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Aluminum

William B. Frank, Aluminum Company of America, Alcoa Center, Pa. 15069, United StatesWarren E. Haupin, Aluminum Company of America, Alcoa Center, Pa. 15069, United States

Ullmann's Encyclopedia of Industrial Chemistry

Copyright 2002 by Wiley-VCH Verlag GmbH & Co. KGaA. All rights reserved.DOI: 10.1002/14356007.a01_459Article Online Posting Date: June 15, 2000

4. Production

The only method now used industrially to produce primary aluminum is the HallHroult process.Production of aluminum before its development was discussed previously in this article. Alternatemeans of producing aluminum are treated in Section Alternate Processes.

4.1. History of the Electrolytic Reduction of Alumina

The technological elements of the process electrolysis of fused salts to produce metals(including aluminum), the use of cryolite as a flux to dissolve alumina, and the use of carbonelectrodes had been exploited for some time prior to 1886. The workable electrolytic processwas discovered independently, and almost simultaneously, in early 1886 by CHARLESMARTINHALL in Oberlin, Ohio, and PAUL L. T. HROULT in Gentilly, France. Both these young scientistswere familiar with the work ofSAINTE-CLAIREDEVILLE.

In less than three years, the invention had been implemented industrially in North America and inEurope. In November 1888 aluminum was first produced commercially by the electrolyticreduction of alumina by HALL and others in a company that later was to become the AluminumCompany of America. At about the same time, HROULT was associated with a company (later tobe known as Alusuisse) that operated aluminum electrowinning cells at Neuhausen, Switzerland.

KARLJ. BAYER, an Austrian chemist, was issued a patent, DE 43977, in July 1887 for animproved method of producing alumina from bauxite. Bauxite [1318-16-7], discovered byP. BERTHIER in 1821, is named for Les Baux, the village in the south of France near which it wasfirst found. With the development of a process to produce pure aluminum oxide from thisabundant ore, the technology was then complete to spur rapid growth of the aluminum industry inEurope and North America in the last decade of the 19th century. The increase in the productionof primary aluminum during the past hundred years is shown in Figure 2 [19].

4.2. Raw MaterialsCarbon. In the industrial electrowinning (separation by electrolysis) of aluminum, part of theenergy for reducing alumina is supplied as electricity and part comes from consumption of thecarbon anode. Carbon is also used as the cathode lining. Because 0.4 0.5 kg of anode isconsumed for each kilogram of aluminum produced, this represents the major carbon requirement.Because the ash from the carbon will contaminate either the aluminum produced or theelectrolyte, high-purity carbon is desirable. Certain impurities, such as vanadium, are particularly

Figure 2. Annual world production of primary aluminum 1854 1982 [7][Full View]

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harmful in that they catalyze air burning of the carbon. Other impurities, such as phosphorus,accumulate in the electrolyte and undergo cyclic redox reactions (partial reduction followed byreoxidation), consuming electric current without producing product. The coke residue frompetroleum refining is quite pure and, therefore, has been the major source of carbon for anodes.The structure of petroleum coke (Petroleum Coke) varies depending on the nature of thepetroleum feedstocks used at the refinery, the refinery flowstream, and the coking conditionsused. This coke produced at about 500 C requires calcining at about 1200 C to remove volatileconstituents and increase its density before it is blended into the anode mix. After calcination, thecoke is ground and mixed with crushed spent anodes and sufficient coal-tar pitch to allowmolding into anode blocks by pressing or by vibrating. They are baked at 1000 1200 C,causing the pitch to carbonize, forming strong carbon blocks. These blocks are made with one ormore sockets into each of which is fastened a steel stub by pouring cast iron around it. Thesestubs both conduct electric current into the anode and support the anodes in the cell. The cost ofprebaked carbon anodes in the United States was about $ 0.40/kg ($ 0.15/ pound) in 1983.

Anthracite has been the major constituent in the cell cathode blocks, although graphite andmetallurgical coke have been used to some extent. The anthracite is calcined at 1200 C or higher,

crushed and sized, mixed with coal-tar pitch, molded into blocks, and baked. These blocks,mortared together with a carbonaceous seam mix, form the pot lining, which is the container forboth the aluminum and the electrolyte. High purity is not as important for the cathode blocksbecause leaching of impurities is very slow. Consumption of cathode carbon amounts to 0.02 0.04 kg of carbon per kilogram of aluminum produced. The life of a pot (typically 2 6 years)

generally is terminated by failure of the carbon pot lining.

Aluminum Oxide. Depending on its purity and losses in handling, 1.90 1.95 kg of alumina areconsumed in producing 1 kg of aluminum. The cost of alumina at United States smelting facilitieswas $ 0.20 0.26/kg ($ 0.09 0.12/ pound) in 1983. The preparation of metallurgical aluminaand its required physical and chemical properties are discussed underAluminum Oxide.

Electrolyte Materials. The electrolyte for electrowinning aluminum is basically a solution ofaluminum oxide in cryolite [15096-52-3]. The presence of cryolite is essential for dissolution ofalumina. Cryolite usually comprises more than 75 % of the electrolyte which typically alsocontains calcium fluoride (4 8 %), excess aluminum fluoride (5 15 %), alumina (1 6 %), andsometimes lithium fluoride [7789-24-4] (0 5 %) and magnesium fluoride [7783-40-6] (0 5 %).These additives lower operating temperature and increase current efficiency.

The mineral cryolite is the double fluoride of sodium and aluminum and has a stoichiometry verynear the formula Na3AlF6 and a melting point of about 1010 C. It has been found in substantial

quantities only in Greenland, and was mined extensively there in the early 20th century but now is

essentially exhausted. Synthetic cryolite can be produced by reacting hydrofluoric acid with analkaline sodium aluminate solution:

Cryolite also can be recovered from used pot linings. The lining is crushed and treated with dilutesodium hydroxide solution to dissolve fluorides. After being filtered, the solution is neutralizedwith carbon dioxide to precipitate the cryolite.

Cryolite is produced directly in reduction cells by reaction of the soda impurity in the feedalumina with added aluminum fluoride (Fluorine Compounds, Inorganic):




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Electrolyte generated by the above reaction must be tapped from the cells periodically. In modern

smelters with dry scrubbing equipment for fume treatment and cell lives greater than 3 years,cryolite is a byproduct rather than a raw material in producing aluminum.

Synthetic cryolite could be purchased in the United States for $ 500 600/t in 1983.

Aluminum Fluoride. Aluminum fluoride, AlF3 [7784-18-1], may comprise as much as 15 wt % of

the electrolyte in excess of the amount represented by the cryolite composition. Aluminumfluoride is consumed during normal operation by three major mechanisms. First, losses ofaluminum fluoride by vaporization are appreciable; the most volatile species present in theelectrolyte is sodium tetrafluoroaluminate [13821-15-3], NaAlF4, having a partial pressure of

200 600 Pa over the operating melt, depending on composition and temperature (Fluorine

Compounds, Inorganic). Second, aluminum fluoride is depleted by hydrolysis:

And finally, aluminum fluoride is consumed by reaction with the soda present in feed alumina(Eq. 1).

Fume capture and scrubbing efficiencies have improved aluminum smelters. Fluoride previouslylost by vaporization of NaAlF

4

and hydrolysis of bath is now almost completely recycled to the

cells. Nevertheless, aluminum fluoride consumption amounts to 0.02 0.04 kg AlF3 per kilogram

of aluminum product. In 1983 technical anhydrous aluminum fluoride could be purchased in theUnited States for about $ 350/t.

4.3. HallHroult Cell for Aluminum ProductionAll commercial production of aluminum today is done in HallHroult cells. Employing prebakedcarbon anodes or self-baking Sderberg anodes. The HallHroult cell with prebaked anodes isshown in Figure 3. Essentially pure alumina is fed into the previously discussed cryolite baseelectrolyte. Electric current deposits aluminum into a pool of molten aluminum held under theelectrolyte in the carbon lined cavity of the cell. Oxygen from the alumina deposits

electrolytically onto the carbon anode dipping into the electrolyte and reacts with (burns) theanode. Cells typically range from 9 to 12 m long, 3 to 4 m wide, and 1 to 1.2 m high. Thermalinsulation surrounds the carbon lining of the cell to control heat losses. Although carbon is thematerial known to withstand best the combined corrosive action of molten fluorides and moltenaluminum, even carbon would have a very limited life in contact with the electrolyte at the sidesof the cell were it not protected by a layer of frozen electrolyte. The thermal insulation is adjustedcarefully to maintain a protective coating on the walls but not on the bottom, which must remainsubstantially bare for electrical contact. Steel collector bars in the carbon cathode conduct electriccurrent from the cell. These bars are inserted into holes that have been sized carefully so thatthermal expansion forms a tight electrical contact, or cemented in place with a carbonaceouscement containing metal particles, or bonded in place with cast iron.

(1)

Figure 3. HallHroult cell with prebaked anodes




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The electrical resistivity of prebaked anodes ranges from 0.005 to 0.006 cm. Current density atthe anode face is 0.6 1.3 A/cm2.

The Sderberg anode (Fig. 4) uses a premixed paste of petroleum coke and coal-tar pitch. Thismixture is added at the top of a rectangular steel casing that is typically 6 8 m long, 2 m wide,and 1 m high. Heat from the electrolyte and heat from the electric current passing through theanode bakes the carbonaceous mix as it progresses through the casing.

The baked portion extends past the casing and into the molten electrolyte. Baked mix replacesanode being consumed at the bottom surface. Electric current enters the anode through either

vertical or sloping steel spikes. These spikes are pulled and reset to a higher level as theyapproach the lower surface. Sderberg anodes have an electrical resistivity about 30 % higherthan that of prebaked anodes. Because of the resulting lower power efficiency and the greaterdifficulty in collecting and disposing of baking fumes, Sderberg anodes are being replaced byprebaked anodes, even though the former save the capital, labor, and energy required tomanufacture the latter.

4.3.1. ElectrolytePure cryolite melts at about 1010 C. Alumina and other additives lower the melting point,allowing operation at 940 980 C. The cryolite-aluminum fluoride-alumina system (Fig. 5) [20]has binary eutectic points at 961 C and 694 C and a ternary one at 684 C. Calcium fluoride and

lithium fluoride further reduce the liquidus temperature (Figs. 6 [21] and 7 [22]). Calcium fluorideis seldom added intentionally. Because of a small amount of calcium oxide impurity in thealumina, it attains a steady-state concentration of 3 8 % in the melt. At this level calcium iscodeposited into the aluminum and emitted in the off-gas at a rate equal to its introduction.Magnesium fluoride accumulates to 0.1 0.3 %, in the electrolyte by the same mechanism ascalcium fluoride. Some operators add up to 5 % MgF2 because it expels carbon dust from the

electrolyte by decreasing the electrolyte's ability to wet carbon.

a) Carbon anode; b) Electrolyte; c) Insulation; d) Carbon lining; e) Current collector bar;f) Thermal insulation; g) Steel shell; h) Carbon block; i) Ledge; j) Crust; k) Aluminacover; l) Removable covers; m) Anode rods; n) Fume collection; o) Air cylinder; p)Feeder; q) Current supply; r) Crust breaker[Full View]

Figure 4. Aluminum electrolyzing cell with Sderberg anode

a) Manifold gas; b) Steel shell; c) Current collector bars; d) Frozen ledge; e) Moltenelectrolyte; f) Coke and tar paste; g) Current supplying pins[Full View]

Figure 5. The Na3AlF6AlF3Al2O3 system [20]

[Full View]




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Calcium fluoride, in addition to lowering the liquidus temperature, decreases the vapor pressureand solubility of reduced species in the electrolyte for better current efficiency. Detrimentally, itlowers alumina solubility and electrical conductivity and increases density, viscosity, and surfacetension of the electrolyte.Lithium fluoride, in addition to lowering the melting point, decreasesthe vapor pressure, density, reduced species solubility, and viscosity; it also increases electricalconductivity. The only negative effect appears to be lowered alumina solubility. Its high cost,however, requires that its benefits be weighed against the price.Aluminum fluoride decreasessolubility of reduced species and lowers surface tension, viscosity, and density. It has theundesirable effects of decreasing alumina solubility and electrical conductivity and increasingvapor pressure. Aluminum fluoride acts as a Lewis acid with sodium fluoride acting as a Lewisbase. Neutrality has been defined arbitrarily as a molar ratio of sodium fluoride to aluminumfluoride of 3 : 1. Control of electrolyte acidity or the NaF : AlF 3 molar ratio, referred to as the

cryolite ratio (Rc), is of importance to cell operation. Lithium fluoride is a slightly weaker Lewis

base than sodium fluoride. Magnesium fluoride and calcium fluoride are weak Lewis acids.

Ionic Structure of the Melt. There is general agreement that molten cryolite is completely ionizedto sodium ions and hexafluoroaluminate ions:

Also it is well established that the hexafluoroaluminate ion dissociates further:

AtRc = 3, [AlF6]3 is about 30 % dissociated [23]. Raman spectroscopy has showed that the

dissociation increases with decreasing cryolite ratio to complete dissociation atRc = 1 [24].

The nature of the species formed when alumina is added is not so well established. Combinationof the results of cryoscopic measurements, Raman spectrographic data, and equilibrium studies

Figure 6. The Na3AlF6Al2O3CaF2 system [21]

[Full View]

Figure 7. The Na3AlF6Li3AlF6Al2O3 system [22]

[Full View]




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and vapor pressure measurements (reviewed in [23]) leads to the conclusion that [Al2O2F4]2 and

[Al2OF6]2 are the two major oxygen-containing ions in the melt. Possible reactions for their

formation are:

4.3.2. Electrode Reactions

Cathode Reaction. Even though Na+ is the principal current carrier, it does not discharge at thecathode. The reversible electromotive force for the formation of liquid aluminum is about 0.24 Vlower than that for the formation of sodium gas at 101.3 kPa (1 atm) for the range of compositions

and temperatures used industrially. BOWMAN [25], using cyclic voltammetry, stationary electrodepolarography, differential pulse polarography, and chronopotentiometry, found evidence only fora reversible three-electron transfer process. There was no evidence for a chemical reaction, eitherpreceding or following the electron-transfer process. This ruled out discharge of sodium at lowactivity followed by a chemical reaction to form aluminum, and also eliminated dissociation ofthe [AlF6]

3 or [AlF4] to form Al3+ and F ions in the double layer preceding charge transfer.

ROLIN, et al. [26] believed this latter process does take place. If so, dissociation must be too rapidto be detected by BOWMAN's techniques.

The cathode overvoltage can be represented by [27]:

whereR is the gas constant, Tis the temperature (K),Rc is the mole ratio NaF : AlF3, Fis the

Faraday constant, and i is the electrode's current density (in A/cm2).

Although this relationship mathematically looks like activation overvoltage, actually the cathodeovervoltage is caused by an increase in the NaF : AlF3 ratio at the aluminum surface [28]. Sodium

ions carry the current while complex aluminum anions discharge. This requires a diffusional fluxof [AlF6]

3 and [AlF4] ions to the cathode interface with a similar diffusional flux of Na+ and F

ions away from the interface. Using Fick's first law, the resulting concentration gradients werecalculated and good agreement was found between the measured overvoltages and theelectromotive forces between the two aluminum half-cells, one containing electrolyte of the bulkcomposition and the other, electrolyte corresponding to the calculated interfacial composition.

Anode Reactions. The primary anode reaction can be written:

However, O2 ion is not present in the bulk electrolyte; instead, oxygen is present as structurallylarge complexes, i.e., [Al2OF6]

2 and [Al2O2F4]2. Thermodynamically, oxygen depositing onto

(2)




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carbon at 950 1000 C should equilibrate to about 99 % CO and 1 % CO 2. However, based on

either net carbon consumption or the volume of gas produced, the primary anode product isessentially all CO2. The high anodic overvoltage implies that reaction kinetics cause this

surprising displacement from thermodynamic equilibrium. Rotating disk [29] and impedance [30]measurements indicate that there is a small diffusional overvoltage, probably caused by reaction

within the pores of the electrode. Using the general treatment for heterogeneous reaction control[31], overvoltage data can be expressed by the relationship:

where v is the number of executions of rate-controlling steps to produce one overall step, p is thereaction order, n is the number of electrons transferred in one overall step, and i0 is the reactionlimiting current density.

The reaction order,p=0.57, was found from measurements of overvoltage versus current density[32]. In industrial practice, the reaction order ranges from 0.4 to 0.6, varying with carbonreactivity and porosity. The reaction-limiting current density, i0, goes from 0.0039 to0.0085 A/cm2 as alumina concentration varies from 2 to 8 wt %. The reaction order also has beendetermined [32] from the rate of change of the limiting current with reactive species concentrationand a similar value obtained:

Measurements [33] of the ordinary combustion of graphite have shown that when oxygen reactsboth in pores and on the surface, a chemical reaction of approximately half-order results. Linearsweep voltammograms showed voltage peaks with increasing current density corresponding todischarge of CO2, COF2, and CF4 [34].

The following anode reaction mechanism is consistent with these observations. Oxyfluoride ionsdissociate within the double layer to oxygen ions:

Oxygen ions discharge upon the carbon surface (surf) forming an activated complex, CO:

(3)

(4)




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The activated complex converts very slowly to CO(g) through an adsorbed (ads) intermediate:

Carbon burning in pure oxygen at 940 970 C has a combustion rate equivalent to between 0.1and 0.2 mA/cm2. One would expect this reaction to proceed at a similarly slow rate. As availablesurface sites become covered with C*O, oxygen is deposed at higher energy (overvoltage) onto acarbon site already bonded to oxygen, producing unstable [C*O2]

, which breaks carbon to

carbon bonds almost immediately as it is formed:

The adsorbed CO2

quickly desorbs as CO2

(g). This mechanism explains both the high anodic

overvoltage and the primary production of CO2 instead of the thermodynamically favored CO.

When there is insufficient alumina in the electrolyte, the cell experiences a phenomenon calledanode effect. Bubbles grow larger and larger on the anode until the electrolyte no longer wets theanode. With a constant potential the current falls to a low value; but with the constant currentsource used industrially, the cell potential rises to 30 V. Current then penetrates the gas film bya multitude of small electric arcs or sparks. The gas produced at the anode changes from CO2 to

CO, with significant quantities (3 25 %) of CF4 and minor amounts of C2F6 generated.

Fluorocarbon compounds most likely deposit on the surface of the anode and trigger the anodeeffect. As alumina is depleted, anode overvoltage increases. At about 1.2 V anode overpotential,sufficient thermodynamic activity of fluorine is produced to cause fluorine to bond to the carbon.Even though these low-surface-energy carbon-fluorine compounds decompose on the surface toCF4 and C2F6 at cell temperature, their rate of formation can exceed the rate of thermal

decomposition, producing high coverage. Once the cell is on anode effect, restoring the aluminaconcentration is not sufficient to return it to normal operation. The gas film must be broken bysplashing aluminum, by interrupting the current momentarily, or by lowering the anodes toexpose new areas not contaminated with fluorine.

4.3.3. Current EfficiencyAccording to Faraday's law, 1 kA h of electric current should produce 0.3356 kg of aluminum, but

only 85 95 % of this amount is obtained. The principal loss mechanism is recombination ofanodic and cathodic products. Reduced species go into solution in the electrolyte at thealuminum-electrolyte interface (Fig. 8). Disagreement exists over whether the dissolved species ismetallic sodium, monovalent aluminum, subvalent sodium, colloidal aluminum, or a combination




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of these species. Equation (4) produces a thermodynamic activity of sodium in the melt, whereasEquation (5) produces an activity of aluminum monofluoride.

Metal going into solution must first diffuse through the metal-electrolyte boundary layer (Fig. 8).The metal is then transported by convection to the vicinity of the anode. Here it reacts with carbondioxide. Chemical reactions appear to be fast compared to mass transport. The rate-controllingstep was previously assumed to be diffusion of dissolved metal through the boundary layer at themetal-electrolyte interface. However, ref. [35] indicates mixed control with diffusion at both thealuminum-electrolyte interface and the bubble-electrolyte interface being important.

There are several mechanisms accounting for additional minor losses in current efficiency. Newcell linings absorb sodium, with Equation (4) maintaining an equilibrium activity of sodium.Fortunately, the lining saturates early in the cell's life but until this occurs, current efficiency is

low. When a metal dissolves in a molten salt it usually imparts electronic conductivity to the melt,thereby lowering current efficiency. Some investigators have found a small electronicconductance for cryolite-base melts but others have not. Such elements as phosphorus andvanadium, which can be reduced partially at the cathode and then reoxidized at the anode, lowerefficiency.

4.3.4. Cell VoltageThe voltage of a HallHroult cell is made up of a number of components:

whereE0 is the thermodynamic equilibrium voltage described under ThermodynamicConsiderations (Section Thermodynamic Considerations); CA is the concentration overpotential

at the anode; SA is the surface overpotential at the anode described by Equation (3); CC is the

concentration overpotential at the cathode described by Equation (2); SC is the surface

overpotential at the cathode, generally negligible;Iis the total cell current;RA is the electrical

resistance of the anodes or anode;RB is the effective resistance of the bath, allowing for fanning

out of current as it flows from anode to cathode and the increased bath resistivity caused by gasbubbles (for details, see 27);RC is the cathode resistance; andRX is the resistance external to the

cell but included in calculating power consumption.

The anode concentration overpotential can be calculated by:

(4)

(5)

Figure 8. Current efficiency loss by metal reoxidation[Full View]




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been refined to the point where cells of over 2.5105 A have been designed and operated withhigh efficiency. Natural convection caused by temperature gradients or composition differences inthe electrolyte are insignificant compared with the movement induced by magnetic forces and gasbubbles. Gas bubbles produce significant stirring in the electrolyte and are the dominant force inbath movement, whereas electromagnetic forces predominate in metal pad movement.

4.4. Thermodynamic ConsiderationsThe thermochemistry of constituents of the electrolyte and of the electrolyte-aluminum system istreated in [23]. Thermodynamic data used in the following section are from theJANAF(JointArmy-Navy-Air Force) Thermochemical Tables [39] and the 1978 supplement to the tables [40].

For pure -alumina, reduction to aluminum can be represented:

In the industrial production of aluminum, a portion of the energy is supplied by the combustion ofcarbon anodes to carbon dioxide. The overall cell reaction can be represented:

Additional energy is required to heat alumina and carbon from room temperature to operatingtemperature. The alumina fed to the cell usually is not pure -alumina. For -alumina, the overallprocess can be represented:

The enthalpy requirement for transformation of the reactants at 25 C (298 K) to products at 960 C (1233 K) is:

This value corresponds to a theoretical energy requirement of 6.25 kW h per kilogram ofaluminum produced.

The reversible cell potential (decomposition voltage) for Reaction (6) can be calculated from therelation:

(6)




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where is the Gibbs free energy change for Equation (6), 3.450105 J; n is the number of

electrons per unit cell reaction, i.e., 3; Fis the Faraday constant, 96.487 kJ V1 mol1; andis the standard electrode potential, in volts, at 960 C for the reaction with all reactants

and products at unit activity. Substituting the appropriate values into Equation (7) and solving forthe cell potential:

The above decomposition potential applies to electrolyte saturated with alumina at 950 C. The

decomposition potential from melts with alumina activity less than unity can be calculated [27]from the equation:

The activity of alumina in cryolite-base melts can be approximated [27] from the equation:

where is the concentration of alumina in the electrolyte, in wt %, and is thesaturation concentration of alumina, in wt %.

4.5. Alternate ProcessesAlthough the HallHroult process has gained industrial dominance, it has several inherentdisadvantages. The most serious are the large capital investment required and the highconsumption of costly electrical power. There are also the costs of the Bayer alumina refiningplant and of the carbon anode plant. Many of the aluminum-producing countries must importalumina or bauxite. The supply of petroleum coke is limited. These deficiencies have spurredresearch to find alternate processes.

The earliest commercial process for producing aluminum was sodiothermic reduction ofaluminum halides. Reduction of aluminum chloride by manganese has been investigated [41].However, neither can compete with the HallHroult process. Many attempts have been made atdirect carbothermic reduction of alumina butthese have resulted in very low yields, owing to theformation of solid aluminum carbide, aluminum suboxide vapor, and aluminum vapor that reactswith carbon monoxide as the temperature drops on the way from the furnace. Yields as high as67 % can be obtained by staging the reactions to produce aluminum carbide at 1930 2030 C,which then reacts with alumina to produce aluminum and carbon monoxide at 2030 2130 C[42].

Better yields result from adding to the furnace a metal (or a metal oxide that is subsequentlyreduced to a metal), such as iron, silicon, or copper, to alloy with the aluminum and lower itsvapor pressure. Of course it is then necessary to extract the aluminum from the alloy. In principlethis can be accomplished by electrolytic refining, by fractional crystallization, or by monohalide

(7)




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distillation. In the latter process the aluminum extraction takes place at 1000 1400 C:

The AlCl gas is transported to a cooler zone, 600 800 C, where pure aluminum is formed:

These combined processes to date have proved noncompetitive. Selective solution of aluminumfrom the alloy by using a volatile metal, such as mercury, lead, bismuth, cadmium, magnesium, orzinc, has been investigated. Following extraction, the volatile metal is distilled, leaving purealuminum. If FeAl3, TiAl3, or Al4C3 is formed, neither electrolysis nor volatile metal extraction

will remove the aluminum from the compound.

The aluminum vapor pressure can be lowered also by alloying the aluminum with aluminumcarbide. Over 40 wt % aluminum carbide is soluble in aluminum at 2200 C [ 43]. Thisobservation led to a process in which an alloy of aluminum and aluminum carbide was producedat 2400 C. When this alloy was tapped from the furnace and allowed to cool slowly, thealuminum carbide crystallized into an open lattice, the interstices of which were filled with purealuminum. Pure aluminum was then removed either by leaching with molten chloride fluxes or byvacuum distillation. The aluminum carbide residue was recycled to the arc furnace. Alternately,the aluminum carbide could be distilled destructively above 2200 C to produce aluminum and aresidue of pure graphite.

In a joint project with the U.S. Department of Energy concluded in 1983, Alcoa investigatedproducing aluminum-silicon alloy carbothermally in a blast furnace. The required hightemperature was obtained by using a pure oxygen blast. Using a low-pressure blast furnace toproduce an aluminum-silicon alloy was found to be infeasible because of severe bridging andinterruption of the burden movement caused by total reflux of Al, Al2O, and SiO vapors [44],

[45]. Addition of iron improved operation but required an improved technique for extracting thealuminum from the dilute aluminum alloy. An alloy having higher aluminum content could beobtained if a significant portion of the process energy was supplied by electrical power [46].Japanese researchers are continuing to explore the aluminum blast furnace concept aiming for alow-silicon, high-iron alloy [47]. This improves the blast furnace efficiency but complicates

extraction of aluminum from the alloy. They propose to overcome this difficulty by extracting thealuminum with lead [48].

In 1976 Alcoa described a smelting process wherein aluminum chloride, dissolved in a moltensodium chloride-lithium chloride electrolyte, was electrolyzed in a bipolar electrode cell toproduce aluminum and chlorine. The chlorine was recycled to a fluid-bed chemical reactor, whereit reacted with alumina, pyrolytically coated with carbon from fuel oil, to produce aluminumchloride, carbon dioxide, and carbon monoxide. This reaction was highly exothermic. Thealuminum chloride was desublimed to separate it from the gas and recycled to the cell to producemore aluminum and chlorine. This process required 30 % less electrical power than their bestHallHroult cells. Because the entire process is a closed system, it is also environmentally more

attractive than the HallHroult process. Problems with the chemical plant and low demand foraluminum caused the pilot plant to be shut down temporarily in 1982. Research on the plant wascontinuing in 1984.



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