (15-9)

Addition of reactions (15-8) and (15-9) produces the overall reaction given in (15-7). This two—stage process offers an improvement in the reaction kinetics of the reduction with carbon due to the gas—solid reactions of (15-8) and (15-9) compared with the solid—solid reaction of (15-7). This point becomes more important at lower temperatures (below 700℃).

By examination of the enthalpies for these equations it is seen that the amount of heat needed to be added to or evolved from the extraction reaction will depend on the △Ho value for reaction (15-8).3 This may be found by reversing reaction (15-1), i.e. dissociation of the metal oxide (2MO = 2M + O2), and a adding reaction (15-4). Metal oxide dissociation is endothermic and will therefore be the controlling factor. However, the large exothermic value of reaction (15-4) produces exothermic reduction of the lesser stable oxides when CO is used as the reducing agent. The reduction of metal oxides with CO alone can be assessed by examining reactions (15-1) and (15-4).

Under standard thermodynamic conditions CO gas will reduce all the oxides above the 6FeO + O2 = 2Fe3O4 reaction line and, as stated above, e.g. Fe. Thus reduction of most metal oxides with carbon is endothermic due to the smaller exothermic nature of reaction (15-3) compared with reaction (15-4) while reaction of FeO and less stable oxides with CO is exothermic; the enthalpy values become more positive in each case with increasing stability of the oxide.

Changing the partial pressures of CO, CO2 and O2 will alter the position of the reaction of the reaction lines on the △Go—T diagram as discussed above. If the CO/CO2 ratio is increased the 2CO + O2 = 2CO2 reaction line is lowered (more negative △G) allowing other metal oxides, e.g. ZnO to be reduced with CO. In the same way reducing the partial pressure of CO will make the reaction line for 2C + O2 = 2CO more negative thereby lowering the minimum reduction temperature of metal oxides with carbon. This later procedure is not normally carried out in practice due to the cost of operating at low pressures.

Coke added to the blast furnace first reacts with preheated air injected at the base of the stack through the tuyeres to produce CO2. At the very high temperatures achieved at the tuyeres, thermodynamics indicates that CO should be formed but the non—equilibrium condition and excess of oxygen result in the production of CO2 which is subsequently reduced to CO by—reaction with coke outside the tuyere zones

inside tuyere zones C+O2=CO2 △Ho298= -397KJ

outside tuyere zones CO2+ C= 2 CO △Ho298= +175KJ

Using an earlier discussion, reduction of metal oxides with CO is the most favorable reaction at the lower temperatures (below 700℃) in the blast furnace while above this temperature reduction with C is thermodynamically favorable. However, due to the high CO/CO2 ratios the 2CO + O2 = 2CO2 reaction line is lowered and the improved reaction kinetics associated with the gas—solid reaction (15-8) compared with the solid—solid reaction (15-7), reduction of metal oxides with CO takes place at higher temperature than 700℃ in the blast furnace. The limit to reduction with CO in the blast furnace is generally determined by the temperature at which slag begins to form. At the onset of slag formation the surface of the metal oxide will be covered with a layer of liquid slag which will prevent contact with carbon monoxide. The only method available for reduction at this stage is for the solid coke to penetrate the slag layer and provide a solid coke—solid metal oxide reaction according to reaction (15-7).4 Thus, although predominately indirect reduction of metal oxides with coke takes place in the blast furnace there is also some direct reduction.

Reduction of metal oxides with hydrogen is of less importance industrially than with C or CO. The lower enthalpy value for the reaction 2H2 + O2 = 2H2O results in less exothermic reduction of metal oxides with hydrogen than with carbon monoxide; most metal oxides being reduced endothermically. Examination of reactions (15-2), (15-3) and (15-5) indicates H2 is better reducing agent of metal oxides than carbon at the lower temperature (below 650℃). From the same thermodynamic reasoning CO should be a better reducing agent than H2 below 800℃ and H2 better than CO above 800℃. However, due to its increased diffusivity, hydrogen is often superior to CO below 800℃ in the reduction of metal oxides while, due to improved reaction kinetics, CO is a more effective reducing agent above 800℃. In certain cases CO and H2 are added as joint reducing agents.

Other reducing agents

In certain cases a metal oxide is reduced with another metal which forms a more stable oxide. These are called metallothermic reduction reactions and are normally used when the metal to be extracted forms stable carbides (Ti, Cr, Nb) on reduction with carbon. These reactions never go to completion, leaving some residual unreacted reducing agent in the final metal product together with some unreduced metal oxide in the oxide or slag phase.5 Metallothermic reductions are normally exothermic; the stronger the reducing agent the more exothermic the reaction. Certain of these reactions may be achieved without any initial heat supply. Si, Al and, occasionally Mg are used as the reducing agents.

The main application of this technique is in the production of low carbon ferroalloys as in the aluminothermic production of ferrotitanium, ferrovanadium, ferroniobium and silicothermic production of ferrochromium in electric arc furnaces; scrap iron or iron being added in the required amount; e.g.

3MO + 2Al = 3M + Al2O3

2MO + Si = 2M + SiO2

Vanadium has been extracted by reducing V2O5 with calcium in a sealed reaction vessel or ‘bomb’ operated with an inert atmosphere.

Thermal decomposition

Oxides of the more noble metals, e.g. Au, Ag, Hg, Pt, Pd have low decomposition temperatures at normal atmospheric pressure and may be easily reduced to the metal by thermal decomposition:

2Ag2O = 4Ag + O2 above 220℃

2 PdO = 2 Pd + O2 above 900℃.

Lesson 20 Leaching (1)

Leach v. 浸出

leachable adj. 可浸出的

Scrap alloys 废合金块

Anodic slimes 阳极泥

Apatite n. 磷灰石

Diffusion-controlled (受)扩散控制的

Chemically controlled 化学反应控制的

Agitation n. 搅拌

agitate v. 搅拌

Settling n. 沉降

Pulp density 矿浆密度

Nonporous adj. 无孔的，致密的

Regenerate v. 使更新，使再生

Acid leaching 酸性浸出

Leaching efficiency 浸出率

Optimum adj. 最适度的，最佳的

Feldspar n. 长石

Sericite n. 云母

Cyanide n. 氰化物

Pyrite cinder 黄铁矿烧渣

Flue dusts 烟尘

Ferric sulphate 硫酸铁

Polysulfide n. 多硫化物

Thiosulfate 硫代硫酸盐

Salt roasting 食盐氯化焙烧法

Chlorine n. 氯，氯气

Abandon v. 抛弃

Corrosion n. 腐蚀

Spent acid 废酸

Sulfuric acid 硫酸

nitric acid 硝酸

sulfurous acid 亚硫酸

aqua regia 王水

Base n. 碱

Bauxite n. 铝土矿

Monazite n. 独居石

Hydroxide n. 氢氧化物

Wolfmite n. 黑钨，锰铁钨矿

Sheelite 白钨矿

Ammonium n. 铵

Ammine n. 氨络合物

1. General Principle and Agent of Leaching

Leaching is the process of extracting a soluble constituent from a solid by means of a solvent. In extractive metallurgy it is the process of dissolving a certain mineral (or minerals) from an ore or a concentrate, or dissolving certain constituents from metallurgical products such as calcines, mattes, scrap alloys, anodic slimes, etc. In this respect, either one or two purposes can be achieved:

⑴ Opening the ores, concentrates, or metallurgical products to recover the metal values.

⑵ Leaching easily soluble constituents (usually gangue minerals) in an ore or a concentrate in order to have it in a more concentrated form. e.g., the leaching of tungsten flotation concentrate with hydrochloric acid to dissolve away calcite and apatite.

It is necessary that the ore be finely ground in order to liberate the leachable mineral.1 Economic factors, usually decide the particle size of ore before processing.

The factors influencing the rate of a leaching process can be summarized in the following points:

① Rate of leaching increases with decreasing particle size of the ore since the smaller the particles, the larger is the surface area per unit weight.

② If a leaching process is diffusion-controlled then it will be greatly influenced by the speed of agitation. On the other hand if it is chemically controlled then it will not be influenced by agitation, provided that enough agitation is done to prevent the solid from settling.2

③ Leaching rate increases with increasing temperature. However, this increase is much less remarkable for diffusion-controlled process than for a chemically controlled process.

④ Rate of leaching increases with increasing concentration of the leaching agent.

⑤ Rate of leaching increases with decreasing pulp density, i.e. when a large volume of leaching agent is added to a small of solids.

⑥ If an insoluble reaction product is formed during leaching, then the rate will depend on the nature of this product. If it forms a nonporous layer, then the rate of leaching will greatly decrease. If, however, the solid product is porous, it will not affect the rate.

2. Leaching Agent

The choice of a leaching agent depends on many factors:

① Chemical and physical character of the material to be leached.

② Cost of the reagent.

③ Corroding action of the reagent and consequently the materials of construction.

④ Selectivity for the desired constituent to be leached.

⑤ Ability to be regenerated, e.g., in the leaching of ZnO by H2SO4 the acid is regenerated during electrolysis.

Selectivity of a leaching agent toward a particular mineral in an ore depends on:

① Concentration of the leaching agent: In some cases little is to be gained in leaching the desired mineral by increasing the concentration of the leaching agent.3 On the contrary, the dissolution of other minerals may be enhanced. For example, in acid leaching of copper oxide ores, acidity control has a large effect on the dissolution of undesirable minerals.

② Temperature. Sometimes an increase in temperature has little effect on increasing the leaching efficiency of the desired mineral, but has a marked effect on increasing the level of impurities. It becomes even necessary to cool the leaching agent to the optimum temperature. This is the case in leaching copper oxide ores at Wood Heights, Nevada. During the summer, the leach solution is circulated over a cooling tower to keep the solution temperature at 20.5℃. Very little increase in copper leaching efficiency is noted above this temperature; however, a marked increase in impurities is immediately apparent.

③ Contact time. Extended contact period between the solvent and the ore may result in an increased percentage of impurities in solution. For example, copper oxide minerals when contacted with dilute sulfuric acid are dissolved first from the ore. Iron and aluminum minerals continue to dissolve as the feldspars, slowly break down under the acid. Thus, the minimum contact time resulting in maximum copper recovery and minimum impurity contamination should be selected.

A new and successful leaching agent may cause a revolution in hydrometallurgy; the introduction of sodium cyanide for leaching gold and silver ores is an example

The following leaching agents are in common use.

Water alone, is used to leach calcines produced by sulfating or chloridizing roasting, such as the leaching of zinc sulfate or treated pyrite cinder, and also in the leaching of Re2O7 from flue dusts in MoS2 roasting.4 Water in the presence of air or oxygen, and at about 150℃ dissolved sulfides, converting sulfates.

Aqueous salt solutions:

① Ferric sulfate:---used for leaching sulfide minerals.

② Sodium carbonate:---used for leaching uranium ores.

③ Sodium chloride:---used for leaching PbSO4.

④ Sodium cyanide:---used for leaching gold and silver from their ores.

⑤ Sodium sulfide :---used for leaching sulfide minerals forming soluble polysulfides.

⑥ Sodium thiosulfade:---used for leaching silver chloride produced by salt roasting of ores.

Chlorine water

Chlorine water was once used in leaching gold ores, but was abandoned when the cyanidation process was discovered. It has been suggested for leaching sulfide ores.

Acids

Sulfuric acid is the most important leaching agent. It is the cheapest acid, has only minor corrosion problems encountered with its use, and is effective in opening most ores. It is used either dilute, concentrated, or sometimes mixed with hydrofluoric acid. In many cases, spent acid from electrolytic processes is adjusted to the required concentration and used.

Other acids such as hydrochloric and nitric, are used only to a limited extent. Sulfurous acid is finding new applications for leaching some ores, such as low grade manganese types. Aqua regia is used for leaching native platinum ores, and in the refining of gold and silver by parting.

Bases

Sodium hydroxide is used chiefly for dissolving aluminum from bauxite, for opening monazite sand, and for leaching wolframite and scheelite ores. Ammonium hydroxide is used for extracting those metals (such as copper and nickel) that form soluble ammines from their ores.

Leaching by bases has the following advantages:

① negligible corrosion problems;

② most suitability for ores containing much carbonate gangue;

③ more-selectivity, since iron oxides will not be leached.

Lesson 21 Leaching (2)

Leaching in place 就地浸出

Leaching in situ 就地浸出

Psf (pounds per square foot) 磅/平方尺

Heap leaching 堆浸

Dump leaching 堆浸

Tonnage n. 吨位，吨数

Asphalt n. 沥青，煤焦油

Dump truck 自卸卡车，翻斗车

Perforate v. 穿孔，打眼

Ditch n. 沟渠，明沟

Pad n. 垫片，基底

PVC polyvinyl chloride 聚氯乙烯

Perpendicular adj. 垂直的，成直角的

Sump n. 槽，池，坑

Copper pyrite 黄铜矿

Zinc pyrite 锌黄铁矿

Percolation n. 渗滤

Percolate v. 渗滤，使渗滤

Percolation leaching 渗滤浸出

Vat leaching 槽浸出

False bottom 假底

Filtering medium 过滤介质

Countercurrent n. 逆流

Inapplicable adj. 不适用的

Impervious adj. 不渗透的

Clog v. 堵塞

Thickener n. 浓缩槽，沉降槽

Pulp leaching 矿浆浸出

Paddle n. 桨，叶片

Pachuca tank 帕秋卡空气搅拌浸出槽

Dorr agitator 道尔型混合式搅拌器

Scraper n. 刮板

Blade n. 叶片

Steam coil 蛇形管蒸汽加热

Pregnant adj. 充满的，富有的

Pregnant solution 富液，母液

Filter-cake 滤饼

Filtrate n. 滤出液

Rigorous adj. 严格的，精确的

Batchwise adv. 分批地

Digestion n. 浸出，浸煮

Curing n. 处理，熟化，固化

Dampen v. 弄湿

Acide curing 拌酸处理

Bake v. 焙，烘，烤

Decantation n. 倾析，倾注

Selenide n. 硒化物

Telluride n. 碲化物

Closed vessel 密闭容器

Withstand v. 经受住

Autoclave n.压煮器，高压釜

Cladding n. 包层，覆层

Turbomixer n. 叶轮式混合器

Hood ring 密封垫圈

Impeller n. 叶轮，轮

PSI pounds per square inch 磅/平方寸

2. Method and Equipment for Leaching

The grade of the ore and the ease with which the mineral values are dissolved in a particular reagent are the controlling factors in determining the choice of the leaching method.1 The most common methods of leaching are discussed in the following sections.

Leaching in place

This method is mainly used for copper ores which are too low in grade to justify mining and transportation expenses. The ore is simply shattered and leached in place over long periods of time. The method makes use of the presence in the ore of iron sulfides, which under the combined action of air and water, undergo oxidation over periods of weeks or even years to form ferric sulfate. The end product, essentially copper sulfate, is collected in tunnels. The oxidation reactions are exothermic and the heat generated facilitates continued oxidation. An example of this process is to be seen in the operation of the Miami Copper Company in Miami, Arizona, for leaching a worked—out copper mine averaging 0.15% copper.

Martin (1964) suggested leaching uranium and rare minerals from subterranean deposits in situ with carbonated water at pressures of 600 psi or higher. Leaching in place is now undergoing serious consideration since planning to introduce nuclear explosives to shatter ore bodies.

Heap or dump leaching

This method was first employed in the Harz Mountains in Germany during the sixteenth century. Depending upon the tonnage processed, an area of about 300 by 400 ft is leveled and then covered with an asphalt layer. Low grade ore is then dumped onto the site by dump trucks to a level of 20-30 ft high. Either water or dilute sulfuric acid is then sprayed at the top of the dump, and leach solution is collected in streams at the bottom of the heap. Sometimes, perforated vertical pipes are introduced at regular intervals inside the heap to facilitate the flow of water and the same time allow air circulation to facilitate the leaching process.

In other cases, ditches running the entire width of the pad at regular intervals are made. A perforated pipe of 4 inch diameter is laid in every ditch and is then covered with gravel. The ore about 1 inch in size is then dumped into the site and leveled by a tractor 20-30 ft high. Two parallel 6 inch diameter PVC pipes are installed perpendicular to the buried pipes and are connected to their discharge ends. The PVC pipes discharge into a 100,000 gallon sump.

Heap leaching is used on a large scale in Rio Tinto, Spain, for leaching copper and zinc pyrite ores. The heap is sprayed with water and left for long periods of undergoing action by air, water, and ferric salts till most of the copper is converted to copper sulfate.

When the pyrite is fully leached, it is loaded on trucks and shipped to sulfuric acid manufacturers.

Heap leaching is also used for low grade uranium ores (Mashbir, 1964). In this case, the ore is an oxide containing 0.05% U3O8. Dilute H2SO4 (35%) is added to the top of the heap, and in about 8 days uranium is solubilized by the combined action of air and acid; acid consumption is about 50 lb/ton of ore. Water is then sprayed at the top of the heap to leach the uranium out of the bed. This washing operation usually takes 30 days, leaving a residue containing 0.0006% U3O8. Uranium is solution is about 0.2-0.6g/l and the recovery is 88.3%. It is worthwhile mentioning that without the use of heap-leaching in the case of uranium, the ore has been abandoned as uneconomical.

Percolation or vat leaching

The material to be leached is placed in a tank equipped with a false bottom covered with a filtering medium. The solvent is added at the top of the tank and is allowed to percolate through the material. These tanks are usually arranged so that a countercurrent system is employed; the new solids being added to the last tank and weak liquid to the first and pumped successively from one tank to another till it reaches the last tank, almost saturated. Tanks having a capacity of 12,000 tons of ore are in common use.

This process is well suited to cases where the material is porous and sandy, and is inapplicable to material which tends to pack into impervious masses. Regularity in the size of particles rather than their actual size is the chief factor governing good percolation. The ideal is that where the particles are of unequal size, the small ones pack in the openings between the larger ones, thereby clogging the channels. Extraction becomes slow and channeling of solutions through the bed takes place. The method is therefore unsatisfactory if much slime is present. Its advantages are minimum solvent consumption, the production of high grade, pregnant solution, and elimination of the use of expensive thickeners or filters. When leaching is finished, the tanks are emptied manually, and a new batch is introduced. The method is used for leaching gold, copper, and uranium ores.

Pulp Leaching

Pulp leaching with agitation pulp of ores, concentrates, etc. is usually prepared for leaching by grinding the material in water (to minimize dusting) to produce the optimum particle size. Pulp densities vary from 40 to 70% solids. The leaching agent is added and the pulp is agitated continuously. Agitation may be accomplished by

⑴ Mechanical driven paddles. This is usually for small leaching tanks.

⑵ Compressed air. Pachuca tanks are suitable for this purpose. They consist of a cylindrical tank of about 12 ft diameter and 45 ft high, with a 60 degree conical bottom and are made of wood or rubber-lined steel (fig. 21-1). A central vertical tube open at both ends is provided. Compressed air is introduced through this tube when the tank is charged with the pulp, causing circulation of the materials up the central tube and down the annular space so that the solids are kept in suspension.

⑶ Combined air and mechanical agitation. For large scale leaching the Dorr agitators are extensively used (Fig. 21-2). They consist of circular flat-bottomed tanks with a central tube open at the bottom, through which compressed air is admitted. This central tube also serves as a shaft for two agitating arms, one at the top and the other at the bottom. The bottom arms are provided with scraper blades set at an angle so that they carry any settled material toward the central tube, where it can be lifted by the compressed air. The upper arms help to distribute the aqueous phase. Agitation by compressed air is especially suitable for cyanidation of gold and silver ores, and for leaching uranium ores where oxygen is essential for the process. Dorr agitators may be equipped with steam coils for heating.

Pulp leaching may be carried out in a single stage or two stages:

⑴ Single stage. This may be batch or continuous process. In the continuous process a fixed proportion of the pregnant solution is removed from the circuit as filter-cake moisture; the remainder is returned as filtrate to the leaching tank. The method has the advantage of high economy as to reagent consumption and is applied particularly for ores that require high reagent concentration for efficient extraction.

⑵ Two—stage. In this method the leach solution from the second stage containing the dissolved values and unused reagent is advanced to the first stage. This method has the advantage of recovering the unused reagent.

Hot digestion.

This method of leaching in heated vessels is used when extremely rigorous treatment of material is necessary. Highly concentrated solutions (either acidic or basic) and high temperatures (at or near the boiling point of the solution) combined with efficient stirring are required. The digester is an open vessel, heated externally, and is operated batchwise. Examples of this process are the leaching of ilmenite or monazite sand in sulfuric acid. In some cases hot digestion is carried out in ball mills to achieve thorough grinding during the treatment, as for example in the acid leaching of wolframite concentrate.

Acid curing

The finely divided raw material is dampened with water to about 1% moisture; then concentrated sulfuric acid is added and the material is left to cure in bins or piles, or is baked by heating. The cured material is then pulped with water. The pregnant solution is separated by filtration, or countercurrent decantation. This method is sometimes used for some uranium ores which is not amenable to standard processes. Anodic slimes from copper electrolysis are also treated in such a way to decompose the selenides and tellurides of copper and silver.

Leaching under pressure.

Two types of pressure leaching are to be distinguished:

⑴ In absence of oxygen. In this case the ore is heated with the leaching agent at a temperature above the boiling point of the solution to achieve a high reaction rate. Therefore, the process must be carried out in a closed vessel that withstands the vapor pressure of solution at that temperature.An example is leaching bauxite with caustic soda solution.

⑵ In presence of oxygen. Here the pressure in the autoclave is due to the solution pressure plus the oxygen pressure (or air pressure if air is used instead of oxygen). In this case the rate of leaching depends on the oxygen partial pressure and not the total pressure. This method is used mainly for leaching sulfide ores or uranium oxide ores.

The apparatus used is a pressure reactor of 10~15ft diameter and 25~50 ft long, with their long axis laid horizontally, and constructed of mild steel with all internal surfaces and parts made of suitably resistant materials.2 Stainless steel cladding is the most usually type of corrosion-resistant material but in some cases titanium, special alloy, or acid resistant brick lining has to be used. Some autoclaves are equipped with heating-cooling coils, and usually insulated. Each tank is equipped with turbomixers. The center mixer is usually equipped with a hood ring and an impeller running at about 140 rpm, to mix air thoroughly with the pulp in case an oxidizing leach is essential.3 The autoclaves are at 80~150 psi and at about 120℃ . The autoclave are mounted on a slope ( about 8o) to provide flow by gravity from one to the next if continuous operation is required.

Lesson 23 Ion Exchange and Solvent Extraction

Cuprammonium 铜氨

Stripping 反萃

Rayon 人造纤维

Elution 洗提

Effluent 流出液

Tantalum 钽

Immiscible 不混溶的

Pulsation 脉冲

Proprietary 专有的

perforate 穿孔

chelating 螯合

niobium 铌

append 附加说明

kerosene 煤油

reversed 逆向的

1. Ion Exchange

Ion exchange can be defined as exchange of ions between a solution and a solid by which ions from the solution are taken up and retained by the solid which gives up an equivalent number of ions to the solution without any physical change in its structure.1

The first attempt to utilize ion exchange phenomena was in the field of water softening around 1906, using natural and synthetic silicates.

The improved synthetic organic resins greatly broadened the application potential for ion exchange processes due to their stability and high capacity. The first attempts to apply ion exchange for metal recover were in connection with recovery of copper from waste liquors of the cuprammonium rayon and brass industry, silver from photographic film manufacturing wastes, and chromium from electroplating wastes. Uranium was the first metal to be recovered from leach solutions by ion exchange on a large scale, and the great amount of research done in this field opened the doors to the wide possibilities of using ion exchange for recovering other metal from leach solutions.

The ion exchange process is especially useful in the treatment of very dilute solutions with metal ion concentration of the 10 ppm or less. For solutions with metal ion concentration above 1% this method is generally not of value.

Ion exchange operation consists of two steps

⑴ Sorption. The solution containing the metal values is passed through a bed of resin, whereby the metal ions to be recovered leave the aqueous phase and enter the resin phase. When the bed gets saturated with the metal ion in the feed, the metal ion will appear in the effluent (break through), and therefore the flow of feed should be stopped.2

⑵ Elution. Passing a small volume of a suitable solution that removes the metal ions completely-from the resin.

After each of these two operations, the bed is washed to remove loosely absorbed ions. In this way a concentrated solution of pure metal ions is obtained which can be processed further to recover the metal, and the resin is generated by washing for reuse.

Now ion exchange in extractive metallurgy is being used to fulfill the following purposes:

⑴ Recovery of metal values from leach solution, e.g. uranium and vanadium.

⑵ Separation of closely related metals from a leach solution, e.g. cobalt and nickel, hafnium and zirconium, rare earths, niobium and tantalum, and the platinum metals.

2. Solvent Extraction

Solvent extraction involves two operations:

⑴ Extraction. The metal values in the aqueous phase are extracted by agitation with an organic solvent immiscible in that phase. The two phases are then allowed to separate; the aqueous phase is discarded or recycled and the loaded organic phase saved.

⑵ Stripping. Recovery of the metal values from the loaded organic phase by agitation with a small volume of suitable solution. The stripped solvent is then recycled. In this way a concentrated solution containing the metal values in a relatively pure form is obtained.

Solvent extraction as a means of separation and purification has for long been familiar to the chemical industry. Only in recent years, however, has it begun to achieve recognition in the metallurgical field as a means of recovering metals in solution selectively.3 The first large-scale use of solvent extraction in metallurgy was in connection with preparing uranium containing 1 ppm of contaminants for the atomic energy program. It has been proved in practice that solvent extraction is one of the most economical methods for metal recovery. Now it is being applied extensively in the following field:

⑴ Recovery of a metal from a leach solution

⑵ Separation of two or more closely related metals

⑶ Purification of leach solution i.e. removal of an unwanted impurity such as iron.

It is very important to choose an extractant in solvent extraction. An ideal extractant should fulfill the following requirements:

⑴ Selectivity

⑵ High extraction capacity

⑶ Easily striped

⑷ Separates easily from water

⑸ Safe to handle

⑹ Stable during storage or when in contact with acids or alkalis, i.e. does not hydrolyze during extraction or stripping.

Concentration of the metal being collected can be effected using the solvent extraction (or liquid-liquid extraction) technique. This was developed for the extraction of uranium from very dilute solution and until recently has been used only for some of the less common metals. It is now being used in copper extraction, however, and is likely to be applied to other base metals in the near future. It is also used in the separations of the platinum metals. Solvent extraction has the potential for separating metal from a relatively impure solution at the same time as it concentrates it but the effectiveness in this function depends on which metal and what impurities are involved.4 Modern reagents can be very specific in their action.

The “solvent” is usually a complex organic compound with a replaceable hydrogen atom. Some of these are organic derivatives of phosphoric acid but the most now appear to be chelating compounds marketed under proprietary trade names “LIX” and “KELEX” ----usually with a number appended.

In use the reagent is dissolved in carrier which is usually based on kerosene. This is agitated, with a similar volume of leach solution partially purified if necessary and with a sulfuric acid content of about 0.1%. Any copper in the leach solution is chelated and enters the kerosene phase. The kerosene is allowed to rise and separate from the aqueous layer. It is then agitated with a smaller volume of aqueous solution containing, this time, about 15% sulfuric acid. The reaction is reversed. The copper escapes from the complex and enters the aqueous solution. If the volume ratios of the solutions have been selected for the greatest effect, the concentration of copper in the concentrated liquor may be up to 50 times that in the leach solution, and this is a suitable feed for electrolysis tank.

Large-scale continuously operated mixer/settler unit are in use as well as various types of columns in which the two immiscible phases pass in opposite directions. In the perforated plate column some form of pulsation is needed to cause the lighter phase (the kerosene) to be injected through the perforations as suitably sized droplets to maintain a large interfacial area in the system. It will be appreciated that the reagent is continuously being regenerated for re-use.

Lesson 24 Electrolysis (1)

Aqueous adj. 水的

Ionic adj. 离子的

Electrolytic cell 电解槽

Diagrammatic adj. 图解的

Cathode n. 阴极，负极

Electro-（词头）电（花、动、解）

Hydro- 水，流体

Anode n. 阳极，正极

Polarity n. 极性

Discharge v. 放电，排料

Inert adj. 惰性的

Platinum n. 铂

Evolve v. 放出，转变，展开

Hydroxyl n. 羟基，氢氧根离子

Dilute v. 冲淡，稀释

Reactive adj. 反应的，活性的

Irrespective adj. 不考虑的，不顾的

Amalgam n. 汞齐，汞合金

Coulomb n. 库仑

Ampere n. 安培

Valency n. 化合价

Equivalent n. 等量，当量，克当量

Diatomic adj. 双原子的

Efficiency n. 效率

Current efficiency 电流效率

Parameter n. 参数

Chromium n. 铬

Chromium plating cell 镀铬电解槽

Comprise v. 包含，由。。。。组成

Leakage n. 漏，泄露

Contributory n. 起作用的因素

Recombination n . 重新组合

Redox n. 氧化还原作用

Undergo v. 进行，经历

Energy efficiency 电能效率

Cell voltage 槽电压

Perform v. 做，使用

Proportion n. 比例，比率

Alternative adj. 选择的

Imply v. 含有。。。意思

1. Basic consideration

The fundamental definition of electrolysis is the decomposition of a liquid (aqueous or molten) ionic compound by the passage of an electric current. An electrolytic cell is shown diagrammatically in Fig. 24-1. The cathode is the electrode at which electrons are consumed and the anode is the electrode at which electrons are produced. In this respect the electrolytic cell is exactly the same as the electrochemical (corrosion) cell but the electrode polarities are reversed.

Fig 24.1 Simple electrolytic cell

In the electrolysis of molten sodium chloride about 804℃ only two species of ion exist, namely Na+ cations and Cl- anions. In practice the operating temperature is reduced to 600℃ by addition of other chlorides but these ions do not interfere with the basic process. Therefore the reaction at the cathode is i.e. reduction, and molten

Na+ + e = Na

sodium (melting point 980℃) is produced. The reaction at the anode is i.e. oxidation, and chlorine gas is produced and collected as a valuable by-product.

2Cl- = Cl2 + 2e

In the electrolysis of aqueous copper sulphate solution the ionization reactions are:

CuSO4 = Cu2+ + SO42-

H2O = H+ + OH-

The cathodic reaction is the discharge of copper ions. The anode reaction, however, depends of the electrode material. If the anode is “inert”, e.g. platinum or carbon, oxygen gas is evolved by hydroxyl ion discharge as for dilute sulphuric acid. If the anode is copper, the dissolution of copper is more favorable than ion discharge and hence Cu2+ ions are produced. This is the basis of electrorefining1 of copper.

The concentration of ions present also affects the nature of the electrode product. Consider the ionization reactions for aqueous sodium chloride solution:

NaCl = Na+ + Cl-

H2O = H+ + OH-

The cathode product, hydrogen, is the same irrespective of concentration since sodium is too reactive to be deposited as a metal from aqueous solution (sodium may be deposited as an amalgam at a mercury cathode). In dilute solution oxygen is evolved at the anode with traces of chlorine. In concentrated solution the discharge potential of Cl- ions is such that mainly chlorine gas is evolved.2

2. Faraday’s laws of electrolysis

Faraday’s first law states that the mass (m) of any substance discharged (deposited or dissolved) at an electrode is proportional to the quantity of electricity passed which is measured by coulombs and 1 coulomb (c) is equivalent to 1 ampere second (AS). Hence:

m = WIt/(nF) (24-1)

Where m is the mass (g) discharged of atoms of relative atomic mass W by the passage of a current I (A) for a time t (s), n is the valence of the atom and F the Faraday constant 96500.3 sometimes equation is expressed as

m = ZIt / F

Where Z is the chemical or electrochemical equivalent, W/n. An element which has more than one valency state has a corresponding number of Z values.

Faraday’s second law may be simply expressed in that 1 mole of ions of any substance is discharged by the valency number of Faradays to produce 1 mole of atoms of that substance.4 Thus 1 mole of copper (63.5 g) is produced from 1 mole of Cu2+ ions by 2 Faradays (Cu2+ + 2e = Cu). Similarly, 1 mole of aluminum (27 g) requires 3 Faradays to discharge 1 mole of Al3+ ions. Care needs to be taken when evolution of diatomic gases is considered. One mole of hydrogen atoms is produced from 1 mole of hydrogen ions (H+ ) by 1 Faraday, but 1 mole of hydrogen gas (hydrogen molecules, H2) requires 2 Faradays since

2H+ + 2e = H2

It should be noted that 1 Faraday is equivalent to 1 mole of electrons.

3. Current efficiency

Equation (24-1) gives the theoretical maximum mass of substance which can be discharged by a given current in a given time. Current efficiency refers to the cathode product which in practice is less than theoretically predicted. This parameter is defined as

Current efficiency = actual mass discharged / theoretical mass discharged × 100% (24-2)

Example

Determine the current efficiency of chromium plating cell in which a current of 10 A flowing for 90 minutes deposits 0.9 g of chromium from chromic acid electrolyte (CrO3) comprising Cr6+ ions. The relative atomic mass of chromium is 52.

Using Faraday’s first law to find theoretical mass deposited:

m = 52 ×10 ×(90 ×60) / (6 ×96500) g = 4.85 g

Therefore

Current efficiency = 0.9 / 4.85 × 100% =18.6%

In general, chromium plating from chromic acid has a current efficiency of between 12% and 20%. Some processes such as aluminum refining operate at current efficiencies in excess of 99%. There are many reasons for the difference between theoretical and actual yields. Some of which depend on operating conditions. For example, loss of current by electrical leakage and conversion to heat may be contributory factors. These may be recombination of electrode products unless these are well separated, e.g. Mg + Cl2 = MgCl2, or simultaneous unproductive unwanted electrode reactions may occur, e.g. hydrogen evolution or Fe2+/Fe3+ redox reactions. The electrode product may react with the environment, e.g. sodium metal in moist air, or undergo chemical or physical reactions with the electrodes or the electrolyte, e.g. re-solution of cathode deposit.

Although current efficiency is usually taken as a measure of the efficiency of cathodic deposition, it is also possible to define anodic current efficiency and this may be defined in the same way as equation (24-2). For the situation where a metal anode dissolves chemically in the electrolyte, the weight loss of the anode is greater than theoretically predicted and a current efficiency of greater than 100% is achieved, which can never be the case at the cathode.5 A high anodic current efficiency has less significance of the overall process than a high cathodic current efficiency. However, a low anodic current efficiency may be significant.

4. Energy efficiency

This parameter is usually expressed as the power (W) required to produce 1 kg of product and has units KWh kg-1. Copper refining has an approximate energy efficiency of 0.2 KWh kg-1 whereas the figure for aluminum refining is 20 KWh kg-1 i.e. less energy is used in refining 1 kg of copper than for the same mass of aluminum.6

Sometimes energy efficiency may be calculated as:

Energy efficiency(%) = Current efficiency(%) × theoretical cell voltage / applied voltage (24-3)

One use of equation (24-3) is to study the interrelationship between energy efficiency and current efficiency. If the applied voltage could be kept constant and the current efficiency increased, the energy efficiency, measured as a percentage, increases, i.e. more current is usefully used decreasing the overall power (current× voltage) needed. If the current efficiency is maintained constant and the applied voltage altered the situation is rather more complicated. Increasing the voltage increases the power input but that part of the applied voltage performing useful work may not change in proportion. This is because part of the applied voltage is needed to overcome polarization and resistance effects in the cell and this fraction may markedly increase.

Power input may be reduced by reducing the applied voltage but there is a minimum voltage below which electrolysis does not proceed at a significant rate. Hence optimum conditions in practice are achieved by a balance between current efficiency and energy efficiency. Even though aluminum refining has a relatively poor energy efficiency there is no alternative low energy route. Energy is not only consumed by electrolysis but also by heat required to keep the electrolyte molten. The only way to improve energy efficiency in this case is to use as large a plant as possible. The most important variable in electrolysis, however, is input current since this determines the output of the cell as implied by Faraday’s first law (24-1).

Lesson 25 Electrolysis (2)

Current density 电流密度

Geometric adj. 几何学的

Electroplate v. 电镀

Activation n. 活化，活性

Polarization n. 极化

Limiting current density 极限电流密度

Substrate n. 底层，基层，基质

Throwing power 电镀能力，着电效率

Reversible adj 可逆的

Reversible electrode potential 可逆电极电位

Electrochemical series 电化序

Emf electromotive force 电动势

Precious adj. 贵重的，宝贵的

Diaphram n. 隔膜，隔板

Anolyte n. 阳极液

Compartment n. 间隔，室，层

Slime n. 泥渣，软泥

Anode slime 阳极泥

Overpotential n. 超电压，超电势

Implication n. 包含的东西

Back emf 反电动势

Decomposition n. 分解

Extropolate v. 外推，推断

Bus-bar n. 母线，导电排

Acidify v. 使酸化

Graph v. 图，图解

5. Current density

This parameter is the current per unit area of electrode and has units A·m-2 but is often quoted as A·cm-2. For example in copper refining the current density is 220 A·m-2 and for aluminum refining 15,000 A·m-2 are required. Normally the measurable geometric area of the electrode is satisfactory for determining current density but in some electroplating processes a rough surface greatly reduces the expected current density. Both anodic and cathodic current densities may be considered but they need not necessarily be equal.

Activation and concentration polarization effects are directly dependent on current density and therefore it may be desirable to reduce current density. If the current were reduced then output would also be reduced, if the electrode area were increased current density would decrease without affecting the applied current. In some cases a high current density is desirable. For example, gas cleaning by promoting cathodic gas evolution. Also in electroplating the current density used is as high as possible without producing ‘burning’ (dark, powdery deposit). The highest current density at which ‘good’ plating is achieved is called the limiting current density. Current density also affects the covering power and throwing power of plated metal on a substrate.

6. Theoretical cell voltage

The voltage, Ecell, is the minimum theoretical voltage required to electrolyze a given electrolyte. In principle as soon as this voltage is achieved between an anode and a cathode, electrolysis i.e. discharge of ions occurs. Within this topic area it is necessary to distinguish between aqueous electrolytes and fused salt electrolytes.

In aqueous electrolysis the theoretical cell voltage is dependent on the reversible electrode potentials of each electrode. The Nernst equation and electrochemical series are used to calculate Ecell. The most simple example of an aqueous electrolyte is water (acidified) at pH = 7, Ec,H2 = -0.413V and Ea,O2 = +0.817V. Therefore Ecell = Ec - Ea = - (0.413) – (0.817) = - 1.23V.

If platinum electrodes dipping into acidified water are connected together directly no current flows and no emf exists between them.1 If an external current is applied oxygen is evolved at the anode and hydrogen at the cathode only when the potential difference between the electrodes has been increased to a theoretical minimum of 1.23V. Another method of calculating Ecell is by using free energy data, and the equation △G = - n·F·Ecell.

2H2O = 2H2 + O2 , △Go298 = + 474 KJ

n = 4, since four electrons are transferred.

4H+ + 4e = 2H2 and 4OH- = 2H2O + O2 + 4e

Therefore

Ecell = - 474000 / ( 4 × 96500 ) V = - 1.23 V

The signs of both △G and Ecell should be noted. The positive value of △G indicates the non-spontaneous direction of reaction, i.e. the reaction as written above should normally go from right to left. Therefore the negative value of Ecell indicates that this voltage must be overcome in order to produce hydrogen and oxygen. Hence applying fractionally more than 1.23 V results in electrolysis.

Calculation of Ecell in this way is particularly important when considering fused salt electrolytes where Eo data are not as readily available as thermodynamic data.2

An interesting variation in electrolytic processes is the electrorefining of nickel from nickel sulfide. The nickel sulfide from the matte is cast into anodes weighting about 250 kg each. The sulfide is regarded as Ni3S2 at this stage rather than NiS and contains about 75% nickel, 20% sulphur, 3% copper, 0.5% cobalt, 0.5% iron with small amounts of lead, zinc, selenium and precious metals. A diaphragm cell is used for electrolysis to separate the impure anolyte from the catholyte.3 Cobalt, iron and copper ions present in the anolyte would deposit at the cathode unless removed. The anolyte is therefore pumped away to be purified and then returned to the cathode compartment. In addition to a diaphragm (which increases electrical resistance) between the compartments the anodes are placed in very large bags to collect the anode slime. The electrolyte used consists principally of nickel sulfate. The anodic reaction for electrolysis is

Ni3S2 = 3Ni2+ + 2S + 6e.

The sulphur produced is collected as part of the anode slime. Nickel ions in the purified catholyte are deposited on the pure nickel cathode according to

3Ni2+ + 6e = 3Ni.

The overall reactions is therefore

Ni3S2 = 3Ni + 2S, △Go298 = + 203 KJ

Since n=0, Ecell = - 0.35 V

7. Polarization and overpotential

Eo for Ni plating from aqueous NiSO4 solution is –0.25V but plating does not begin until the cathode potential is about –0.45V.4 Since the overpotential, is defined as

= Ep + Er (Ep =-0.45V, Er =-0.25V) (25-1)

Therefore o = -0.20V, which is the amount by which the cathode potential must be made more negative before nickel can plate out. Overpotential is defined by equation

= o + e + r (25-2)

Each type of polarization acts against the applied voltage, e.g. the applied voltage must include an extra contribution with Ecell to overcome the activation energy for an electrode reaction. In addition, local electrode concentration and resistive effects set up ‘back emf’ as indicated by and respectively which must also be overcome by the applied voltage. These polarization effects take place at both cathode and anode and the applied voltage, , may be defined as:

E’app = Ecell+ a + c (25-3)

8. Decomposition voltage

The decomposition voltage is the minimum voltage at which appreciable electrolysis occurs. This is best explained by reference to Fig. 25-1. The graph of current against voltage shows that as the voltage increase there is little change in current. From Faraday’s first law (equation (24-1)) this implies little product. Suddenly, however, a small change voltage produces a marked increase in current (product). Extrapolating as shown gives the decomposition potential ED, which is the minimum applied voltage necessary to produce a significant yield and takes account of resistance in the external circuit, Econt (contacts, busbars, etc) and electrolyte resistance, Esol.5 Therefore the decomposition voltage may be defined as

ED = Ecell+ a + c + Eres (25-4)

Where Eres = Econt + Esol

Fig. 25-1 Graph of current against voltage for electrolysis

showing decomposition voltage ED

Electrolytic decomposition of acidified water has been shown to take place at a theoretical applied voltage of 1.23V. However polarization and, to a lesser extent, resistive effects contribute a further 0.5V to this figure in practice.6

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