The Limiting Phenomena at the Anode of the Electrowinning of Zinc from Zinc Chloride in a Molten Chloride Electrolyte

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PROEFSCHRIFT

ter verkrijging van de graad van doctor aan de Technische Universiteit Delft,

op gezag van de Rector Magnificus prof.dr.ir. J.T. Fokkema, voorzitter van het College voor Promoties,

in het openbaar te verdedigen op dinsdag 28 september 2004 om 10.30 uur

door

Steven Christian LANS

mijnbouwkundig ingenieur, geboren te Gouda.

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Dit proefschrift is goedgekeurd door de promotor: Prof. Ph.D. Pr.Eng. Dr.Habil. M.A. Reuter

Toegevoegd promotor: Dr. A. van Sandwijk

Samenstelling promotiecommissie: Rector Magnificus, voorzitter

Prof. Ph.D. Pr.Eng. Dr.Habil. M.A. Reuter, Technische Universiteit Delft, promotor Dr. A. van Sandwijk, Technische Universiteit Delft, toegevoegd promotor

Prof. dr. ir. G.J. Witkamp, Technische Universiteit Delft

Prof. Dr.-Ing. M. Stelter, Technische Universität Bergakademie Freiberg, Duitsland Prof. dr. ir. G. Van Weert, Oretome Ltd., Canada

Dr. L.J.J.Janssen, Technische Universiteit Eindhoven Dr. sc. J. Vandenhaute, Umicore, België

This research was supported with financial aid by Umicore ISBN 90-9018394-9

Cover design: Ingrid van der Kamp

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Dit proefschrift wordt opgedragen aan Suzanne, omdat je er altijd bent, maar er nooit bij kan zijn.

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SUMMARY

The objective of this research is to investigate the possibilities and technological viability for the electrowinning of zinc from zinc chloride. This research contributes to development of an alternative process, because it provides:

• A clear understanding and overview of the present zinc industry and future developments. • A thorough literature investigation, leading to:

Understanding the reasons to abandon the proposed process route previously used (molten salt electrowinning in particular).

Definition of the practical problem and focusing the research on the anode, where the main loss of energy occurs.

• An ingenious experimental set-up enabling visual observations of the electrowinning process that contributed to a creative approach of interpreting the bubble phenomena at the anode.

• A basic model that enables a systematic analysis of the experimental data.

Zinc is extracted from ores and products when they have served a useful life, via physical separation and extractive metallurgy. An analysis of both hydro- and pyrometallurgical processing routes is provided, which leads to a motivation to develop an alternative process for the winning of zinc. The incentives for an alternative process can be summarised as: • Growing environmental awareness.

• Economical considerations, such as reduced energy consumption. • Possibility to process complex zinc-containing materials.

A dry chlorination process can possibly meet these criteria. In order to position molten salt electrowinning of zinc in the current state of technology and research in this area, commercially operated molten salt electrowinning processes and new developments are discussed.

The decomposition potential of zinc chloride is low compared to zinc sulphate, thus theoretically the electrical energy consumption per kilogram of zinc can be reduced. The main properties of molten salts which make them attractive for electrolysis compared to aqueous electrolytes are these high conductivities and diffusivities. In addition, it is possible to extract zinc in the molten state. Zinc chloride itself is not an appropriate electrolyte, because it has a low conductivity, high viscosity, high vapour pressure, is very hygroscopic and has a high solubility for zinc metal, thereby reducing the current efficiency. An electrolyte with more suitable properties is prepared by the addition of alkali chlorides.

When reviewing the existing literature data, it appeared that the energy consumption lies in the range between 1.8 and 8.8 kWh⋅kg-1_{zinc. The electrowinning of zinc from zinc chloride}

has received attention for about a century and has been performed under different conditions, on different scales and with different cell designs. Experimental results have been used to validate and discuss the existing knowledge critically, indicating that scale-up causes an

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increased cell potential. It is proven that the anodic current density has been evaluated incorrectly in literature, since the effective anode area has not been defined correctly.

On the basis of a systematic experimental study into the cathodic reaction of these results it can be concluded that processes occurring at the cathode are not the reason for the relatively high energy consumption of the electrowinning process. The current efficiency is sufficiently high (generally > 98%) and the electrochemical study proved that the process is ohmically limited at the conditions employed. The cathodic current density can be extremely high without experiencing mass transfer limitations, whenever freezing of the electrolyte is prevented. When the electrolyte composition is prepared to have a melting point sufficiently above the operating temperature (e.g. by increasing the ZnCl2 concentration, or by the

addition of LiCl), very high current densities can be applied at relatively low potentials, i.e. j > 1 A⋅cm-2_{, at E}

cell < 2.5 V and current efficiencies > 95%.

On increasing the anodic current density, the energy consumption of the process rises considerably, due to a higher cell potential. The experimental results prove that the bubble phenomena at the anode give considerable rise to the energy consumption of the electrolysis of zinc chloride under the experimental conditions employed. Because the chlorine bubbles growing at the anode surface area tend to stick at the surface, and because of the large volumetric production of gas, a large ohmic overpotential for the anodic reaction emerges. This conclusion could be drawn on the basis of electrochemical measurements and could be visualised in the see-through furnace.

Different concepts have been applied in order to be able to remove the chlorine bubbles from the surface, but none of these were successful. To minimise the energy consumption for the electrowinning process, the ratio of the anodic to cathodic surface area ratio (A:C ratio) is the most important parameter. The ratio of the effective conductivity to the specific conductivity of the electrolyte was determined to be about 0.4 and the anodic overpotential contributes for about 90% to the total overpotential for an A:C ratio of about 1. The anodic overpotential is only about 50% of the total overpotential when the A:C ratio is about 24.

A model has been developed to give clear insight into the contribution of the different terms to the total ohmic drop and it has been determined that, in general, the largest part of the ohmic drop is due to the layer of chlorine adhering to the anode. The ohmic drop has been modeled by representing the interelectrode gap as three layers, i.e. a layer of gas of certain thickness attached to the electrode causing coverage, the plume (a rising mixture of gas and electrolyte) and a layer of pure electrolyte. The dimensions of the plume are characterised by the plume angle, which has been measured by image analysis. The effects of anode material, anode height, pressure and temperature on the electrowinning have been determined.

The velocity of the plume has been modeled by solving the momentum balance, to determine whether the rising velocity of the plume could affect the layer of bubbles adhering to the anode and covering the anode. The model cannot be used to demonstrate the interaction between the two layers, but provides a good and reliable estimate of the plume velocity in the vertical direction.

In general, it can be concluded that the electrowinning of zinc from zinc chloride is technically feasible. This work has clearly demonstrated that the bottle-neck of the process is the anodic reaction, but that it can be more attractive than the conventional aqueous process with regard to energy consumption.

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SAMENVATTING

De doelstelling van dit project is om de mogelijkheden en technische levensvatbaarheid van de elektrowinning van zink uit zinkchloride te onderzoeken. Dit onderzoek draagt bij aan de ontwikkeling van een alternatief proces, omdat het voorziet in:

• Een duidelijk begrip en overzicht van de huidige zinkindustrie en toekomstige ontwikkelingen.

• Een grondig literatuur onderzoek wat heeft geleid tot:

Begrip van de redenen om in het verleden niet de voorgestelde procesroute te ontwikkelen (specifiek voor gesmolten zout elektrowinning).

Definitie van een praktisch probleem en onderzoek gericht op de anode, waar de grootste energieverliezen optreden.

• Een ingenieuze experimentele opstelling waarmee het mogelijk was om het elektrowinning proces te aanschouwen, wat heeft bijgedragen aan een creatieve benadering om de bellenfenomenen aan de te interpreteren.

• Een eenvoudig model dat een systematische analyse van de experimentele resultaten mogelijk maakte.

Zink wordt gewonnen uit ertsen èn uit producten aan het einde van hun levenscyclus, via fysische scheidingsmethoden en extractieve metallurgie. Een analyse van zowel hydro- als pyrometallurgische processen wordt besproken, wat heeft geleid tot een motivatie om een alternatief proces te ontwikkelen voor de winning van zink. De stimulans om een alternatief proces te ontwikkelen kan als volgt worden samengevat:

• Groeiend bewustzijn van het milieu.

• Economische overwegingen, zoals energie besparing.

• Mogelijkheden tot verwerking van complexe, zinkhoudende materialen.

Een droog chlorinatie proces kan mogelijk voldoen aan deze criteria. Om de gesmolten zout elektrowinning van zink te positioneren in de huidige stand van techniek in dit vakgebied, worden commercieel werkende gesmolten zout elektrolyse processen en nieuwe ontwikkelingen besproken.

De decompositiepotentiaal van zinkchloride is laag vergeleken bij die van zinksulfaat, dus kan theoretisch de elektrische energie consumptie per kilogram zink worden gereduceerd. De voornaamste eigenschappen van gesmolten zouten die aantrekkelijk zijn vergeleken met een waterig elektrolyt zijn het hoge geleidingsvermogen en hoge diffusiesnelheden. Bovendien is het mogelijk om zink te winnen als gesmolten metaal. Zinkchloride zelf is geen geschikt elektrolyt, want het heeft een laag geleidingsvermogen, hoge viscositeit, hoge dampspanning, is erg hygroscopisch en is een goed oplosmiddel voor metallisch zink, waardoor de stroom efficiëntie gereduceerd wordt. Een elektrolyt met meer geschikte eigenschappen wordt bereid door de toevoeging van alkalichloriden.

Toen de data uit de bestaande literatuur werden bestudeerd bleek de energie consumptie te liggen tussen 1.8 and 8.8 kWh⋅kg-1_{zink. De elektrowinning van zink uit zinkchloride heeft al}

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ongeveer een eeuw de aandacht van onderzoekers en is uitgevoerd onder verschillende condities, op verschillende schaalgrootten en met verschillende celontwerpen. Experimenteel resultaten uit eigen werk zijn gebruikt om de huidige kennis op dit gebied, welke aangeven dat opschaling een hogere celpotentiaal veroorzaken, te valideren en kritisch te recenseren. Er wordt bewezen dat de anodische stroomdichtheid fout gedefinieerd is in de literatuur, omdat het effectieve anode oppervlak niet juist gedefinieerd is.

Op basis van een systematische experimentele studie betreffende de kathodische reactie kan geconcludeerd worden dat de processen aan de kathode niet de reden kunnen zijn voor de relatief hoge energie consumptie van het elektrowinning proces. De stroomefficiëntie is hoog genoeg (in het algemeen > 98%) en de elektrochemische studie heeft bewezen dat het proces ohm’s gelimiteerd is onder de condities waarbij gewerkt is. De kathodische stroomdichtheid kan extreem hoog zijn zonder waarneembare massa transport limitaties, zolang bevriezing van het elektrolyt wordt voorkomen. Als de elektrolytsamenstelling zodanig wordt bereid dat het smeltpunt voldoende boven de operationele temperatuur ligt (bijvoorbeeld door de ZnCl2

concentratie te verhogen, of door additie van LiCl), kunnen erg hoge stroomdichtheden worden bereikt bij relatief lage potentialen, dwz. j > 1 A⋅cm-2_{, bij E}

cell < 2.5 V en

stroomefficiëntie > 95%.

Als de anodische stroomdichtheid verhoogd wordt, neemt de energieconsumptie behoorlijk toe als gevolg van een hogere celpotentiaal. De experimentele resultaten leveren het bewijs dat belfenomenen aan de anode de oorzaak zijn voor de aanzienlijke stijging van de energieconsumptie ten behoeve van de elektrolyse van zinkchloride onder de toegepaste condities. Omdat de chloorbellen die aan de anode worden geproduceerd aan het oppervlak blijven plakken, en vanwege de grote volumes geproduceerd gas, dient zich een grote ohmse overpotentiaal van de anodische reactie aan. Deze conclusie kon worden getrokken op basis van elektrochemische metingen and kon worden gevisualiseerd in de doorkijk oven.

Verschillende concepten zijn toegepast in een poging om de chloorbellen van het anodeoppervlak te verwijderen, maar geen van allen waren succesvol. Om de energieconsumptie voor het elektrowinning proces te minimaliseren is de anode tot kathode oppervlakteverhouding (A:C ratio) de belangrijkste parameter. De verhouding van het effectieve geleidingsvermogen tot het specifieke geleidingsvermogen van het elektrolyt is bepaald en is ongeveer 0,4. De anodische overpotentiaal draagt voor ongeveer 90% bij aan de totaal overpotentiaal voor een A:C ratio van ongeveer 1. De anodische overpotentiaal is slechts ongeveer 50% van de totale overpotentiaal als de A:C ratio zo’n 24 bedraagt.

Er is een model ontwikkeld om de verschillende termen die bijdragen aan de totaal ohmse val te bepalen. Het grootste gedeelte van de ohmse val is ten gevolge van een laag chloorbellen die kleeft aan het anodeoppervlak. De ohmse val is gemodelleerd door de interelektrode tussenruimte te representeren als drie lagen, namelijk een laag gasbellen hechtend aan het anodeoppervlak die de oorzaak zijn van bedekking en afscherming, de pluim (een mix van opstijgende gasbellen en elektrolyt) en een laag puur elektrolyt. De dimensies van de pluim worden bepaald door de pluimhoek, die is gemeten met behulp van beeldanalyse. De effecten van anodemateriaal, anodehoogte, druk en temperatuur zijn bepaald.

De snelheid van de pluim is gemodelleerd door de impulsbalans op te lossen, om te bepalen of de pluimsnelheid een effect zou kunnen hebben op de laag bellen die aan het anodeoppervlak gehecht zitten en deze bedekken. Het model kon niet gebruikt worden om de interactie tussen deze twee lagen te demonstreren, maar biedt een goede en betrouwbare schatting van de pluimsnelheid in verticale richting.

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In het algemeen kan worden geconcludeerd dat de elektrowinning van zink uit zinkchloride technisch mogelijk is. Dit werk heeft duidelijk laten zien dat de anodische reactie de bottleneck van het proces is, maar dat het een aantrekkelijkere optie kan zijn dan het conventionele waterige proces op basis van een vergelijk van de energieconsumptie.

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1 INTRODUCTION

This chapter will discuss the background of this project i.e. to position the dry chlorination process route among the current principles in zinc production and to discuss the current practices and developments in molten salt electrowinning.

The latter is achieved by considering the total zinc cycle, from raw materials to end-use, the recycle and includes the association of zinc with other metals as well. Subsequently, the focus is more on the metallurgical extraction of zinc, providing an analysis of both hydro- and pyrometallurgical processing routes. With regard to processing of complex ores and secondary materials, growing environmental awareness, energy and economical considerations, it can be motivated that there is a need to carry out research into an alternative process for the extraction of zinc. Thus, based on the provided discussions, the scope is zooming in to a dry chlorination process and specifically molten salt electrometallurgy. In order to position molten salt electrowinning of zinc in the current state of technology and research in this area, commercially operated molten salt electrowinning processes and new developments will be discussed.

1.1 Principles and practices in zinc production

This section has been prepared not only to discuss the extractive metallurgy and recycling of zinc but also illustrates the role and position of zinc in the material cycle of metals and materials, based on the Zinc College Course Notes for the International Zinc Association (Reuter, Dutrizac & Lans, 2002).

1.1.1 The zinc cycle

The production of zinc starts by mining zinc ore, followed by mineral processing and finally extractive metallurgy. Zinc is applied in a variety of products, e.g. as zinc die-castings, but very often associated with steel products to prevent corrosion, or associated with copper in case of brass. The list of multi-component, multi-material products is endless; consider cars, cell phones, computers, etc. At the end of life of these products, they can serve as a raw material to recover, amongst others, zinc metal again, by the sequence: collection, sorting, component recycling and secondary metallurgy to make a product once again. The total cycle of zinc is depicted in Figure 1-1.

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In ores, zinc is often associated with lead, as well as minor, but important metals as silver, germanium and indium. Figure 1-2 shows that zinc is one of the bulk metals that is being produced, usually closely tied with lead production. It is shown that with processing of these ores, other metals are introduced into the system downstream as well, for which no specific ores and extraction routes are available. It can be concluded that the zinc process plays an essential role in the primary production and recycling of these metals as well (Verhoef et al., 2003). Besides these valuable by-products, some undesirable elements are contained in the ore that have to be separated and end up in residues. For example, iron is often disposed of as jarosite, which is not sufficiently stable to be environmentally liable. It is not allowed to produce jarosite anymore in The Netherlands, because of environmental legislation. Incentives for alternative technology arise from environmental considerations, e.g. focusing on prevention of hazardous residues, prolonging the “license to operate”.

Sulfide ores Oxide ores Pb Zn Sn Ni Cu Ti Cr Fe Al Mg Mn Cd Ga _Ge In Ag Bi As Au_Mo_PdPt Bi As Se Te Co Rh Ru Os Ir Rh In Ga V Co Mg Fe Hg Co Fe Hg Ru Os Ir Rh Cu As Sb Cr Au Ag Zn Pb As Zn Cu Sb Fe V Sn Fe Cr Nb Ta Mg Mg Fe Fe Al Al Al As V Cu Ti Cr Mg Mn Al Al Fe BBr Mn Sb Mn Cu Li Cr Cu Mn Ti Zn Mg Ag Cl Ni Fe Zr Sn Pb Ni Cu As Co V Sb Zn Ni Pt Te Se Fe Bi Pb Zn W Ag Nb Ta Mg Sulfide and oxide ores Interconnected Carrier Metal Cycles Hg Sulfide ores Oxide ores Pb Zn Sn Ni Cu Ti Cr Fe Al Mg Mn Cd Ga _Ge In Ag Bi As Au_Mo_PdPt Bi As Se Te Co Rh Ru Os Ir Rh In Ga V Co Mg Fe Hg Co Fe Hg Ru Os Ir Rh Cu As Sb Cr Au Ag Zn Pb As Zn Cu Sb Fe V Sn Fe Cr Nb Ta Mg Mg Fe Fe Al Al Al As V Cu Ti Cr Mg Mn Al Al Fe BBr Mn Sb Mn Cu Li Cr Cu Mn Ti Zn Mg Ag Cl Ni Fe Zr Sn Pb Ni Cu As Co V Sb Zn Ni Pt Te Se Fe Bi Pb Zn W Ag Nb Ta Mg Sulfide and oxide ores Interconnected Carrier Metal Cycles Interconnected Carrier Metal Cycles Hg

Co-elements that also have considerable own production infrastructure. Valuable to high economic value; some used in high tech applications

Co-elements that have no, or limited own production infrastructure. Mostly highly valuable, high-tech metals e.g. essential in electronics.

Co-elements that end up in residues, or as emissions. Costly because of waste management or end-of-pipe measures.

Carrier metals.Bulk metals, generally of lower value Sulfide ores Oxide ores Pb Zn Sn Ni Cu Ti Cr Fe Al Mg Mn Cd Ga _Ge In Ag Bi As Au_Mo_PdPt Bi As Se Te Co Rh Ru Os Ir Rh In Ga V Co Mg Fe Hg Co Fe Hg Ru Os Ir Rh Cu As Sb Cr Au Ag Zn Pb As Zn Cu Sb Fe V Sn Fe Cr Nb Ta Mg Mg Fe Fe Al Al Al As V Cu Ti Cr Mg Mn Al Al Fe BBr Mn Sb Mn Cu Li Cr Cu Mn Ti Zn Mg Ag Cl Ni Fe Zr Sn Pb Ni Cu As Co V Sb Zn Ni Pt Te Se Fe Bi Pb Zn W Ag Nb Ta Mg Sulfide and oxide ores Interconnected Carrier Metal Cycles Hg Sulfide ores Oxide ores Pb Zn Sn Ni Cu Ti Cr Fe Al Mg Mn Cd Ga _Ge In Ag Bi As Au_Mo_PdPt Bi As Se Te Co Rh Ru Os Ir Rh In Ga V Co Mg Fe Hg Co Fe Hg Ru Os Ir Rh Cu As Sb Cr Au Ag Zn Pb As Zn Cu Sb Fe V Sn Fe Cr Nb Ta Mg Mg Fe Fe Al Al Al As V Cu Ti Cr Mg Mn Al Al Fe BBr Mn Sb Mn Cu Li Cr Cu Mn Ti Zn Mg Ag Cl Ni Fe Zr Sn Pb Ni Cu As Co V Sb Zn Ni Pt Te Se Fe Bi Pb Zn W Ag Nb Ta Mg Sulfide and oxide ores Interconnected Carrier Metal Cycles Interconnected Carrier Metal Cycles Hg

Carrier metals.Bulk metals, generally of lower value

Figure 1-2 Interconnection of metal cycles (Verhoef et al., 2003)

It is postulated that the production of zinc from primary materials includes the extraction of other metals as well. When considering the applications of zinc (e.g. galvanised steel and brass), even more metals (materials) enter the zinc cycle. Then, the cycle of zinc depicted in Figure 1-1 can be expanded to a three-dimensional figure, Figure 1-3, relating design and manufacture of a product and the life cycle of that product to the resources required. To realise complete recycling in a sustainable manner, it is necessary to combine knowledge of fundamental metallurgy and thermodynamics, the dynamics of processes and markets, innovativity in product design and manufacturing, and environmental control and policies. Figure 1-3 demonstrates the interaction between those cycles (Schaik et al., 2002):

• the resource cycle representing products and the raw materials, energy efficiency, together with the other influencing factors such as legislation and infrastructure used to create these • the technological cycle represents all technological disciplines to achieve recycling,

including design, manufacture and separation technologies

• the life cycle represents large scale systems, where company/plant/sector boundaries are crossed.

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Thus, it can be argued that another incentive to develop new technology is to deal with complex (secondary) materials in order to close the zinc cycles.

Figure 1-3 View of the various cycles that describe the life of a product

1.1.2 Hydrometallurgical production of zinc

The most common process to produce zinc is the roast-leach-electrowinning (RLE-process). In order to produce zinc electrolytically from concentrates, first a roasting step is required. The concentrate is oxidised in a fluid bed roaster. This roasting step is followed by neutral leaching. The oxides are dissolved in a dilute sulphuric acid solution and in a next unit operation electrolysed. Subsequently, the neutral leach residue is further leached to recover the remaining zinc content. The spent solution is treated to precipitate an iron containing residue, jarosite, or alternatively, goethite or hematite.

Figure 1-4 depicts a flowsheet for the treatment of zinc concentrates and secondary zinc-containing material. The flowsheet can be further extended by several other alternative processes to treat waste products or secondary materials. A literature survey produced a list of 40 (hydro- and pyrometallurgical) possible technologies to treat secondary zinc-containing material (Honingh et al., 2000). Figure 1-4 also incorporates a block for pyrometallurgical processes, which are most commonly used to process complex zinc ores, or other zinc-containing materials (secondary materials). It shows that hydro- and pyrometallurgical processes cannot be assessed isolated from each other, especially when taking primary as well as secondary materials into consideration. Hydrometallurgical processes can treat pyrometallurgically produced zinc oxides again.

Zn-Concentrate Roasting Acid Roasting Acid Neutral Leach Electrolyte Cleaning and Electrolysis Zn Neutral Leach Electrolyte Cleaning and Electrolysis Zn Further Leach ((super)hot/reductive) Jarosite

Jarosite GoethiteGoethite HematiteHematite

Pyro-treatment Dumping

ZnO•Fe2O3 (Zn-silicate)

Fe2O3

FeOOH X[Fe3(SO4)2(OH6)

Residue X[Fe3(SO4)2(OH6) FeOOH Fe2O3 ZnO•Fe2O3 (Zn-silicate) Residue

Pressure Leach Alternatives Pressure Leach AlternativesPressure Leach_Alternatives Pressure Leach Alternatives

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Roast-leach-electrowin process

Zinc sulphides and iron sulphides (e.g. pyrite, FeS2) react in a fluid bed roaster with air

(oxygen) at a temperature of about 910 °C. Zinc oxides and iron oxides react further to form zinc ferrites, which are difficult to leach in normal neutral leaching conditions. The reactions are:

ZnS + 3/2 O2 → ZnO + SO2

2 FeS + 7/2 O2 → Fe2O3 + 2 SO2

2 FeS2 + 11/2 O2 → Fe2O3 + 4 SO2

ZnO + Fe2O3 → ZnO⋅Fe2O3 (zinc ferrite)

The sulphur dioxide gas is converted into sulphuric acid, which can be used to leach the zinc oxide (calcine). The advantage of chlorinating roasting for the alternative process has the advantage that elemental sulphur is produced (Section 1.2), thus a sulphuric acid plant is not required.

Usually, the leaching step is divided into two steps; a relatively cold, neutral leach and a (super) hot acid leach. In the neutral leach, calcine (mainly zinc oxides and zinc ferrites) is contacted with dilute sulphuric acid at a temperature of 65°C; zinc oxides dissolve:

ZnO + 2 H+ → Zn2+ + H2O, or

ZnO + H2SO4 → ZnSO4 + H2O

A thickener separates the solids from the liquids, whereupon the zinc sulphate solution (pH=5) is sent to the purification unit. The solid residues are zinc ferrites and Pb/Ag-residues. The solid residues, including the zinc ferrites, of the neutral leach may be subjected to a (super) hot acid leach (50-150 g⋅dm-3_H

2SO4; 90 ºC) during which zinc ferrites dissolve:

ZnO⋅Fe2O3 + 8 H+ → Zn2+ + 2 Fe3+ + 4 H2O

Alternatively, the zinc ferrites can be treated pyrometallurgically to recover the zinc. The solid residues of the hot acid leach are the Pb/Ag-residues, which can be pyrometallurgically treated. However, as a result of the hot acid leach, apart from zinc, iron is dissolved as well. Because iron gives rise to various problems in the production process, it has to be removed. The jarosite process is the most applied processing operation for precipitating high concentrations of dissolved iron from zinc processing solutions. It remains the least costly of all the iron-removal technologies, at least if residue long-term disposal costs are not considered. The precipitation of jarosite is carried out at a temperature of 90-98 ºC and at pH=1.5. The reaction produces acid, which has to be neutralised with calcine in order to maintain the pH at a constant value. Sodium or ammonium is added in order to initiate the precipitation reaction:

3 Fe3+ + Me+ + 2 SO42- + 6 H2O → MeFe3(SO4)2(OH)6 + 6 H+, (Me = K+, NH4+, H+, Na+)

In the goethite process, air is dispersed in the neutral leach in order to oxidise any Fe2+ to Fe3+_{, which then precipitates as ferric hydroxide. The hematite process is currently only}

operated in Japan, with the objective to precipitate Fe2O3, by hydrolysis of a neutralised

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alternative dry chlorination process route, iron can be removed as hematite, when chlorinating roasting operation is carefully controlled, as will be discussed in Section 1.2.

The zinc sulphate solution is further purified using zinc dust to precipitate metallic Cu, Cd, Ni, Co. There are two two-step purification processes commonly used: ‘Hot-cold Purification’ and ‘Reverse Purification’. Hot-cold purification implies the addition of arsenic oxides and zinc dust at a high temperature of 90 ºC which yields a Cu/Co/Ni-cement, followed by the addition of copper sulphate and zinc dust at lower temperatures (75-80 ºC) in order to precipitate cadmium. On the other hand, reverse purification involves the addition of zinc dust at a relatively low temperature of 55 ºC, which yields a Cu/Cd-cement, followed by the addition of Sb-tartrate at a more elevated temperature (75-80 ºC) to produce a Co/Ni-cement. The knowledge of precipitating impurities by zinc dust can be applied to a molten chloride mixture as well, which will be discussed in Section 1.2.

Electrowinning

Electrolysis is conducted at a temperature of around 35 ºC according to the following reactions:

H2O → 2 H+ + 1/2 O2 + 2e- (anodic reaction)

Zn2+ + 2e- → Zn + (cathodic reaction) Zn2+ + H2O → Zn + 2 H+ + 1/2 O2 (total reaction)

It is only the high overpotential of hydrogen evolution that enables the electrodeposition of zinc. Impurities in the zinc sulphate solution can dramatically reduce this overpotential and cause the formation of hydrogen rather than the deposition of zinc. Especially metallic ions such as antimony, germanium, cobalt and nickel can cause extensive re-solution of electrodeposited zinc, leading to a decrease in the cathode current efficiency of zinc production. Cadmium and copper co-deposit with the zinc to give an impure zinc product. The fluoride ions pit the aluminium starting sheets, and this results in a very firm attachment of zinc to it. This hampers the automatic stripping of the zinc plates from the starting sheets. Zinc electrodeposition is also very sensitive to the presence of many organic compounds, which can cause very spongy zinc deposits or extensive dendritic growth (Kerby et al., 1980). Iron in cell electrolyte increases the rate of anode corrosion and increases the lead and iron contents in cathode zinc (Krauss, 1985).

Additives are added to the electrowinning feed to optimise the cell operation. For example, glue and strontium carbonate are added to counteract changes in zinc polarisation behaviour, deposition structure and current efficiency (Tang et al., 1997), caused by the remaining impurities. The formation of acid aerosols is eliminated or minimised by production of a foam layer (by the addition of foaming reagents) on top of the electrolyte, thus trapping most of the mist (Alfantazi et al., 1997).

Typically, lead-silver (0.5-1.0%) alloys are used as anodes in the tankhouse. Lead contamination of the zinc cathodes from these anodes is reduced by the formation (30–60 days) of a protective, hard, dense layer of PbO2/MnO2 (Prengaman et al., 2000). Acid and

oxygen are produced at the positive anode, zinc of high purity (>99.995%) is deposited on the cathodes. A typical tankhouse may contain up to about 20,000 lead-silver anodes and 20,000 aluminium cathodes depending on the plant capacity and the area of the electrodes (1.1 to 3.2 m2). The cell potential is usually about 3.5 V. A contribution to the cell potential is due to the

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ohmic potential drop, which depends on the electrical conductivity of the solution (0.20-0.35 Ω-1_cm-1_{industrially), resulting in 22-38% of the total cell potential (Alfantazi, 2001). The}

electrolyte conductivity increases with increasing temperature and acid concentration, but decreases with increasing metal sulphate concentration. It has been determined that the conductivity is a linear function of the product of the concentration of hydrogen ions and free water molecules, which is related to water activity (Tozawa, 1993). The interelectrode gap is about 90 mm. Current densities of 400-600 A⋅m-2_{are usually applied and a temperature of}

about 40°C is maintained. Cooling towers are used in order to keep the temperature constant, as the temperature would rise upon current passage. The energy consumption is in the range of 3-3.5 MWh per ton of zinc. The electrodeposited zinc used to be manually stripped, and this limited the size of the cathodes to approximately 1.1 m2. Since stripping has become mechanised, the surfaces of the cathodes have increased (to a maximum of 3.2 m2). After stripping, the zinc sheets are recast in standard commercial blocks. Typical operational data on the electrowinning of zinc from a zinc sulphate solution are given by Table 1-1.

It has been postulated in Section 1.1 that the objective of this work is to look into the possibilities to operate the molten salt electrowinning of zinc at a considerably higher current density, e.g. double, or triple as is employed in aqueous electrowinning. At the same moment, the objective is to reduce the energy consumption, e.g. < 2 kWh⋅kg-1_{, justifying a large}

investment in order to deal with very corrosive chlorides at high temperatures. Since zinc can be produced above its melting point, stripping of cathodes is not required anymore. Impurities in the zinc chloride are still not desirable, because they may affect the purity of the product. On the other hand, impurities that have a detrimental effect in aqueous electrowinning, will not be as harmful in molten salt electrowinning (e.g. impurities lowering the overpotential of hydrogen, or causing dendritic growth of the zinc deposit). The effect of impurities on the molten salt electrowinning operation are beyond the scope of this research.

Table 1-1 Typical tankhouse data

Zn-purity 99.995% Power 3-3.5 kWh⋅kg-1 Current 400-600 A⋅m-2 Temperature ca. 40 ºC Electrode area 1.1 to 3.2 m2 Cathode (Al) Anode (Pb-Ag) Electrode gap 90 mm Electrolyte Zn = 50 g⋅dm-3_{, pH = 5}

Additives glue, antimony, foaming agents (e.g. strontium carbonate)

1.1.3 Pyrometallurgical production of zinc

Pyrometallurgical processes can process both primary complex Pb/Zn concentrates and secondary zinc-containing material very efficiently. Contained impurities that would usually cause severe problems in electrolytic zinc production are treated often without many problems. In the Imperial Smelting Process, which is the dominating pyrometallurgical zinc-producing process, sintering is performed to make the material suitable for smelting. The zinc produced by the Imperial Smelting Process will be further purified by distillation (Zn-refining). Steel consumer products containing zinc (e.g., cars) are mostly treated for their iron content in steel plants (e.g., in mini-mills or basic oxygen converters). The flue dusts

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produced by these plants (e.g., electric arc furnace dust – EAF) contain a significant amount of zinc, which can be further enriched by, for example, the Waelz process and can subsequently be reused as a raw material. A flowsheet illustrating this processing route is depicted in Figure 1-5. Flue dust, created in pyrometallurgical processes, can be treated in several ways, depending on its zinc content and is receiving increasing attention.

Recycled Steel Recycled Steel Steel Plant Sintering (Dwight-Lloyd) Zn/Pb Concentrate Secondary Materials Sinter Sintering (Dwight-Lloyd) Zn/Pb Concentrate Secondary Materials Sinter Imperial Smelting Furnace Coke/Briquettes Hot Blast Zn Slag / Bullion/ Flue dust Imperial Smelting Furnace Coke/Briquettes Hot Blast Zn Slag / Bullion/ Flue dust Zn-Refining Zn-Refining Galvanising Scrap

Sheet steel Galvanising Product

(e.g. car )

Galvanising Scrap

Sheet steel Galvanising Product

(e.g. car )

Waelz Kiln / Oxide Preparation Waelz Kiln / Oxide Preparation

Figure 1-5 Overview of pyrometallurgical processing (closure of the zinc cycle)

The Imperial Smelting Process is the dominating carbothermic zinc process used to produce zinc and lead simultaneously. The raw material for the process can be of complex composition, such as mixed ores and secondary raw materials containing impurities that would cause severe problems in electrolytic zinc production.

The burden is fed at the top through a double bell-and-hopper lock. Preheated blast air is introduced through tuyeres near the bottom of the reactor where coke is being oxidised. The reaction is strongly exothermic, thus providing heat for the reduction of oxides in the upper part of the furnace. As CO rises through the shaft, it reacts with ZnO and with other oxides like PbO, FeO, and Fe2O3 (Figure 1-6). The reactions are:

ZnO + CO → Zn + CO2

PbO + CO → Pb + CO2

Fe2O3 + CO → 2 FeO + CO2

FeO + CO → Fe + CO2

The top gas mixture is about 8-9% zinc and a ratio of CO/CO2 of about 2-3, the remainder

being nitrogen. Within this composition, the Zn vapour is stable from 1000 ºC to 1040 ºC. Ultimately, the zinc in the gaseous product must be condensated in order to obtain liquid zinc. The hot gas mixture is cooled very fast in a shower of liquid lead at about 500 ºC (splash condenser) and re-oxidation of the zinc is minimised. The Pb/Zn mixture is then cooled, on which the liquid zinc metal forms a separate phase, which can be recovered.

From the bottom of the reactor, a lead bullion consisting of a range of metals; i.e., Pb, Ag, Cu, Bi, Sn, Sb, As and Au, is withdrawn at a temperature of between 1200-1250 ºC, as well as a slag containing PbO, ZnO, FeO, Fe2O3, Al2O3, CaO, and SiO2. The slag contains those

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The ISP accepts a wide range of secondary materials including Waelz oxides, electrolytic zinc plant residues, galvanisers' ashes and drosses and non-ferrous flue dusts (Lee, 2000) The oxide materials can to a certain extent be incorporated with the sinter plant feed, but can also be directly injected through the tuyeres (Schneider et al., 2000).

850o_C

Hot Blast

950-1000o_C

Slag (CaO-SiO₂-FeO)

Lead Bullion (Pb-Ag-Cu-Bi-Sn-Sb-As-Au-Ag)

1200-1250o_C Fe₂O₃+CO=2FeO+CO₂ Zn+CO2=ZnO+CO PbO+CO=Pb+CO₂ ZnO+CO=Zn+CO2 ZnO+CO=Zn+CO₂ C+CO2=2CO FeO+CO=Fe+CO₂ O₂+2CO=2CO₂ 1000-1040o_C 8-9% Zn CO/CO₂=2-3 460o_C Top Air 700o_C

Coke (700o_{C)/Sinter(200}o_C)/Brickets

Pb (2% Zn)

450o_C

Pb (2.3% Zn)

560o_C

Cooling Launder/Separation Bath Zn (Cd) 470o_C

Figure 1-6 Schematic overview of the Imperial Smelting Furnace

1.2 Chlorination processes

A chlorination process has similar stages of operations as the conservative aqueous process route, i.e. roasting (chlorinating) → leaching → purification → electrowinning. Chlorinating agents, e.g. Cl2 and HCl, are highly reactive and chlorination is usually fast and complete.

Sphalerite and galena will be completely chlorinated by chlorine within one hour at a temperature of 600 ºC (Kershner & Donaldson, 1961). Elemental sulphur is produced instead of SO2 and H2SO4. Also, it is possible to selectively chlorinate non-ferrous sulphides and

producing Fe2O3, facilitating easy removal in the downstream process.

Wet chlorination processes have the advantage that chlorides have a relatively high solubility (e.g. compared to sulphates), leaching rates are fast at moderate temperatures and a high redox potential is possible across the entire pH-range (Peek, 1996). In case of dry chlorination processes, the leaching step is not applicable.

Purification of a mixture of metal chlorides can be carried out by conventional cementation techniques, from a chloride melt, as well as from aqueous chloride solutions, to recover amongst others the precious metals, copper, cadmium, lead. Iron, usually the main contaminant, can be separated from the zinc chloride by means of several techniques (Devilee, 1997):

• Cementation with metallic zinc (Monk & Fray, 1973b; Devilee et al., 1999)

• Selective electrolysis (Iwanec & Welch, 1969; Monk & Fray, 1973b; Devilee, 1997) • Oxidation to hematite (Iwanec & Welch, 1969; Monk & Fray, 1973a)

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• Solvent extraction (Bloom & Hastle, 1965) • Wet cementation (Monk & Fray, 1974) • Pyrohydrolysis (Peek, 1996)

Electrowinning of zinc from a fused chloride melt or from chloride solutions is much more favourable, because these media have a much higher conductivity than aqueous sulphate solutions. Because of the high solubility of zinc chloride in these media, high current densities can be employed. Lixiviant regeneration is possible, chlorine evolving at the anode can be recycled to the chlorination operation. Baik and Fray (2001) discuss the possibility of the production of oxygen at the anode and zinc and HCl on the cathodic side of a two compartment membrane cell. The literature investigation of the electrowinning of zinc from fused zinc chloride will be discussed in detail in Chapter 2.

A review of dry chlorination processes has been published by Parsons (Parsons, 1978). In this paper, a description of the following options for the treatment of zinc-containing materials by means of chlorination is given:

• Swinburne-Ashcroft process (high-grade zinc-lead ores, <30% gangue)

• Baker-Burwell process (low-grade zinc-lead ores, slimes of a Zn-Pb-Cu-Ag plant) • Malm process (low-grade copper, zinc and lead ores)

• Mines Branch (Canada) investigation (complex zinc, copper and lead ores, 7-21% Zn) • NIM (South-Africa) process (zinc flotation pyrite residue)

Parsons states that the most serious problems of the presented investigations were due to the corrosion of equipment. In addition, moisture was a problem, because it caused a loss of elemental sulphur as sulphates or hydrogen sulphide. Also, the chlorinated products are difficult to handle, due to the sticky nature. These and other processes will be described in further detail in the next sections.

The Ashcroft-Swinburne Process

The Ashcroft-Swinburne Process (Swinburne & Ashcroft, 1899; Ashcroft, 1926, 1935; Parsons, 1978) implements chlorination of zinc sulphide with chlorine, forming ZnCl2 and

elemental sulphur. The zinc sulphide is mixed with fused zinc chloride and treated with chlorine in a converter, the temperature is maintained above the melting point of ZnCl2. Iron,

one of the main impurities in a concentrate is oxidised to the ferric state and precipitated as Fe2O3 on the addition of ZnO. In a separate operation, coarse zinc powder was added in a

stoichiometric amount to remove the majority of impurities as a heavy alloy in granular form. A flowsheet is depicted in Figure 1-7.

As an alternative to this dry chlorination process, the wet Ashcroft process (Ashcroft, 1926, 1934; Parsons, 1978) has been proposed. After chlorination in a converter, the sulphur gas and vaporised metal chlorides are condensed in an aqueous solution with a high concentration of ZnCl2 at a temperature of 130 ºC. The sulphur is recovered as floating layer. Antimony,

bismuth and arsenic precipitate as sulphides. The molten chlorides are first treated to recover silver and gold by substitution of copper or lead. Lead is then recovered by the addition of metallic zinc. Subsequently, different wet methods have been reported to remove residual impurities from the solution. The chlorides are leached or washed with cold water to separate

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it from gangue material. Copper is precipitated by the addition of iron, which can be removed as iron hydroxide when ZnO is added, although the ferric hydroxide slimes were difficult to filter. A flowsheet is depicted in Figure 1-8. Another version of the process includes partial chlorination, separating a mixture of chlorides from FeS and gangue (Ashcroft, 1934; Parsons, 1978).

Converter

Gangue/Fe2O3/Pb/Alloy Residue

Purification Purification Cl2 Electrowinning Electrowinning ZnS ZnCl2 Zn-dust Condenser Condenser S2 S Off-gas Zn ZnO Mixed chlorides/ Fe2O3 Evaporator Evaporator Iron Hydroxide Converter Converter Precious Metal Recovery Precious Metal Recovery Cl2 Electrowinning Electrowinning ZnS ZnCl2 Solution Wet Condenser Wet Condenser S2 Bi, As, Sb S Zn Mixed chlorides/ Gangue Lead Recovery Lead

Recovery _LeachingWater Water

Leaching Copper_Cement Copper

Cement _RemovalIron Iron Removal Pb Zn H2O Fe ZnO Au, Ag Pb Gangue Cu H2O ZnCl2

Figure 1-7 Flowsheet of the dry Ashcroft process Figure 1-8 Flowsheet of the wet Ashcroft process

The Baker-Burwell process

This process (Parsons, 1978) used a rotary kiln to chlorinate the sulphide material, mixed with pebbles, with an excess of chlorine at 150 ºC. The produced sulphur reacted with chlorine to form sulphur monochloride. The solids were leached with water to dissolve the chlorides and filtered to remove gangue and precious metals. The chloride solution was treated to recover the metals by substitution, similar to the flowsheet depicted in Figure 1-8. Metallic zinc was recovered by electrolysis of the zinc chloride aqueous solution. The process was abandoned, due to problems with sulphur plugging and marketing of the sulphur monochloride.

The Malm process

The Malm process (Parsons, 1978) operated a gas tight, porcelain lined, rotary kiln, counter-currently fed by chlorine and sulphide materials. The operating temperature was about 60-70 ºC and the sulphur remained in the chlorinated mass. The chlorination continued until 40-70% of the metals were converted into chlorides. The partially chlorinated material was fed to a multiple hearth furnace, which operated at 400 ºC. Air was introduced into this furnace, roasting the iron chloride to iron oxide, while the liberated chlorine reacted with the unreacted ore, completing chlorination. The chlorides were then leached by water and the gangue material was removed. Purification was carried out by the addition of iron to precipitate Au, Ag and Cu, subsequently Zn metal was added to remove Pb, followed by zinc oxide addition to remove iron. The purified zinc chloride solution was dried and fused salt electrolysis was applied. Sulphur was not separately recovered from this process. This process offered little flexibility to treat other feedstocks, e.g. ores with a large amount of sulphur.

The Mines Branch investigations

The process (Parsons, 1978) involved reaction of zinc-bearing sulphides with chlorine gas at 250 ºC in a rotary kiln, to form a sulphide-free residue and sulphur monochloride gas. The residue was roasted with air at 450 ºC to convert iron chloride into oxide. The ferrous chloride reacted to F2O3, FeCl3 volatilised and was reacted with fresh ore to reduce it to FeCl2. The

calcine from the oxidation step was leached with water to dissolve zinc and copper chlorides, followed by leaching with brine to dissolve lead and silver chlorides.

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NIM (South-Africa) process

This process (Parsons, 1978) describes the chlorination-roasting process recovering a ZnCl2

-CuCl2 mixture, elemental sulphur and Fe2O3. A pyrite concentrate is first dried and then

completely chlorinated in a fluidised bed reactor at 600 ºC. The ferrous chloride is then oxidised in a second fluidised bed reactor operating at 450 ºC.

1.3 Present state of technology in molten salt electrowinning

Electrowinning is an established process for metal extraction. High temperature electrolysis is generally used for aluminium and metals that are extremely reactive and cannot be extracted in an aqueous solution (e.g. Al, Mg, Na, Li). These metals are deposited above their melting points. Molten salt processes involving refractory metals (e.g. Ti, Ta, Nb) focus on electrodeposition of solid metal.

The main properties of fused salts which make them attractive are their extremely high conductivities and diffusivities, compared to aqueous electrolytes, coupled with low viscosities. This allows very high currents to be passed through the electrolyte with only modest voltage drops across the electrolyte leading to low energy consumption (Fray, 1980). The advantages and disadvantages of molten salt electrowinning are listed in Table 1-2 (Minh, 1985).

Table 1-2 Advantages and disadvantages of molten salt electrowinning (Minh, 1985)

Advantages Disadvantages

• high ionic conductivity • high temperature operation • high diffusivity • high vapour pressure of salts

• low viscosity • corrosiveness

• good solvent for metal salt • lower current efficiency

• faster electrode kinetics • inert atmosphere required in many cases • smaller voltage loss due to polarisation

• no competing cathodic reaction of water

• solidification of electrolyte on power failure

• operation at higher current densities

Since the materials of construction for a chloride process will be much more expensive than for the sulphate process route, high current densities are required to jusitify a chloride process route. Operation at high current density allows an electrowinning cell that is relatively small compared to the sulphate cells for equal production capacity, reducing capital costs.

A primary consideration in metal electrowinning from molten salt is the minimisation of the electrical energy consumption per unit weight of metal produced. In order to develop an energy efficient process, it is necessary to know where the major losses occur. Electrolyte properties play an important role, as well as the cell design. The design can range from monopolar cells with a common arrangement of electrodes (vertical or horizontal), to inclined electrodes, to rotating electrodes forcing a centrifugal field to improve product separation (Cox & Fray, 1996; Morris & Fray, 1994; Copham & Fray, 1990). Product separation can also be accomplished with diaphragms. Instead of monopolar cells, bipolar cells can be used to increase the energy efficiency.

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In order to give a good overview of the possibilities, a short overview of commercial molten salt electrowinning processes will be given.

1.3.1 The Hall-Héroult process for the production of aluminium

The Hall-Héroult cell and electrode reactions

The Hall-Héroult cell is depicted in Figure 1-9. The operating temperature is in the range of 950 to 1000 ºC. The overall reaction to produce aluminium from aluminium oxide is:

2 Al2O3 + 3 C → 4 Al + 3 CO2

The cathode reaction for the production of aluminium is not very straightforward, since aluminium is found in various anion complexes. Kisza et al. (1999) determined the kinetics and mechanism of the cathodic reaction, which takes place in three-steps:

preceding chemical step: AlF63- ↔ AlF4- + 2 F-

low-frequency charge transfer step: AlF4- + 2e- → AlF2- + 2 F-

high-frequency charge transfer step: AlF2- + e- → Al + 2 F-

The exchange current density of the second charge transfer step is >20 A⋅cm-2_{and it does not}

contribute to the cathodic overpotential. The first charge transfer step is influenced by the preceding chemical reaction and diffusion. The exchange current density decreases from ~8.5 A⋅cm-2_{in an industrial electrolyte without Al}

2O3, to ~1 A⋅cm-2 in an industrial melt containing

8 wt% of alumina and does not contribute to the cathodic overpotential. The cathodic overpotential (50-100mV) is due a concentration gradient of the NaF/AlF3 ratio at the

interface (Kisza et al., 1999).

The carbon for the anodic reaction is provided by prebaked anodes, which are made of mixture of petroleum coke aggregate and coal tar pitch, with preformed electrical connection sockets. Alternatively, Söderberg anodes are baked in the Hall-Héroult cell by waste heat, but are less favourable, due to their lower power efficiency and difficulties in handling emissions from the baking process. The anode is consumed during the process, according to the elementary anode reaction (neglecting the formation of some CO):

C + 2 O2- → CO2 + 4e-

However, uncomplexed oxygen ions do not exist in the bulk electrolyte. Grjotheim et al. (1982) give an overview of the structural constituents of cryolite-alumina melts. Kisza et al. (1999) proposed a three-step mechanism:

preceding chemical step:

Al2OF106- ↔ Al2OF62- + 4 F-

charge transfer step with an intermediate adsorption: Al2OF62- + C ↔ 2 AlF3 + CxOads + 2e-

charge transfer and electrochemical desorption: Al2O2F42- + CxOads ↔ CO2 + 2 Al2OF4 + 2e-

It was shown that at low alumina concentrations (2 wt% Al2O3) the first charge step is the

rate-determining step and shifts to the second charge transfer step at higher alumina content. The anodic overvoltage is difficult to measure, since the anode is consumed during

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electrolysis. Also, construction material (e.g. coke type, carbon or graphite) and porosity play an important role in the determination of the overvoltage, e.g. 0.56 V for graphite and 0.37 V for petrol coke at a given current density (Grjotheim et al. 1982). Grjotheim et al. (1982) give an overview of the Tafel coefficients for anodic overvoltage on carbon and graphite in cryolite melts.

When the alumina concentration in the cell is nearly depleted, the “anode effect” occurs. A gas layer covers the surface of the anode, indicating a decreased wettability of the anode by the electrolyte. The carbon fluorides, CF4 and C2F6 are evolved in addition to CO and CO2.

Operation with a constant current causes the potential to rise to 30-50 V and arcs or sparks may be produced due to lack of contact. Nowadays, the anode effect is prevented by point feeding of Al2O3 in the electrolyte, to provide a well mixed electrolyte with a high enough

concentration of Al2O3.

Figure 1-9 Schematic view of a Hall-Héroult cell (left: reproduced from Jarrett, 1987a) and picture of a

potline with the anode immersed into the electrolyte

Properties of the electrolyte

Cryolite (Na3AlF6) is the most important component of the electrolyte, because it has a high

solubility for Al2O3. Solid fluxes, are usually added to promote special properties. The

composition of the electrolyte is optimised in order to improve faradic efficiency and minimise the ohmic drop, i.e. the properties of the electrolyte should:

• decrease the solubility of reduced species in the melt • increase (or not decrease) the solubility of Al2O3

• lower the liquidus temperature • increase the conductivity

• decrease the density to provide better product separation • decrease the vapour pressure to minimise vapour loss

Table 1-3 shows the effect of various additives on the properties of cryolite. The physical and chemical properties of the electrolyte impact cell design, electrolyte flow, cell life, productivity and power efficiency. Some species are introduced with the feed and build-up in the electrolyte e.g. CaF2 and some MgF2 (Haupin, 1987a). They have the beneficial effects of

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electrolyte. However, the detrimental effects are that it lowers Al2O3 solubility, increases

density, viscosity and surface tension of the melt. A common additive is AlF3, which

beneficially lowers the liquidus temperature, solubility of reduced species, viscosity and surface tension of the melt. It has the undesirable side effect of decreasing Al2O3 solubility

and electrical conductivity and increasing the vapour pressure. The acidity of the electrolyte, expressed as the ratio NaF/AlF3, is an important characteristic in cell control and operation

(Haupin, 1987a). Figure 1-10 shows the ternary system of Na3AlF6-AlF3-Al2O3, which is used

to determine Al2O3 solubility at a certain liquidus temperature. The physical properties

mentioned above can be plotted in the phase diagrams, because they relate to the possible phases in the system of all constituents of the electrolyte, an overview of these diagrams has been given by Grjotheim et al. (1982).

Table 1-3 The qualitative effect of various additives on the properties of cryolite, ↑ = increase, ↓ = decrease, / = highly non-linear decrease, 0 = highly non-linear increase (reproduced from

Haupin, 1987a) Increased variable Al2O3 solubility Electrical conductivity

Density Viscosity Liquidus temperature Metal solubility Surface tension Vapour pressure CaF2 ↓ ↓ ↑ 0 ↓ ↓ ↑ ↓ AlF3 ↓ ↓ ↓ ↓ / ↓ ↓ ↑ LiF ↓ ↑ ↓ ↓ ↓ ↓ ↑ ↓ MgF2 ↓ ↓ ↑ ↑ ↓ ↓ ↑ ↓ NaCl ↓ ↑ ↓ ↓ ↓ ↓ ↓ ↓ NaF / ↑ / / / ↑ ↑ ↓ Al2O3 - ↓ ↓ 0 ↓ ↓ / ↓ Operating Temp. ↑ ↑ ↓ ↓ - ↑ ↓ ↑

Figure 1-10 Ternary system Na3AlF6-AlF3-Al2O3 (reproduced from Haupin, 1987a)

Energy considerations

The minimum required energy to produce aluminium can be determined from thermodynamic calculations of the Gibbs energy. When considering the decomposition of pure Al2O3 into Al

and oxygen at 1000 ºC, 8.69 kWh⋅(kg Al)-1_{is required. When considering the Hall-Héroult}

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required. In practice, the energy consumption is related to the cell potential and current efficiency, according to:

CE E M CE F n E W cell Al cell ₌ _⋅ ⋅ ⋅ ⋅ ⋅ = 2.98 3600 [kWh⋅(kg Al) -1_] Eq. 1-1

The (absolute) cell potential can be written as follows:

Ecell = E0’ + ηca + ηan + ηconc.ca + ηconc.an + I⋅(Rca + Relec + Ran + Rextern) Eq. 1-2

Thus, several overpotentials add to the decomposition potential, which can be determined from Gibbs energy thermodynamic calculations and the Nernst equation. The contribution of anodic and cathodic overpotential have been discussed previously in this section. The concentration overpotential is due to depletion of reactive species at the electrode. The resistance of the anode, electrolyte, cathode and bus bars, etc. will increase the potential on increasing current, according to Ohm’s Law. The resistance of the electrolyte depends on the composition, operating temperature and also gas bubbles. In practice, the total cell potential in a modern cell mounts to about 4.1 to 4.2 V, which is depicted in Figure 1-11, showing the main contributions to the cell potential (data extracted from Kvande & Haupin, 2000).

Figure 1-11 Energy consumption across a Hall-Héroult cell for 93% CE (Kvande & Haupin,2000)

The cathodic CE is determined by the aluminium loss and is defined as the ratio of the actual metal output over the theoretical amount, based on Faraday’s Law. Aluminium is soluble in the electrolyte, the amount ranging from 0.06 to 0.21 wt% Al, depending on electrolyte composition and temperature (Grjotheim et al., 1982). Due to dissolution or dispersion of metal, the electronic conductivity of the metal-salt mixture may increase, which also contributes to a decreased CE (Haarberg & Thonstad, 1989). The metal can react anywhere in the melt with dispersed or dissolved gas from the anode. The degree of convection of the melt (basically due to bubble evolution and electromagnetic forces acting on the aluminium) and the interelectrode distance are important parameters in relation to the CE (Figure 1-12). In practice, an interelectrode gap of about 5 cm is maintained. The CE is up to 95% in modern cells.

A Hall-Héroult cell is internally heated by the currents passing through it. The insulating construction materials of the Hall-Héroult cell are covered by a frozen layer of electrolyte, to

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prevent chemical attack by the molten fluoride and resist melt penetration. To maintain this frozen layer, the operation temperature is actually pre-determined by the liquidus temperature of the melt. The simplified energy balance is given by:

Q H n

W = ⋅∆ + Eq. 1-3

Thus, the input of electrical energy is determined by the heat loss from the cell (Q), plus the change in enthalpy of the overall reaction (∆H) time the number of moles reacted, which is given by: 100 3 CE F I n= ⋅ Eq. 1-4

If the CE decreases for some reason, the available heat for the surroundings increases, resulting in melting of the solid crust. Since the area for heat transfer is determined by the cell dimensions, the requirement of the frozen ledge to a large extent determines the thermal efficiency (and kWh/kg Al) of the process. Figure 1-13 shows the heat losses for a centre-fed cell. Cells that are fed from the side (breaking away the crust between the anode and the frozen side layer) are more heavily insulated, requiring 70% of the heat loss to be through the anodes and top crust (Jarrett, 1987b).

In modern cells, the CE is up to 93% and the cell potential is 4.1 V at minimum, resulting in an energy requirement of 13.1 kWh⋅(kg Al)-1_.

Figure 1-12 Recombination of anodic and cathodic products

(reproduced from Haupin, 1987c) Figure 1-13 Heat loss distribution of a centre-fed cell

(reproduced from Jarrett, 1987b)

1.3.2 Magnesium, sodium and lithium production Electrode reactions

Magnesium is electrolytically produced from MgCl2, which is prepared from an aqueous

MgCl2 brine, carnallite (MgCl2⋅KCl⋅6H2O) deposits, or (carbo-)chlorination, or chlorinating

leaching of magnesia (MgO) or magnesite (MgCO3). The extraction consists of preparation of

The Limiting Phenomena at the Anode of the Electrowinning of Zinc from Zinc Chloride in a Molten Chloride Electrolyte

CONTENTS

SUMMARY

SAMENVATTING

1 INTRODUCTION