Aluminum Production

CHAPTER 2.5

Aluminum Production Alton T. Tabereauxa and Ray D. Petersonb a

Alcoa Primary Metals, Muscle Shoals, Alabama, USA Aleris International, Rockwood, Tennessee, USA

b

NOMENCLATURE a activity (thermodynamic concentration) db bubble layer (diameter) thickness (cm) CE current efficiency (%) E
standard equilibrium potential or standard EMF (V) EBemf back electromotive force or counter electromotive force potential (V) Eext external cell energy losses (kWh/kg Al) Eint total internal cell reaction and heat loss energy (kWh/kg Al) Epol electrode polarization potential (V) Erev reversible equilibrium voltage or Nernst potential (V) F Faraday’s constant (96,485 kJ/t equivalent) g gaseous state
DGreac Gibbs energy of reaction at standard conditions (kJ/mol Al2O3) DG Gibbs energy of reaction at nonstandard conditions (kJ/mol Al2O3)
DHtot standard enthalpy of reaction (kJ/mol Al2O3) I cell current or line current (kA) ia anode current density (A/cm2) ic anode limiting current density (A/cm2) icc anode critical current density (A/cm2) IRan electrical resistance of the anode (O) IRbub electrical resistance of the gas bubbles (O) IRca electrical resistance of the cathode (O) IRel electrical resistance of the electrolyte (O) IRext electrical resistance of the external bus bar system (O) K parabolic rate constant (mg/cm2/h) К electrical conductivity of the electrolyte (1/Ocm) m mass of oxide film (mg/cm2) n number of electrons transferred in the reaction as written R universal gas constant ( J/mol K) T oxidation time (h) T temperature of the electrolyte (K) Tb temperature electrolyte (
C) Van anode voltage drop (V) Vbub gas bubble voltage drop in the electrolyte in the anode-to-cathode distance (V)

Treatise on Process Metallurgy, Volume 3 http://dx.doi.org/10.1016/B978-0-08-096988-6.00023-7

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Vca cathode voltage drop (V) Vcell total cell operating voltage (V) Vel electrolyte voltage drop (V) Vext external cell voltage drop between cells (V) VRI ohmic voltage drop (V) xA fraction current efficiency, x ¼ %CE/100

GREEK a alpha crystalline state of alumina g gamma crystalline state of alumina hca concentration overvoltage at anode (V) hcc concentration overvoltage at cathode (V) hsa surface reaction overvoltage at anode (V) hsc surface reaction overvoltage at cathode (V) r density (kg/m3) f fraction of anode covered by bubbles

SUBSCRIPT an anode Bemf back electromotive force bub bubbles ca cathode el electrolyte ex external g gas state l liquid state RI resistance and current s solid state

SUPERSCRIPT b bulk value e equilibrium rev reversible

Aluminum is one of the most commonly used metallic elements. Due to its high corrosion resistance and mechanical strength to mass ratio, aluminum alloys are used as a major structural material in aircrafts, buildings, machinery parts, beverage cans, and food wraps. Aluminum is the most recyclable of all materials. It is four times more valuable than any other recycled consumer materials and there is no limit to the number of times aluminum can be recycled. Recycling aluminum requires only 5% of the energy used to make primary aluminum, as recycled metal requires only 2.8 kWh/kg of metal produced while primary aluminum production requires about 13 kWh/kg.

Aluminum Production

2.5.1. HYDROMETALLURGY OF THE BAYER PROCESS Aluminum production is accomplished in two processes: the Bayer process for refining bauxite ore to obtain aluminum oxide, or alumina, and the Hall–He´roult electrolytic process for smelting the alumina dissolved in cryolite to produce pure aluminum metal. The Bayer process marks the beginning of the modern field of hydrometallurgy. It is the principal industrial means of refining bauxite to produce alumina (Al2O3) that is pure enough for aluminum electrolysis. Karl Bayer invented the process in 1887 when working at the Tentelev chemical plant near Saint Petersburg in Russia. He improved the process for manufacturing alumina for the textile industry where it was used either as a substance that fixes dyes or as a mordant in dyeing cotton. Bayer discovered that alumina contained in bauxite could be dissolved selectively by heating it with a solution of sodium hydroxide under pressure in an autoclave to form sodium aluminate solution. Then aluminum hydroxide could be precipitated from an alkaline solution when seeded with alumina hydrate to act as a nucleus. The precipitated hydrate is easily filtered, washed, and calcined to produce high-purity crystalline alumina, and the caustic alkaline liquor is recycled. Today, the industrial Bayer process is virtually unchanged and nearly all the world’s alumina supply, over 80 million tons in 2011, for aluminum production is derived from it.

2.5.1.1. Impact of Different Bauxites on the Bayer Process The Bayer process is basically used for the extraction of aluminum hydrate from the bauxite ores with the mass ratio of alumina to silica (A/S) above 9. The sinter process is widely used to process the poor-grade diasporic bauxite ores with A/S below 7, in China and Russia, by sintering the bauxite ore with sodium carbonate and limestone to form sodium aluminate and calcium silicate. In the process, the sinter is then leached with water, caustic soda solution, and recycled liquor to dissolve the soluble sodium aluminate. The resulting slurry is then filtered and the precipitation of hydrate is promoted by bubbling carbon dioxide gas into the supersaturated sodium aluminate solution, i.e., the carbonation process. Meanwhile, the caustic in the solution is recycled to extract alumina from the bauxite in the sintering process. When operated in conjunction with the Bayer process to recover alumina and soda from red mud, it is called the combination process. High-temperature sinter processes were developed in Russia for treating high-silica nepheline concentrates by sintering with limestone for the production of cement, soda, potash, and alumina. Before the 1970s, alumina was produced in two different grades, i.e., floury and sandy. During the past decades, the sandy alumina, with low fine content (90% greater than 45 mm), was more desirable because of the increased concerns of the environmental influence and energy costs. Sandy alumina can be obtained by the Bayer process, with

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adjustments in the precipitation step. However, it is difficult to produce sandy alumina with a low fine particulate content and narrow size distribution by the sinter process. Ores that contain high concentrations of aluminum hydroxide minerals, >35%, are called bauxite. It is the raw material for almost all production of alumina. Bauxite occurs in three main forms depending on the number of molecules of water of hydration and the crystalline structure. The three structures are gibbsite, bo¨hmite, and diaspore. The major difference between them is that bo¨hmite and diaspore have a different crystalline structure that requires even higher temperatures and pressures for complete dehydration than gibbsite as shown in Table 2.5.1. Gibbsitic ores are plentiful in a number of countries such as Australia, Brazil, Guinea, Guyana, India, Jamaica, Surinam, and Venezuela. Recent discoveries of bauxite have been found in Cambodia, Saudi Arabia, and Vietnam. More than 99% of the bauxite ores in Russia and China are bo¨hmitic and diasporic ores that are characterized by challenging processing demands, high alumina and silica content, and low A/S mass ratios. The energy usage for the low temperature digestion of gibbsite bauxite is 7.5–12 GJ/t, and the high-temperature digestion of bo¨hmite and/or diaspore bauxite is 11–18 GJ/t. Table 2.5.1 Composition and Process Differences in Bauxites Ores Unit

Gibbsite (Hydragillite)

Böhmite (Gibbsite– Böhmite Mixtures)

Diaspore

Process

Bayer

Bayer

Soda-Lime Sinter

Composition

a-Al2O33H2O

a-Al2O3H2O

b-Al2O3H2O

Alumina content

%

45–65

47–85

47–85

Silica

%

1–5

2–5

4–16

19

12

4

Monoclinic

Orthorhombic

Orthorhombic

2.42

3.01

3.44

150

200–250

450

Alumina/silica ratio Crystal system Density Temperature, rapid dehydration

3

g/cm


C

Na2O in liquor

g/l

120–150

205–245

240–360

Pressure

MPa

1

1

3.5

Energy

GJ/t

7.5–12

11–16

34–45

Sandy alumina, lower soda loss, normal organics

Floury alumina, high soda loss with DSP, high organic oxalate

Floury alumina, high fines, high 45 mm

Comments

843

Aluminum Production

A large amount of alumina has to be produced in order to produce aluminum in the Hall–He´roult process; it requires about 4 t of bauxite to produce 2 t of alumina and it takes almost 2 t of alumina to produce 1 t of aluminum metal. The two largest operating costs per ton of alumina in Bayer plants are bauxite and fuel.

2.5.1.2. Bayer Process The purification of bauxite to produce alumina in the Bayer process consists of five steps as illustrated by the schematic shown in Figure 2.5.1. 2.5.1.2.1 Crushing, Mixing, and Desilication Bauxite ore is first mechanically pulverized and milled to reduce the particle size and then screened. The crushed ore is mixed with the process liquor containing caustic soda and processed in a grinding mill to produce a slurry containing 35–40% solids. Sodium hydroxide is the main chemical used in the process to extract aluminum hydrate from 1. Mixing

Refining bauxite to alumina Caustic soda solution

Slurry mixer Digester

5. Calcination Calcining kiln Precipitators

2. Digestion Thickener Filter

Pressure reducer and heat exchanger

3. Clarification 4. Precipitation Settling tank

Filter

Red mud to disposal pond

Figure 2.5.1 The Bayer process for the production of alumina from bauxite.

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bauxite. Prior to digestion, the soluble silica is removed from the process solution or slurries by an initial desilication step in which the slurry is kept near atmospheric boiling for several hours to encourage the precipitation of the solubilized silica to the desilication product (DSP) at lower temperatures, 90–100
C, to reduce excessive scaling in the heat exchangers. The DSP 3Na2O3Al2O35SiO2X (where X represents a variety of inorganic anions, sulfate, carbonate, chloride, aluminate, hydroxide, as well as organics) formed is the main source of soda loss in 80–90% of the reactive silica into the DSP, with the remainder being transferred in the digestion step. The loss of soda from the liquor is a major economic penalty, and it is linear with the reactive silica content in bauxite. For this reason, bauxites containing >8% reactive silica by weight are usually considered to be uneconomical for processing.

2.5.1.2.2 Digestion Digestion conditions are tailored to the aluminous phase distribution in the bauxite. Digestion times are often determined not by the kinetics of gibbsite dissolution but by the kinetics of residual DSP formation and, more importantly, the reduction of the soluble silica in solution. The slurry of aluminum-bearing minerals in bauxite is pumped into a digester where the aluminum-containing compound, sodium aluminate (NaAlO2), is solubilized and extracted by washing with a solution of hot concentrated sodium hydroxide at 110–270
C under pressure (depending on the type of bauxite). These conditions are maintained for a time ranging from half an hour to several hours. The time of digestion is often not dictated by alumina dissolution kinetics, but kept to a minimum to avoid excessive dissolution of quartz which provides additional reactive silica. The addition of 1–2% CaO, or lime, facilitates the extraction from ores containing >8% bo¨hmite and diaspore. Digestion converts the aluminum-bearing minerals in bauxite to aluminum hydroxide, Al(OH)3, which behaves in an amphoteric manner, neutralizing a base as it dissolves in the strong caustic solution of sodium hydroxide at high temperature and pressure according to the chemical equations: Gibbsite: Al2 O3 3H2 OðsÞ þ 2NaOHðaq:Þ ¼ 2NaAlO2 ðaq:Þ þ 4H2 OðlÞ

ð2:5:1Þ

Bo¨hmite and diaspore: Al2 O3 H2 OðsÞ þ 2NaOHðaq:Þ ¼ 2NaAlO2 ðaq:Þ þ 2H2 OðlÞ

ð2:5:2Þ

The hot slurry containing the sodium aluminate solution passes through a series of flash tanks that reduce the pressure and recover heat that can be used in the refining process. Bauxite is charged to digestion to achieve a specific alumina supersaturation to ensure stability through these trains of flashing and clarification stages.

Aluminum Production

2.5.1.2.3 Clarification The slurry is pumped into a tank where the solution is clarified as the insoluble oxide impurities will not dissolve in the caustic and are separated from the pregnant liquor containing sodium aluminate by a process known as settling. The solution is cooled, filtered, and the coarse particles above 100 mm are removed in a settling tank or cyclone and sent to a long-term storage containment area. This bauxite residue called “red mud” that accumulates in the bottom of the tank consists of fine sand, iron oxide, and oxides of trace elements like titanium. Polymer flocculent filter aids are added to accelerate the settling and thickening and also to improve the density of the underflow solids of the bauxite residue from the thickeners. It settles out as solid slurry from the caustic suspension during clarification from the pregnant liquor. The fine particulate trapped on filters is washed to recover alumina and caustic soda that can be reused. 2.5.1.2.4 Precipitation The cooled (from about 100 to 75
C) filtered liquor is pumped through a series of sixstory tall precipitation tanks. Solid seed crystals of alumina hydrate from the classifiers are added to the liquor through the top of each tank. The undesired nucleation of hydrate particles (<1-mm particles) occurs in liquor. It is desired to have the seed crystals cause cementation of small aluminum hydrate crystals (1- to 45-mm particles) and grow to larger agglomerated crystals (>45-mm particles) as they settle through the liquid and dissolved alumina attaches to them. The precipitated crystalline aluminum trihydroxide (Al2O33H2O), commonly referred to as hydrate or gibbsite, is then precipitated from the cooled supersaturated liquor. The hydrate often grows into the agglomerates of either hexagonal tablets from seeded sodium aluminate solutions. Some occluded soda, Na2O, impurity is trapped in hydrate crystals after precipitation. The primary control theory for precipitation is to maximize yield (grams of alumina precipitated/liter of sodium alumina solution) while maintaining the product soda, 45 mm, and seed balance. This is achieved first by controlling the liquor fill temperature which impacts the occluded soda, superstation of liquor, fines generation, and oxalate precipitation, and second, controlling the seed charge which impacts the amount of fines in product, seed balance, and particle attrition. The white fluffy solid precipitate is filtered and then concentrated by evaporation resulting in an aluminum hydroxide filter cake. The alkaline liquor and hydrate seeds are recycled in the process. 2.5.1.2.5 Calcination Aluminum hydroxide is washed, filtered, and calcined in either a rotary kiln, stationary gas suspension calciner or fluid bed calciners to about 950–1000
C to drive off the chemically bonded water from hydrate and form anhydrous alumina crystals in the process. The industry is discontinuing the use of rotary kilns in favor of using stationary calciners,

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which consume about 33% less energy, 3.0 GJ/t compared with 4.5 GJ/t alumina. The calcination step requires 25% of the total energy in the Bayer Process. 2AlðOHÞ3 ðsÞ ¼ Al2 O3 ðsÞ þ 3H2 OðgÞ

ð2:5:3Þ

The type of alumina crystals formed at these temperatures is commonly referred to as the gamma (g-Al2O3) crystalline state of alumina. The technical pure alumina is classified as smelter grade alumina (SGA) and contains at most 0.01–0.02% SiO2, 0.01–0.03% Fe2O3, and 0.3–0.6% NaO2 as impurities. Alumina has to be calcined above 1250
C to obtain complete conversion to the alpha crystalline state of alumina (a-Al2O3) in order to make chemically inert, ceramic grade alumina. The largest waste product generated in bauxite refining is the tailings or bauxite residue. A Bayer refinery produces about the same amount of red mud as it does alumina, in terms of dry weight. It settles out as a solid slurry from the caustic suspension during clarification or it is settling out of undissolved solids. The bauxite residue slurry is then pumped to a large on-site impoundment area or a residue storage area that allows water to evaporate after several years. When the mud has dried to a solid enough consistency, it is covered with dirt or mixed with soil. Extensive efforts are being made to find solutions to reduce the quantity of residue as well as to find ways to use it in commercial products [1]. Bauxite residue represents one of the alumina industry’s most important disposal and environmental problems as there are now more than 80 alumina refinery plants in the world with bauxite residue storage areas with an estimated inventory close to 4 billion tons globally.

2.5.1.3. Alumina Properties The grade of alumina (particle size, attrition, alpha/gamma content, and surface area) are influenced by precipitation and calcining conditions, and it is usual to differentiate between the two main grades, i.e., floury alumina, which is either produced by the sinter process or is highly calcined and contains mostly a-Al2O3 alumina; sandy alumina which is produced by the Bayer process and calcined to a lesser degree contains mainly g-Al2O3. There is a clear trend now in the industry toward the production of sandy alumina. Major properties of alumina important to aluminum smelters which have the potential to impact aluminum productivity are provided in Table 2.5.2.

2.5.2. ELECTROMETALLURGY OF ALUMINUM 2.5.2.1. Introduction Aluminum is a highly reactive metal that forms a strong chemical bond with oxygen. Aluminum cannot be produced by the electrolysis of an aluminum salt dissolved in water because of the high reactivity of aluminum with the protons of water and the subsequent

Table 2.5.2 Key Characteristics of Alumina Quality That Impact the Aluminum Electrolysis Process Alumina

Property

Flowability

The ability to flow alumina powder through smelter handling systems, especially the point feeders

Solubility

Surface area

Segregation

Change in Property

Angle of repose

33–35


Particle size distribution: increases with higher fines content

Attrition index

10–15%

Particle shape and strength: increases with transport and conveying

Fines

6–8%, 45 mm

Precipitation: particle size distribution; increases with higher attrition

“Superfines”

0.5–3%, 20 mm Precipitation: particle shape and strength; increases with higher attrition

The process needs rapid dissolution of alumina in the electrolyte without sludge formation Fines and superfines

See above

See above

% Alpha content

2–15%

Increases with higher calcination temperatures

% Water content

0.65–0.9% LOI

Increases with lower calcination temperatures

A high surface area is needed for efficient adsorption and removal of fluorides in the cell off-gases in the gas cleaning system Surface area (BET)

Dusting

Typical

60–80 m2/g

Increases with low calcination temperature

A minimum of airborne fines coming off the alumina during handling and cell feeding Fines

See above

See above

Attrition index

See above

See above

LOI and water content

See above

See above

Through storage handling, fume systems, and pot feeding using volumetric dumps Loose bulk density

0.96 g/cm3

Precipitation: particle size distribution and particle shape

3

Precipitation: particle size distribution and particle shape

Vibrated bulk density 1.10 g/cm

Chemical Low impurities are required to retain the electrolyte chemistry and manufacture-specific aluminum products composition 0.30–0.45% Washing and filters: impacts the control of the electrolyte chemistry in cells % Na2O %CaO

0.01–0.04%

% Fe2O3

0.015–0.03%

% SiO2

0.01–0.03%

Digestion: affects all metal products including LME ingot grades

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formation of hydrogen gas. As in aqueous solution, protons (Hþ) are preferentially reduced before Al3þ ions, leading to hydrogen evolution. Thus, the reduction of Al3þ requires to be performed by electrolysis in a molten aluminum salt in the absence of water. The direct reduction to aluminum with carbon, as used to produce iron, is not chemically possible, since aluminum is a stronger reducing agent than carbon. There is a possible indirect process, the carbothermic reduction to aluminum using carbon and Al2O3 which forms intermediate aluminum carbide (Al4C3) that can further yield aluminum metal at a temperature of about 2000
C. This process uses less energy and yields less CO2 than the Hall–He´roult process, the major industrial process for aluminum extraction. However, this process has not yet been perfected and is still under development by companies in the aluminum industry. Therefore, aluminum must be extracted from a solution of purified alumina in a molten salt by electrolysis. The electrolytic process was discovered and patented in 1886 separately and almost simultaneously by Charles Martin Hall and Paul L.T. He´roult. Aluminum oxide is dissolved in molten cryolite, and the molten mixture is electrolyzed using carbon electrodes by passing a direct electric current through it producing pure aluminum metal. Therefore, it is called the Hall–He´roult process for manufacturing aluminum. The key discovery in the development of the aluminum process was that cryolite or sodium hexafluoroaluminate (Na3AlF6) was a suitable solvent for the dissolution and electrolysis of alumina.

2.5.2.2. Electrolyte Composition and Liquidus Temperature 2.5.2.2.1 Dissociation of Cryolite All molten salts have a high ionic character and generally disordered crystal structures where the volume has undergone an expansion of approximately 20%. The ions are free to move, and they may form ion pairs with opposite charges and electrostatic attractive forces. Sometimes these attractive forces can form weakly bound complex ions when there are mixtures of different anions or cations. Cryolite is one example of this, and when solid, the cationic lattice is only sodium ions, while the aluminum ions are ionically bonded to six fluoride ions forming the hexafluoroaluminate anion, AlF6 3 . The mechanism and the degree of dissociation of hexafluoroaluminate ions have been developed from cryoscopy, density, viscosity, and Raman spectroscopy [2]. Most investigators agree that cryolite completely ionizes to form hexafluoroaluminate (AlF6 3 ) anions, which further dissociates to form tetrafluoroaluminate (AlF4  ) as well as sodium (Naþ) and fluoride (F) ions according to the equations: Na3 AlF6 ðsÞ ¼ 3Naþ þ AlF6 3

ð2:5:4Þ

The hexafluoroaluminate ion then dissociates partly as a consequence of the melting process.

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Aluminum Production

AlF6 3 ¼ AlF4  þ 2F 2

K¼

ð2:5:5Þ

ðaAlF4  ÞðaF Þ 4a ¼
 aAlF6 3 ð1  aÞð1 þ 2aÞ2 3

where a is the degree of dissociation of the hexafluoroaluminate ions, and for the cryolite composition, it is about 0.3. The aluminum fluoride added in excess of cryolite reacts with F ions in order to form AlF4  ions according to the following reaction: AlF3 þ F ! AlF4 

ð2:5:6Þ

These ions are generally considered as “complex,” but it is more correct in the molten media to refer to them as ion pairs, ion associations, or clusters. Sometimes they are calculated via association constants. At any one instant, a small fraction will have undergone complete dissociation and there will be a small, but finite concentration of discrete aluminum cations and oxygen anions. Under the influence of a potential gradient, which exists during electrolysis of molten salts, the dissociation of the complexes is enhanced. The fact that dissociation occurs is supported by measurement of cathode polarization in aluminum electrolysis and generally it is reported as being very small. The amount of polarization can be determined from the sodium content. The oxygen-containing species has the lowest concentration, and therefore this anode is more likely to undergo concentration polarization and depletion of the electroactive species at the interface. 2.5.2.2.2 Liquidus Temperature of Cryolite Phase diagrams indicate the rate of change in the depression of the liquidus temperature relative to the composition and the degree of dissociation of ionic species of the molten salts. Pure cryolite melts congruently at 1012
C as indicated in the NaF–AlF3 phase diagram developed by Solheim and Sterten [3] shown in Figure 2.5.2. A low melting point compound chiolite (Na5Al3F14) exists at high AlF3 and melts incongruently at 734
C. A third compound, NaAlF4, is formed in the gas phase from the evaporation of cryolite. It can also be produced by chemical reactions in the solid state. Aluminum fluoride, as well as other compounds, is added to the cryolite melt to decrease the liquidus temperature and change several electrolyte properties that enhance the electrolytic process. The molten mixture of cryolite and additives in the electrolyte is commonly referred to as “bath.” The change in the liquidus temperature of cryolite versus AlF3 is nonlinear as it decreases sharply at higher AlF3 concentrations as demonstrated in Figure 2.5.2. The surplus AlF3 in molten cryolite is expressed as the weight ratio of NaF/AlF3, or alternatively as the molar ratio of NaF/AlF3. An excess concentration of AlF3 typically in the range 8–14% is maintained in the cryolite melt to improve the production efficiency of the electrolysis process. Electrical resistance within the

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1100 Pure cryolite

Temperature (°C)

1000

Liquid

900 Na3AIF6 (solid) + liquid

800

AIF3 (solid) + liquid

Na5AI3F14 (solid) + liquid

734 °C

Na3AIF6 (solid) + Na5AI3F14 (solid)

700

600 25 0

Chiolite

695 °C

Na5AI3F14 (solid) + AIF3 (solid)

30

40 32

50

60 AIF3 (mol %)

AIF3 (wt.%)

Figure 2.5.2 The excess AlF3 side of the Na–AlF3 phase diagram.

anode–cathode distance (ACD) of the electrolyte provides sufficient heat to keep the cryolite molten. The chemistry of the electrolyte determines the operating temperature of the aluminum electrolysis cells and most commercial cells operate near 960
C. The melting point of cryolite is lowered by the effect of the concentration of each additive on the electrolyte, for example: alumina ¼ 5.6
C/% Al2O3, aluminum fluoride ¼ 1.0–5.0
C/% AlF3, calcium fluoride ¼ 2.9
C/% CaF2, magnesium fluoride ¼ 3.8
C/% MgF2, and lithium fluoride ¼ 8.7
C/% LiF. Sterten and Mæland [4] determined that the addition of Al2O3 to a cryolite melt lowers the liquidus temperature of the binary mixture, shown in the Na3AlF6–Al2O3 phase diagram in Figure 2.5.3, to the eutectic temperature of 966
C at 10 wt.% Al2O3. Skybakmoen et al. [5] determined by weight loss measurements the solubility of alumina in cryolite that describes the solidus line located on the right side of the phase diagram in Figures 2.5.3 and 2.5.4 at alumina concentrations higher than 10% and developed the alumina solubility equation: ½Al2 O3 sat ¼ 11:9ðT
C=1000Þ4:8

851

Aluminum Production

1040

Temperature ( °C)

1020 Liquid 1000 AI2O3 + liquid 980 Na3AIF6 + liquid 960 Solid 940

10 15 5 Alumina concentration (wt.%)

0

20

Figure 2.5.3 Phase diagram for the system Na3AlF6–Al2O3.

990

Liquidus temperature (°C)

980 970 6 wt.% AIF3 8%

960

10%

950

12%

940

14%

930 920

16%

5 wt.% CaF2

910 0

Saturation

2

4

6 AI2O3 (wt.%)

Figure 2.5.4 Alumina saturation in the system NaF–AlF3–Al2O3.

8

10

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Alton T. Tabereaux and Ray D. Peterson

The decrease in Al2O3 solubility at higher concentrations of AlF3 is evident in the NaF–AlF3–Al2O3 shown in Figure 2.5.4. It is not practical to operate aluminum cells with AlF3 concentrations, greater than 12–13% due to the decrease in solubility of alumina in cryolite at the lower electrolyte temperatures that eventually results in undissolved alumina deposits in cells (sludge). An equation to calculate the solubility of alumina in cryolite melts containing combinations of AlF3, CaF2, MgF2, and LiF for all practical concentrations was published by Skybakmoen et al. [5] where t is given in
C and the concentrations of components are given in wt.% while the concentration of AlF3 is given in excess wt.% relative to that in cryolite.    1000 %Al2 O3 ðsatÞ ¼ exp A þ B t1

20

where A ¼ 2.464  0.007 (%AlF3)  1.13  105 (%AlF3)3  0.0385(Li3AlF6)0.74  0.032(CaF2)  0.04 (%MgF2) þ 0.0046[(%AlF3)(LiAlF6)]0.5 and B ¼ 5.01 þ 0.11 (%AlF3)  4.0  105(%AlF3)3  0.732(Li3AlF6)0.4 þ 0.085[(AlF3)(LiAlF6)]0.5. The phase diagram for the ternary mixture Na3AlF6–AlF3–Al2O3 shown in Figure 2.5.5 was first developed by Foster [6], and later it was revised by Skybakmoen et al. [5]. The dotted line represents the maximum solubility of Al2O3 in the system. There is a 1.5% difference in the temperature of the alumina–cryolite multivariant line starting at 10% alumina. The compound chiolite (Na5Al3F14) melts at 730
C. Foster

10

wt .%

Al

2O 3

1080

a - Al2O3

1060

1040 1020 1000 980

h - Al2O3

wt.% AlF3

Figure 2.5.5 Phase diagram for the system Na3AlF6–AlF3–Al2O3.

30

760 780

20

800

820

10

840

(1011)

860

Na3AlF6

880

1000

900

920

940

0

960

98

Cryolite

P¢

P(739)

E¢ Chiolite E(696)

40

Aluminum Production

found that the lowest liquidus temperature of the mixture, or ternary eutectic, was located at 37.3 wt.% AlF3, 3.2 wt.% Al2O3, and 59.5% Na3AlF6 and 684
C. Compositions containing 20 wt.% or less AlF3 precipitated a-Al2O3 in the primary phase field of alumina while those containing 25% or more precipitated Z-Al2O3. The electrolyte typically contains from 9% to 11% AlF3, 4% to 6% CaF2, and 1.5% to 4% Al2O3. The normal operating temperature of the electrolyte in aluminum cells is in the range 945–965
C and has a corresponding liquidus temperature of 945–950
C. The difference between the operating temperature and the liquidus temperature of the electrolyte is referred to as the “superheat.” Aluminum cells typically require about 5–10
C superheat in order to have sufficient energy to dissolve alumina powder in the melt. The pool of molten aluminum metal pool in the cathode is about 300
C above its freezing point, 660
C. Because the electrolyte is so close to its freezing point, a ledge of frozen electrolyte is formed attached to the sidewalls of the cells, being one of the coldest zones of cells. The formation of the side-ledge protects the cell lining from the corrosive electrolyte. If the superheat is higher than normal for an extended period, it causes the protective layer of frozen cryolite ledge on the cell’s sidewalls to melt. The ledge plays an important role in regulating the heat balance and bath chemistry control of cells: The thickness of the cryolite sidewall ledge changes according to the thermal behavior of the cells. Whenever the cell superheat increases (difference between the bath temperature and the electrolyte liquidus temperature), some sidewall ledge melts, which increases cell heat losses. Similarly, whenever the superheat decreases, some bath freezes increasing the sidewall ledge thickness which decreases cell heat losses. • The frozen sidewall ledge material is cryolite, containing almost no excess AlF3%. Therefore, the melting and freezing of almost pure cryolite ledge material has an impact on variations in the chemical composition of the total AlF3 concentration in the electrolyte. The composition of the cryolite electrolyte provides about 0.150 g/cm2 difference in density between molten electrolyte and liquid aluminum, thus ensuring a physical separation between the two liquids which reduces the reoxidation of the aluminum by the CO2 gas formed on the anode surfaces.

2.5.2.3. Dissolution of Alumina in Cryolitic Bath The relatively high maximum solubility of alumina in cryolitic bath, 13 wt.% at 1000
C, results from the favorable stereochemistry of forming different stable oxyfluoride aluminate complex anions, Al2 OF6 2 and Al2 O2 F4 2 , with octahedral or tetrahedral coordination of the large O2 and F anions about the small Al3þ cations [2,7,8]. These solute polyhedrals are stable because of the nearly identical ionic diameters for oxygen and fluoride anions. Cryolite dissolves the alumina by a chemical reaction and, while numerous

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0.5

0.4

Al2OF62-

Anion molar fraction

[Al2OF6]20.3 Al2O2F420.2 [Al2O2F4]20.1

Al2O2F64Al2OF106-

0.0 0

2

4

6

8

10

12

Mole ratio NaF/AlF3

Figure 2.5.6 Dissolved oxyfluoroaluminates anions versus AlF3 concentration (or NaF/AlF3 molar ratio).

species have been proposed, the resulting complex anion is generally considered to be an aluminum fluoroaluminate (Al2 OF6 2 ). Sterten [9] established that the resulting complex anion depends on the electrolyte chemistry as shown in Figure 2.5.6. As alumina dissolves in cryolite at low alumina concentrations and high AlF3 concentrations, by forming oxyfluoride aluminate ions, Al2 OF6 2 with a 2:1 ratio of aluminum to oxygen by the equation: Al2 O3 ðsÞ þ 4AlF6 3 ! 3Al2 OF6 2 þ 6F

ð2:5:7Þ

At higher concentrations, alumina dissolves in cryolite by forming oxyfluoride aluminate ions Al2 O2 F4 2 with a 1:1 ratio of aluminum to oxygen by the equation: 3 Al2 O3 ðsÞ þ AlF6 3 ! Al2 O2 F4 2 2

ð2:5:8Þ

Cells are generally operated with 2–4 wt.% Al2O3 in the electrolyte. Saturation ranges between 6% at 960
C and 13% at 1000
C. Alumina solubility varies largely with the excess % AlF3 in the electrolyte and temperature. The cryolite electrolyte has the ability to dissolve alumina, but of course only up to a certain percentage at normal operating temperatures. Above this level, the excess alumina concentration added to bath will

Aluminum Production

be deposited undissolved onto the bottom of cells as sludge, disturbing the current flow and generating metal pad motion, which results in a decrease in current efficiency. Note that when the alumina concentration is below 10% and the temperature of the electrolyte decreases below 960
C cryolite will freeze out on the coldest part of the cell, mainly on the sidewalls. However, if the alumina concentration is greater than 10%, then a-alumina freezes out of solution and it is difficult to redissolve in the melt. 2.5.2.3.1 Alumina Dissolution Mechanism Alumina additions to the cell electrolyte are called “feeding” and are typically performed by adding SGA to the top surface of the molten cryolite-based bath in the electrolysis cells, whereupon it is expected to dissolve and disperse rapidly in the bath. In actual practice with a few exceptions, SGA alumina is not used directly for aluminum electrolysis. Rather it is first sent to the plant’s dry scrubbing system where it is used as an adsorbent to capture gaseous and particulate fluorides from the cell exhaust gas. The reacted alumina, containing fluorides and moisture, is then transported to the electrolysis cells where it is added to the molten bath to produce aluminum. The adsorbed moisture and fluorides react and are released when added to the electrolyte, resulting in localized agitation and stirring of the molten bath and thus they assist the dissolution of alumina in cryolite. One of the difficulties experienced in adding SGA alumina or reacted alumina to cryolite melts in aluminum cells is that the alumina particles can form clumps or rafts that float on the top surface of the melt and then gradually sink, depending on the fines content, thus slowing the alumina dissolution process. The general solubility steps for the dissolution of alumina crystalline particles in cryolite melts in aluminum cells include: • Addition: About 1–2 kg of alumina particles is dumped onto the surface of the molten cryolite melt by point feeders with clumping or raft formation on the top surface of the bath. • Wetting: Alumina particles are wetted by the melt. • Heating: Alumina particles are heated from about 100 to 960
C. • Crust: Cryolite crust formation and subsequent melting occur around alumina particles. • Dissolution: Alumina particles dissolve into the boundary melt. • Distribution: Melt impregnated with dissolved alumina from the addition area is transferred to every area of electrolyte in the cell. Alumina dissolution rates in cryolitic electrolytes, as shown in Figure 2.5.7, have been the subject of intensive study for several decades. Alumina dissolution rate can be affected by the physical, chemical, morphological and microstructure properties of the ore, the dynamics of the feeding process, the chemistry of the electrolyte, and the superheat of electrolyte in the cells. Models have been developed by Haverkamp and Welch [10] for the dissolution of alumina powders in cryolite based on the rate transfer and diffusion. The shape of the

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6

Al2O3 (wt.%)

5 4 3 2

Alumina A

1

Alumina B

0

0

10

20

30

40

50

60

70

80

Time (min)

Figure 2.5.7 Typical dissolution rates for SGA alumina in molten bath.

curves generated gave a reasonable fit to experimental data on dissolution. It is evident that alumina “B” is initially slower than alumina “A” in the test, but they both eventually reach the same maximum concentration. The initial difference in alumina dissolution can increase the formation of deposits of undissolved alumina when feeding alumina very rapidly to aluminum cells that are operating at high amperage. Dando et al. [11] in a study subjected SGA alumina samples to thermal pretreatments prior to adding to molten electrolyte, in order to study the interplay between HF evolution, raft formation and dissolution, formulated several conclusions upon feeding alumina to the surface of molten electrolyte: • Moisture is released by the alumina in several stages: a rapid short-lived initial release from gibbsite in the SGA, followed by a rapid, broader rise in evolved moisture from alpha mineralization of the undissolved SGA, again followed by a sustained baseline evolution of moisture due to dissolution of the alumina. • SGA is rapidly mineralized to its alpha phase by the moisture in the alumina combined with vapor-phase atmolite (NaAlF4). • Cohesive rafts are formed by interparticle platelets. The data presented in this report suggest that the initial dispersion of alumina during feeding is one of the most important factors for preventing raft formation. Any clumping of SGA immediately following feed shots would promote raft formation, owing to the rapid mineralization, wetting, and interparticle platelet growth shown above. Raft formation promotes slow dissolution of the SGA by promoting formation of alpha phase rafts that subsequently sink to the bottom of the cell, promoting so-called sludge or muck formation. Furthermore, reductions in alumina dissolution rates will directly impact dissolved alumina availability and replenishment in the molten bath within the ACD of cells. Any delays in the rate of dissolved alumina replenishment will increase both anode effect frequency and duration. Given that aluminum smelting cells are operated at low dissolved alumina

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concentrations, 2–4%, and relatively low superheats (less than 10
C), the dissolution characteristics of the SGA can significantly impact the operating performance of the cells.

2.5.2.4. The Industrial Production of Aluminum 2.5.2.4.1 Faraday’s Law and Aluminum Production According to Faraday’s law, one Faraday (26.80 A h) of electricity will theoretically deposit one gram equivalent (8.99 g) of aluminum. The theoretical quantity of aluminum produced per kiloampere (kA) is calculated by the equation,   s molecular mass Al kA day Aluminum production ðkg=kAÞ ¼ ð2:5:9Þ ð#electrons  FaradaysÞ ð26:981586,4001Þ ¼ 8:0534 ¼ 3  96,485 A loss in aluminum production results mainly from recombination of anodic and cathodic products in the electrolyte. Also, some of the current is being consumed by parallel reactions that do not give net aluminum formation. The cell current efficiency is determined by the dividing actual Al production per unit time by the theoretical Al production for the same time period. For example, the actual aluminum production per day for one cell operating at 350 kA and 95% current efficiency is calculated: Aluminum ðkg=dayÞ ¼ ð8:0534kg=kA350kA0:95Þ ¼ 2678

ð2:5:10Þ

Because the aluminum production depends on the magnitude of the electrical current, there has been a steady increase in amperage and size of industrial aluminum cells from 40 kA cells (7 m length) in the 1940s to 200–250 kA cells (9 m length) in the 1980s and 350–400 kA cells (20 m length) in the 2000s. The aluminum production per unit time primarily depends on three factors: the potline amperage, the current efficiency of the electrolysis process, and the number of operating cells. In this manner, the annual production capacity for a potline of 300 prebake cells operating at 350 kA and 95% current efficiency is 293,214 mt/year, as indicated in Figure 2.5.8. 2.5.2.4.2 Current Efficiency—Aluminum Back Reaction The amount of aluminum predicted by Faraday’s law is never obtained in practice. In parallel with the aluminum production reaction, a quantity of metal made at the cathode dissolves into the electrolyte at the boundary layer at the bath–metal interface and is transported to the reaction zone where it is oxidized, recognized in the industry as the back reaction, by CO2 gas forming CO gas and Al2O3 as products, as shown in Figure 2.5.9. Industrial aluminum cells prior to the 1970s operated from 85% to 88% current efficiency, while modern cells now can operate 95–96% current efficiency.

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Figure 2.5.8 Annual aluminum production versus amperage for a line of 300 cells.

The corresponding concentration of CO produced by the back reaction in these cells is normally about 10%. 2AlðdissolvedÞ þ 3CO2 ðgÞ ¼ Al2 O3 ðdissolvedÞ þ 3COðgÞ

ð2:5:11Þ

The electrolytic production of aluminum from Al2O3 and the back reaction of aluminum to form CO gas and Al2O3 are expressed in the following equations: Al2 O3 ðdissolvedÞ þ 1:5CðsÞ ¼ 2AlðlÞ þ 1:5CO2 ðgÞ

ð2:5:12Þ

2ð1  xÞAlðlÞ þ 3ð1  xÞCO2 ðgÞ ¼ ð1  xÞAl2 O3 ðdissolvedÞ þ 3ð1  xÞCOðgÞ ð2:5:13Þ Combining these two reactions gives the overall aluminum cell reaction:     1 3 3 1 3 1 Al2 O3 ðdissolvedÞ þ xCðsÞ ¼ AlðlÞ þ 2  CO2 ðgÞ þ  1 COðgÞ 2 4 4 x 2 x ð2:5:14Þ Dissolved metal must diffuse away from the aluminum interface before it can be oxidized. The loss in current efficiency due to the back reaction is directly related to the dissolution of the aluminum product into the melt at the boundary, transport to the reaction zone, and ultimate oxidation of the metal with dissolved CO2 near the anode surface,

Aluminum Production

Figure 2.5.9 Diffusion and transport of aluminum in the electrolyte up to the reaction zone in an aluminum cell.

as shown in Figure 2.5.9. In order to minimize the dissolution of aluminum in the electrolyte, cells are purposely operated with a high excess % AlF3 and low dissolved alumina concentration in the melt; low electrolyte temperature, a balanced current distribution of all anodes, and a sufficiently wide interelectrode distance. The industry current efficiency of aluminum electrolysis for new cell technology has improved from about 82% in 1900 to 96% in 2000, as shown in Figure 2.5.10, while the current efficiency in the older cell technologies has lagged behind by about 5%. The abrupt increase in current efficiency in 1960–1965 was due to the advent of modern cell technologies and improved cell operating practices. It is interesting to note that current efficiency has not improved in aluminum cells over the past 10 years or so. Perhaps, we have reached the maximum obtainable current efficiency with the current industrial Hall–He´roult prebake cell design and technology? The maximum current efficiency for Søderberg cells is only 92% due to its poorer quality anode carbon, inferior magnetic design, and large gas bubbles that increases stirring and mixing at the bath–metal interface. The aluminum production per day for one aluminum electrolysis cell operating at 350 kA and 95% current efficiency is 2678 kg/day, which represents a 5%, or 121 kg/ day, loss in aluminum per day. Aluminum production=cell ¼ ð8:0534kg=kA350kA0:95Þ ¼ 2678kg=day ð2:5:15Þ

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100

Current efficiency (%)

95 90 85 80 75 70 00

19

20

40

19

19

60

19

80

19

00

20

10

20

Figure 2.5.10 Primary aluminum current efficiency from 1900 to 2010.

The aluminum production per year for a modern potline of 300 cells operating at 350 kA and 95% current efficiency is: ð8:0534kg=kA350kA0:95300cells365days=yearÞ ð1000kg=tÞ ¼ 293,214mt=year

All production=potline ¼

ð2:5:16Þ Specific energy consumption ¼

ð4:2V350kA24hÞ ¼ 13:00kWh=kg Al ð2678 kgAl=dayÞ

ð2:5:17Þ

The production of 1 ton of aluminum typically requires 420 kg of carbon, 1920 kg Al2O3, and 16 kg AlF3, and the specific energy consumption is 13.20 kWh/kg Al, as shown in Figure 2.5.11. The theoretical carbon consumption is only 333 kg C/t Al. The excess carbon consumption is mainly due to air oxidation of the hot anodes in the hooded cells. The reaction occurs preferentially in the pitch binder matrix and leads to the physical loss of coke particles to generate dust that floats on top of the electrolyte. Thus, anodes are carefully covered with a mixture of crushed bath and alumina to prevent excessive oxidation. The unused top part of anodes, “butts,” are removed from the cells when new anodes are set, crushed, and recycled in the carbon plant to make new anode carbon. Furthermore, the theoretical alumina consumption is only 1.89 kg/t Al. The excess alumina usage is due to the impurities in the metallurgical grade alumina including water and a small amount of alumina dust may be lost during transportation and transferring to/from silos.

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Aluminum Production

4000 kg Bauxite

1920 kg Al2O3

Aluminum electrolysis cell

420 kg Carbon

13.2 kWh/kg

1.42 kg CO2 + CO

1000 kg Al

Heat 6.7–9.7 kWh

Figure 2.5.11 Typical materials and energy consumption for the electrolytic production of 1 t aluminum.

2.5.2.4.3 Industrial Cell Design 2.5.2.4.3.1 Magnetohydrodynamics

Electrolysis cells in aluminum plants are arranged in long rows, called potlines where the cells are connected in an electrical series circuit with the transformer and rectifier systems located at one end of the potline. Such arrangements allow carrying out most cell operations by using overhead multipurpose cranes. A typical aluminum smelter consists of one to three potlines. Modern 350-kA prebake cells are positioned in a side-by-side arrangement in potlines, as shown in Figure 2.5.12. In most designs, the cells are situated above ground with solid aluminum bus bars located at ground level below them with current flow from one cell to another. The smelting process requires large amounts of electricity. A reliable and uninterrupted electrical power supply is a critical issue for aluminum smelters. Alternating or AC current supplied from the grid must be transformed into direct or DC current, which requires the use of large rectifiers, transformers, and sophisticated monitoring systems located adjacent to the potline building. The most inherent risk in aluminum production is a loss of electrical power. A failure of electricity supply lasting more than 2–3 h can cause the electrolyte in the cells to cool to the point where its electrical resistance is too great when power is restored resulting in shutdown of all cells. It is expensive and time consuming; usually it will take several months, to restart a frozen potline because the solidified aluminum and electrolyte must be physically broken out of the cells. Thus, significant business losses will be incurred due to the interruption in the event of a potline freezing. The remaining lifetime of the cathodes will be shortened due to the extra thermal stresses inevitable with shutdown and restart of cells.

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Figure 2.5.12 Potline of 350-kA prebake aluminum cells.

Magnetohydrodynamics (MHD) forces are generated in the aluminum metal caused by the interaction between the magnetic fields produced by the passage of high-amperage DC electrical current through nearby conductor bars and the electrical current flow in the aluminum metal in electrolysis cells. The large electric currents, 300–500 kA, in modern prebake cells pass through the electrically conducting liquids (electrolyte and metal) and generate powerful magnetic fields in the aluminum pool. These strong MHD forces increase metal velocity, metal wave height and frequency, and distortion of the aluminum–bath interface and thus require higher voltage and operate at lower current efficiency. These cells typically maintain a 25-cm deep pool of liquid aluminum in cells to minimize the impact of the MHD forces. A major achievement in the aluminum industry has been the development of advanced one-quarter size (3D) computer mathematical models of the interaction of MHD fields, forces, and resulting metal pool behavior in aluminum cells. This is exemplified by the model derived by Severo et al. [12] that is used for cell design and retrofit studies, as shown in Figure 2.5.13. These programs are used to develop new magnetic compensation in cells by rearranging the bus bar conductor systems to reduce the magnitude of the magnetic fields inside the cells and thus allow significant increases in the potline amperage without losing aluminum productivity or requiring extra voltage. 2.5.2.4.3.2 ACD and Cell Voltage

Electrolysis occurs in the electrolysis cell ACD or between the bottom horizontal surface of the anodes and the upper aluminum metal surface. The upper part of the aluminum cell shown in Figure 2.5.14 is referred to as the anode superstructure. The anodes, carbon blocks, are attached to long rods that are suspended and clamped to the anode beam located

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Aluminum Production 4.0 3.5 2.5

Bz

1.5 5

4

0.5

4

2 3

-0.5

2

0

-1.5

0

1

-2

-2.5

1

2

1

-4

5

-

1

4

0

3

-3.5 [mt]

Magnetic field component BZ, middle of the metal pad [N/m3] 5

15

25

35

45

55

65

75

85

95

105 115 125 135

1 0 -1 -5

-4

-3

-2

-1

0

1

2

3

4

5

Horizontal electromagnetic force field, middle of the metal pad

Figure 2.5.13 Magnetic modeling results for Bz magnetic field and horizontal forces in the aluminum metal pad.

along the length of both sides of the anode superstructure. Electrical current enters the cell through the large aluminum conductor bus connected to the anode beam. The anode beam moves all the anodes up and down simultaneously, thus changing the ACD, typically 3.8–5.0 cm, from the bottom surface of the anodes to the top surface of the aluminum pool. The ACD is not measured, but the average value can be calculated. Changing the ACD changes the cell voltage and energy input into the cell; most of the cell heat is generated in the electrolyte in this zone. Steel current collector bars attached to the bottom cathode lining carry the electrical current from the cell to the next cell in the series. 2.5.2.4.3.3 Alumina Feeding

The electrolytic process consumes alumina at a nearly constant rate. Alumina is fed to the aluminum cells at specific intervals at a rate equal to 1.92 times the aluminum production

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Air cylinder

Alumina

Anode current supply

Anode rods Crust Removable breaker doors

Carbon anode Electrolyte Molten aluminum Frozen ledge Cathode bars Insulation

Carbon lining Steel shell

Figure 2.5.14 Prebake aluminum electrolysis cell.

rate by point feeders located inside the anode superstructure. The computer process control system activates point feeders (crust breaker and alumina dosing devices) according to sophisticated algorithms to keep the alumina content in the electrolyte near the target value. The alumina concentration dissolved in the electrolyte cannot directly be measured; thus, the amount of alumina added to cells is regulated indirectly using an underfeed/overfeed computer control algorithm based on changes in the cell voltage due to the anode overvoltage at low and high alumina concentrations in the electrolyte. To add alumina to the bath, fast action pneumatic air cylinders activate crust breakers to open holes in the crust layer on top of the molten bath at two or more positions along the center line of the cell. Next, the point feeder adds 1.5–2 kg of alumina from a volumetric dispenser directly to 18–20 cm of molten bath where it dissolves. There are usually from two to six point feeders per cell depending on the aluminum metal production rate, which is determined by the potline amperage, as shown in Figure 2.5.15. The advantage of point feeders is that small quantities of alumina are added to the electrolyte at each break-and-feed. This method minimizes the risk of sludge formation in the center area of the cell, and there are minimal emissions of dust and fluorides during the alumina feeding operation.

2.5.2.4.3.4 Changing Anodes

Prebake anodes are manufactured by compressing a mixture of petroleum coke aggregates and coal tar pitch binder into blocks typically 40–60 cm wide by 120–150 cm long and 50–60 cm high. Petroleum coke is used because of its high purity. These blocks are baked in anode-baking furnaces above 1100–1200
C to convert the binder pitch into dense carbon. The carbon blocks are attached to an iron anode assembly by pouring

865

Number of point feeders per cell

Aluminum Production

- Prebake cells - Søderberg cells

8 7 6 5 4 3 2 1 0

0

100

200

300 Current (kA)

400

500

600

Figure 2.5.15 Number of point feeders versus potline amperage.

molten cast iron into two to six holes in the anode block for fluted steel stub connections. An aluminum, or in some instances copper, rod is attached to the top of the anode assembly. The top section of the anode rods are clamped to the anode beam and thus carry the electrical current from the anode beam to the molten cryolite melt in the cells. Many smelters now manufacture or cut “slots” in the bottom surface of anodes to divert gas bubbles, thus reducing the electrical resistance in the electrolyte and reducing the total cell voltage of operating cells. From 18 to 40 prebaked carbon anodes are required per cell depending on the potline amperage to maintain the anode carbon current density in the range 0.7–1.2 A/cm2. The carbon anodes are consumed at a rate of about 1.9 cm/day, depending on current, by the electrolysis due to the reaction to produce CO2. One or two anodes are replaced each 24–48 h in the cell with new anodes after 21–28 days depending on the anode dimensions and amperage, i.e., the anodic current density. A crushed bath–alumina mixture is used to cover the top and sides of the new carbon anodes to avoid excessive oxidation and also to form a crust on top of the electrolyte to reduce heat losses and emission of fluorides from the electrolyte. Søderberg cells are another type of aluminum electrolysis technology with only one large anode per cell that is continuously produced by a self-baking process by the cell process heat. However, Søderberg cells are being phased-out in the aluminum industry now due to their lower production efficiency, inherent environmental problems, and the emission of polycyclic aromatic hydrocarbons (PAHs) from the large self-baking anodes. 2.5.2.4.3.5 Metal Tapping

Positive aluminum-containing ions deposit continuously at the negatively charged cathode, the top surface of the aluminum pool, where they convert to liquid metal

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and thereby slowly raising the metal volume and level. A portion of the aluminum metal pool is removed every 24–48 h by siphoning the metal into a large crucible. The aluminum metal is weighed and sent to the cast house for further processing. The height of the metal level in the cells is maintained at a target value as the metal height impacts the heat loss out of the cell sidewalls because molten aluminum metal has a higher heat loss out the sides than molten cryolite. The metal depth also dampens the impact of the magnetic forces on the metal pool. 2.5.2.4.3.6 Cathode Lining

The lower cathode components consist of a rectangular reinforced steel box lined on the inside with carbon, refractory bricks, and insulating materials. The cell is thermally insulated internally on the bottom and lower sides to reduce heat losses and maintain the optimum cell heat balance. As aluminum plants continue to increase potline amperage to produce more metal, it becomes necessary to increase the heat transfer out of the sidewall of cathodes in order to maintain the frozen cryolite ledge to protect the sidewall lining material. Cell cathodes are now being constructed using higher thermal conductivity silicon carbide sidewall blocks, steel fins attached along the outside sidewalls, and air pipes that blow compressed air along the sidewalls. The cell cathode linings have lifetimes generally from 1700 to 3000 days. Higher cathode life requires excellent construction and good quality materials.

2.5.2.5. Electrode Reactions for Aluminum Electrodes 2.5.2.5.1 Anode Reactions 2.5.2.5.1.1 Electrochemical Production of Carbon Dioxide

In the electrochemical reaction to produce aluminum, the carbon anode is continuously consumed by the subsequent reaction with oxygen. There are two possible anode gas products, CO2 and CO, formed in the reactions: 1 3 3 Al2 O3 ðdissolvedÞ þ CðsÞ ¼ AlðlÞ þ COðgÞ 2 2 2 1 3 3 Al2 O3 ðdissolvedÞ þ CðsÞ ¼ AlðlÞ þ CO2 ðgÞ 2 4 4

ð2:5:18Þ ð2:5:19Þ

The reversible voltage potential at 1000
C to produce carbon monoxide at the carbon electrode surface in reaction (2.5.18) is 1.065 V compared with 1.181 V to produce carbon dioxide in reaction (2.5.19). Thus, the production of CO in reaction (2.5.18) is thermodynamically favored. However, in practice, the reaction kinetics and polarization result in the generation of CO2 gas at the anode electrode surface. At a low anodic current density, oxygen forms a stable C–O surface compound, and at a low gas evolution it can detach slowly and produce CO gas, as shown in Step “a” of Figure 2.5.16. But at the normal anodic current density used in commercial aluminum

Aluminum Production

Figure 2.5.16 Formation of CO2 at the carbon electrode surface.

cells, 0.70–1.2 A/cm2 and high alumina concentration, additional oxygen atoms react with the CO surface compound before it can detach forming an unstable CO2 stable surface compound that detaches rapidly producing CO2 gas, as shown in Step “b” of Figure 2.5.16. Desorption of CO from the carbon surface must be kinetically hindered. The formation of carbon dioxide at high current density decreases the effective surface coverage favorable for the formation of carbon monoxide, and therefore, at the high current densities the amount of carbon monoxide formed decreases. Welch and Richards [13] measured the anodic overpotentials for the discharge of oxygen-containing anions on a carbon anode in molten cryolite–alumina mixtures, and the results are shown in Figure 2.5.17. It was determined that the rate-determining step is the two-electron transfer reaction in which oxygen-containing anions are discharged. The steepness of the polarization curves demonstrates how quickly it changes in modern prebake cells that operate at low alumina concentrations and higher current density due to increases in potline amperage. The polarization inflection for a change in electrode process at lower current intensity below 0.05 A/cm2 for all alumina concentrations and carbon types indicates the formation/desorption of CO gas at the electrode surface. A hysteresis before and after electrolysis indicates that there is a slowly desorbing intermediate on the surface which is consistent with the CO formation mechanism. Welch and Richards [14] found that the anodic overvoltage follows the classical Tafel equation and concluded that the overvoltage was caused by slow transport of oxygen-carrying ions through the double layer. Welch indicated that anode carbon used for electrodes in aluminum electrolysis cells is disordered as the oxygen bonded to carbon atoms is in a strained five-membered ring, a poorly bonded carbon which is expected to have a higher rate of oxygen desorption from the anode surface. In practice, this has been proven out as more CO is evolved at low current density for poorly baked anodes which have a higher reactivity to carbon dioxide.

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Figure 2.5.17 Anodic overpotential versus current density and alumina concentration.

2.5.2.5.1.2 Parallel Reactions at the Anode

2.5.2.5.1.2.1 Electrochemical Production of Carbon Monoxide Hume et al. [15] have shown that carbon monoxide is electrochemically produced and coevolved at areas on the carbon anode electrode at localized areas of low current density as illustrated by the diagram shown in Figure 2.5.18. Because of the simultaneous rapid acceleration in both the consumption rate and the rate of carbon monoxide formation, it is unlikely to be solely due to the Boudouard reaction. During this process, the current can be carried by two reactions, and this is supported by the anode polarization gradient. Thus, direct electrochemical formation of carbon monoxide becomes an important mechanism at low current densities. These conditions exist on the sides of anodes, which contribute up to one-third of the electroactive area during electrolysis, and depending on the depth of immersion of the electrodes, they can carry more than 10% of the total anode current. At much lower current density, Zoric et al. [16] have illustrated that the current intensity decreases rapidly from 0.70 to 0.2 A/cm2 on the bottom corners of anodes and decreases from 0.2 to 0.08 A/cm2 on the sides of anodes. 2C þ 2O2 ! 2CO þ 4e

ð2:5:20Þ

The electrochemical formation of carbon monoxide by reaction involving the carbon electrode is one cause of excess carbon consumption. As seen from Equation (2.5.20), the electrochemical formation of carbon monoxide doubles the rate of carbon consumption.

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Figure 2.5.18 Carbon monoxide production at different anodic current densities.

2.5.2.5.1.2.2 Production of Carbon Monoxide by the Boudouard Reaction The endothermic reaction of carbon dioxide with carbon and the subsequent transformation into carbon monoxide is favored at higher temperatures and also doubles the carbon consumption rate according to the equation: CO2 ðgÞ þ CðsÞ ¼ 2COðgÞ

ð2:5:21Þ

The Boudouard, or the carboxy reaction, is temperature dependent with the thermodynamically favored product being all CO at temperatures greater than 900
C. However, the Boudouard reaction is not a major function in prebake aluminum cells as it appears that the polarized carbon anodes immersed in cryolite in aluminum electrolysis cells are protected from attack by CO2. It is possible that CO2 generated at the anode interface by electrochemical can attack the carbon anode, especially on the submerged sides of anodes that have a lower current density, as shown in Figure 2.5.18. The subsurface Boudouard reaction does occur to a greater extent within the anode pores in Søderberg cell anodes. These anodes have a significant higher permeability that allows CO2 to penetrate deeper into the pores of the anode. 2.5.2.5.1.2.3 Electrochemical Production of Perfluorocarbons 2.5.2.5.1.2.3.1 Anode Effects An anode effect is phenomena observed in many processes

involving the electrolysis of molten salts and is not always well understood. It is a condition produced by polarization of the anode in the electrolysis of fused salts and characterized by a sudden increase in voltage and a corresponding decrease in amperage.

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Alton T. Tabereaux and Ray D. Peterson

The onset of the anode effect in Hall–He´roult aluminum cells is primarily due to the depletion of the oxygen-containing ionic species at the surface of carbon anodes causing an increase in anode polarization. Prior to the anode effect, the alumina concentration in typical modern cells decreases about 30% of its normal value. Thus, oxygen-containing ions will be arriving at the electrode surface at approximately two-thirds of the normal rate. Once a resistive layer is formed on the anode surface, it causes an increase in voltage at a constant current or a decrease in current at a constant voltage. Once an anode effect occurs in operating cells, its anodic current density has exceeded its critical current density (ic). The cell critical current density is primarily a function of the concentration of dissolved alumina, anode dimensions, and amperage. However, it is also influenced by anode immersion in electrolyte, electrolyte flow, gas bubbles, temperature, and anode spacing. The critical current density calculated by Thonstad and Richards for industrial cells is calculated by the following equation [17]. ic ¼ 1:88wt:% Al2 O3 þ 1:54

ð2:5:22Þ

The decrease in alumina concentration also causes an initial deterioration of the wetting and increased gas bubble coverage, causing the current density at the active parts of the anode to increase leading to the risk of local alumina depletion, which results in an eventual anode effect. Fluorocompounds start being codischarged at the carbon anode surface forming carbon–fluoride intermediate compounds analogous to polytetrafluoroethylene, – {F2C–CF2}–, on the anode surface. These have a strong dewetting effect causing a rapid rise in voltage and increase the start of fluorides being oxidized and forming intermediate compounds at the anode surface with the discharge of perfluorocarbon gases, carbon tetrafluoromethane (CF4), and carbon hexafluoroethane (C2F6). Multiple anode cells are complicated in that each anode in the cell has a different anode current density due to difference in their individual ACD, causing anodes to have an unequal current distribution. Anodes also have different alumina concentrations in the anode–cathode spacing due to differences in their location with respect to the point source of alumina feeding and also differences in bath flow patterns in cells. The anode current distribution data shown in Figure 2.5.19 demonstrate that the anode effect may often start on only one of the prebake anodes prior to the onset of the anode effect, as indicated by the decrease in current on two anodes about 1–2 min prior to the actual anode effect. Once blockage of the electrical current by a film of CF4 gas occurs under one or more of the anodes, it significantly increases the current density on the remaining anodes resulting in an escalating anode effect sequence on all the prebake anodes. The anode effect is described as a blockage effect which inhibits the normal current flow between the anode and the electrolyte. A high percentage of current flow shifts from the bottom surface of anodes to the side surface,

871

Aluminum Production 35000

30000

Anode effect

25000

Pair Amps

20000

15000

10000

10:10:33

10:12:22

10:14:11

10:16:00

10:17:49

10:19:38

10:21:27

10:23:16

10:25:05

10:26:54

Figure 2.5.19 Anode current distribution in a prebake cell before and during an anode effect.

causing large fluctuations in current flow between anodes which induces instability in the aluminum metal pool. It takes from 5 to 10 min for the current distribution to become almost equal again on all of the anodes. Aluminum electrolysis cells operate at constant current; thus, the formation of a highly electrical resistive carbon–fluoride intermediate film on the bottom surface of carbon anodes causes the cell voltage to increase very rapidly from about 4.2 to >30 V, causing the electrolyte temperature to increase to >1000
C, which is observed practically. This generates a lot of heat especially as localized arcing often occurs because of the film. Once the anode effect starts, it proceeds rapidly as indicated by the fast increase in cell voltage. 4Na3 AlF6 ðlÞ þ Al2 O3 ðdissolvedÞ þ 3CðsÞ ¼ 4AlðlÞ þ 6NaFðdissolvedÞ þ 3CF4 ðgÞ

ð2:5:23Þ

2.5.2.5.1.2.3.2 Cell Gas Composition During Anode Effects Electrochemical reactions are still

occurring during anode effects as this is the only way current can flow through the electrolyte. Tabereaux et al. [18] measured the change in anode gas composition during

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Figure 2.5.20 Product anode gas composition changes during anode effects.

anode effects, as shown in Figure 2.5.20. The gaseous mixture consists primarily of carbon monoxide, 60–70%, and carbon dioxide, 20–30%; the measured CF4 content from both prebake and Søderberg cells was from 16% to 20%, and the C2F6 generation was small, 0.0–0.05%. The increase in the carbon monoxide product is due to the higher voltages and changes in the electrode surface reactions. 2.5.2.5.1.2.3.3 Anode Effect Termination The cell remains under the influence of the anode

effect until alumina is added, and the cell current is interrupted by a strong electrical short of the cell current by lowering the anodes until they make electrical contact with the aluminum metal waves or, alternatively, by inserting wooden poles under some of the anodes. These actions cause the current density on the anodes to decrease below the critical current density, which allows adherent gas bubbles formed at the anode surface to collapse or become detached.

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Aluminum Production

2.5.2.5.1.2.3.4 Formation of COF2 The standard potential required for the decomposition

of cryolite and the formation of CF4 is 2.54 V and the potential for formation of C2F6 is 2.78 V during anode effects, which are higher than the 1.191 V potential necessary for the formation of CO2. The changeover from solely oxide ion discharge (forming CO or CO2) to codischarge of the fluoride intermediate, COF2, in reaction (2.5.24) occurs at a much lower anode potential of 1.86 V. 4Na3 AlF6 ðlÞ þ Al2 O3 ðdissolvedÞ þ 3CðsÞ ¼ 4AlðlÞ þ 6NaFðdissolvedÞ þ 3COF2 ðgÞ

ð2:5:24Þ

The electrode potential for COF2 reaction was detected in the late 1970s. However, its existence was not confirmed until Dorreen and coworkers [19,20] detected the intermediate compound using mass spectroscopy. It is shown in Figure 2.5.21 that the COF2 formation occurs during electrolysis about 5 min prior to the full onset of the anode effect. After the rapid rise in cell voltage, the COF2 immediate compound decomposes at 960
C to form CF4 during anode effects. The formation of COF2 and subsequent CF4 formation continues at a fixed rate during the anode effect and this agrees with the measurements of CF4 emissions by Tabereaux et al. [18] from aluminum cells at smelters. Welch [21] indicates that the sharp rise in voltage is due to the coevolution formation of the fluoro compound, COF2 intermediate, that passivates the anode surface (DG
¼ 45.8 kJ at 960
C). 1.0⫻10−8

Partial pressure (Torr)

Electrolysis

Anode effect

69 (CF4 /C2F6) 1.0⫻10−9

47

COF2

66 1.0⫻10−10

1.0⫻10−11

50

55

60

65

70

75

Time (min)

Figure 2.5.21 Formation of COF2 and CF4 prior to anode effect.

80

85

90

874

Alton T. Tabereaux and Ray D. Peterson

2COF2 ðgÞ þ C ! CF4 ðgÞ þ 2COðgÞ

ð2:5:25Þ

Mechanistically, the formation of the COF2 product is likely to be similar to the formation of carbon dioxide, simply because of the high polarization, which is usually brought about by lowered rates of oxide arrival at the electrode surface through concentration polarization and fluoride ions codischarge. This can occur with either the surface edge or the bridge carbon atoms. 2.5.2.5.1.2.3.5 PFC Emission Rates A nearly universal definition of an anode effect has been established under the guidance of the International Aluminium Institute (IAI) to ensure that aluminum plants are using the same methodology when reporting PFC emissions. Anode effect and anode effect duration: An anode effect is typically considered to begin when the cell voltage exceeds a defined voltage threshold (6.0–8.0 V), and the anode effect is considered to end when the cell voltage drops below a second voltage threshold (6.0 V) and remains below this voltage level for a defined time (15 min). However, if the anode effect reoccurs within 15 min, as indicated by the cell voltage exceeding the threshold value, then it is considered as a repeat anode effect and is not counted as a new anode effect. The anode effect duration is the sum of the individual minutes when the cell is on anode effect. Anode effect duration refers to the total minutes per cell and day during which the voltage in the cells is above the threshold. The rate of PFC emission, kg CF4/min, has been established for different cell technologies based on actual PFC emission measurements from commercial cells in aluminum potlines for the duration over which cells have their anode effect as reported by Marks and Bayliss [22].

kg CF4 =tAl ¼ SðAEFÞðAEDÞ

ð2:5:26Þ

where AEF is the average anode effect frequency, AED is the average anode effect duration in minutes, and S is a technology-specific factor that is equivalent to the slope of the graph for (AEMAED) versus kg CF4/t Al per AEM, which is the emission rate in kg CF4/AE minute. Specific default emission rate values have been determined based on plant measurements by the IAI for five different categories of cell technology. These technologyspecific factors are used primarily as default values for aluminum smelters that have not measured their emissions rates. Aluminum smelters are encouraged to make their own PFC measurements in order to determine the specific slope value for their individual aluminum smelter. There are considerable variations (30%) in these factors. The cause for the wide variations in measured emission rates between plants that have the same cell technologies is not well understood. It is known that the low slope emission rate for the Søderberg cells is due to electric shorting between the aluminum metal waves, as indicated by large variations in cell voltage during anode effects, and the anode carbon during the anode effect

875

Aluminum Production

and the PFC emissions stop during the shorting. In addition, it is known that PFC emissions actually begin on one or more individual anodes prior to the cell voltage reaching the AE threshold value. The method used to terminate the anode effects also to some extent influences the PFC emission rates as differences in emissions have been measured in plants using slow anode up/down pumping versus plants using fast aggressive anode down moves to terminate the anode effect quickly. 2.5.2.5.2 Cathode Reactions Measurements indicate that the smaller and more ionic mobile Naþ ions carry most of the current through the electrolyte. However, the deposition of aluminum is favored over sodium as the reversible decomposition is favorable. Thus, aluminum is the primary product produced at the top surface of the cathodic aluminum pool in the electrolysis cells. The most probable cathodic process involves a charge transfer at the cathode surface in which aluminum-containing anions (hexafluoroaluminate) are discharged to produce aluminum metal as well as F ions to neutralize the charge of the current carrying Naþ ions [23]. The most probable overall cathodic reactions are: AlF4  þ 3e ! Al þ 4F þ

3Na þAlF6

3



þ

ð2:5:27Þ





þ 3e ! 3Na þ 3F þ AlðlÞ þ 3F

ð2:5:28Þ

Sodium is codeposited at the cathode along with aluminum, Naþ þ ðAlÞ ) NaðAlÞ þ e

ð2:5:29Þ 3þ

Since the reaction involves an interfacial transfer of the Al ions, the rate of aluminum deposition must be matched by the rate of transfer of Al3þions from the solution. Polarization effects at the cathode contribute much less to overvoltage than at the anode. The ion complexes AlF6 3 and AlF4  have higher ionic mobility than their anodic counterparts, which lowers the concentration polarization effect. In addition, there are no gas bubbles at the cathode which influence both resistance and concentration polarization. 2.5.2.5.2.1 Cathode Parallel Reactions

2.5.2.5.2.1.1 Production of Sodium at the Cathode The primary cathodic reaction in cryolite-based melts is the reduction of Al3þ containing species. The next most favored cathodic reaction is the deposition of sodium which forms an alloy with aluminum. Sodium is produced at the bath–metal interface due to the chemical reaction occurring when reaching thermodynamic equilibrium and sodium partitioning into the two phases based upon equilibrium constants for the system. 3NaF þ Al ¼ AlF3 þ 3Na

ð2:5:30Þ

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Alton T. Tabereaux and Ray D. Peterson

During electrolysis, the content of sodium increases with increasing cathodic current density. This effect is related to the cathodic concentration overvoltage which is mass transfer controlled. It is caused by the fact that the sodium ion is the carrier of current, while aluminum is being deposited. This leads to the formation of a concentration gradient in the boundary layer at the cathode, with enrichment of NaF and depletion of AlF3 at the interface. The addition of lithium fluoride has been shown to result in a decrease in the cathodic overvoltage. The cathode overvoltage reflects the concentration gradient with respect to the concentration of NaF and AlF3 at the bath–metal interface. Hence, there should be a relationship between the overvoltage and the sodium content in the aluminum metal in operating cells. The magnitude of the overvoltage is dependent on the cell design and on the convection pattern in the cell. Modern aluminum cells with good magnetic compensation and low flow rates at the bath–metal interface have a high sodium content, >120 ppm in the aluminum cathode pool. Strong relationships exist between the sodium content in the aluminum pool in cell and the cell current efficiency. 2.5.2.5.2.1.2 Production of Other Metals at the Cathode Metal impurities, e.g., iron, silicon, manganese, copper, and vanadium, in the alumina and anode carbon dissolve in cryolite to form ions that are codeposited at the aluminum cathode surface. These metals are more noble than aluminum, i.e., they have reversible decomposition potentials that are more favorable for reduction than aluminum. Iron and silicon are specific concerns as they are the highest impurities present in alumina and coke and will eventually be contaminates in the aluminum pool at the cathode. Fe2 O3 þ 2Na3 AlF6 ! 2FeF3 þ Al2 O3 þ 6NaF FeF3 þ 3e ! FeðaluminumÞ þ 3F

ð2:5:31Þ ð2:5:32Þ

Impurities in the electrolyte with multiple valence states, such as phosphorus, can be reduced at the cathode and then reoxidized at the anode, thus consuming electrical current without producing any aluminum. 2.5.2.5.2.1.3 Production of Alkali Metal at the Cathode The alkali and alkaline earth metals (lithium, magnesium, and calcium) dissolve in cryolite to form ions (LiF, MgF2, and CaF2) but are not reduced at the cathode because they are less noble than aluminum, and thus, their concentration increases in the electrolyte. The concentration of alkali and alkaline earth metal fluorides is normally well below the saturation concentration in cryolite. However, it has been found that some reduction of these metals does take place as an equilibrium concentration of lithium, magnesium, and calcium in the aluminum metal pool in cells relative to the concentration of these metals present in the cryolite electrolyte, indicating that the process may be mass transfer controlled, or a chemical equilibrium at the bath–metal interface. Given enough time to reach

877

Aluminum Production

thermodynamic equilibrium, any metal fluoride will partition into the two phases based upon equilibrium constants for the system, for example, lithium: 3LiF þ Al ! 3LiðlÞ þ AlF3

ð2:5:33Þ

Peterson and Tabereaux [24] determined that the concentration of alkali and alkaline earth metal in aluminum relative to % alkali metal fluoride dissolved in the electrolyte and excess AlF3 concentration, when expressed as the CR ratio (% NaF/% AlF3) can generally be expressed as a second-order polynomial expression written as: LiðlÞ ¼ 8:94 þ 17:38ðCRÞ  7:01ðCRÞ2  2:26ð%LiFÞ  0:13ð%LiFÞ2 þ 6:4ðCRÞð%LiFÞ ð2:5:34Þ MgðlÞ ¼ 15:31 þ 14:62ð%MgF2 Þ  1:59ð%MgF2 Þ2 þ 42:09ðCRÞ  14:54ðCRÞ2 þ 12:62ð%MgF2 ÞðCRÞ

ð2:5:35Þ

CaðlÞ ¼ 6:71  0:04ð%CaF2 Þ2  14:2ðCRÞ þ 6:55ðCRÞ2 þ 0:69ð%CaF2 ÞðCRÞ

ð2:5:36Þ

2.5.2.6. Thermodynamics for Aluminum Electrolysis 2.5.2.6.1 Standard-State Gibbs Free Energy The driving force for the heterogeneous electrochemical reaction for the production of aluminum is the standard Gibbs energy gradient at the reaction interfaces, and it is calculated by the change in the standard Gibbs energy at equilibrium, DG
, for the reaction to form the products, Al and CO2, from the reactants, Al2O3 and C, at 1250 TK and where x is the fraction current efficiency according to the equation:     3 1 3 1 1


DGreaction ¼ 2  DGCO2 þ  1 DGCO  DGAl2 O3 4 x 2 x 2


ð2:5:37Þ

Substituting the individual numerical values of the standard Gibbs energy for the formation of compounds from the elements into the above equations gives the following equation for the standard Gibbs energy as a function of current efficiency [2].   36:1 DGreac 1250 ¼ 378:4  kJmol1 x


ð2:5:38Þ

Accordingly, at 100% current efficiency, the standard Gibbs energy for the formation of aluminum from alumina is 342.3 kJ/mol, and for every 10% increase in the reoxidation of aluminum in industrial cells, there is a 4 kJ/mol decrease in the standard Gibbs energy.

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Alton T. Tabereaux and Ray D. Peterson

2.5.2.6.2 Reversible Decomposition Potential, Nernst Voltage The magnitude of the cell potential is a measure of the driving force behind a reaction. The larger the value of the cell potential, the further the reaction is from equilibrium and the sign of the cell potential indicates the direction in which the reaction must shift to reach equilibrium. Consider the electrochemical reaction between alumina and carbon to produce aluminum: Al2 O3 þ C ¼ Al þ CO2 E ∘ ¼ 1:20V

ð2:5:39Þ




The fact that E is negative indicates that the system does not favor the products of the reaction in the direction as written. The reaction will require about 1.20 V using a carbon electrode in order to proceed to form aluminum as a product. The Nernst equation describes the relationship between the cell potential at any moment in time and the standard-state cell potential. The standard cell potential of the reaction is directly proportional to the change in the standard Gibbs energy of the electrochemical reaction in the Nernst equation.  
DGreact
E ¼  ð2:5:40Þ nF


Substituting n ¼ 12, F ¼ 96,485 J/V equivalent and the numerical values for DG as a function of TK for an alumina-saturated electrolyte, according to Haupin [25], gives the following equation for the reversible potential for aluminum production.


E ¼ 1:896 þ 0:000572TK

ð2:5:41Þ

However, during normal cell operations, the bath is not saturated with alumina. Consequently, the Nernst equation is used to correct the standard potential for aluminum electrolysis for the actual activities for the reactants to give the reversible equilibrium potential ERev for aluminum electrolysis. At 960
C, the equilibrium potential for cells at alumina saturation is 1.191 V for saturated electrolyte, 8% Al2O3; however, the equilibrium potential is 1.222 V at 2.6% Al2O3 alumina. 2 3   a4Al r3CO2 RTK
5 ð2:5:42Þ  ln 4 ERev ¼ E ¼ 3 nF 2 aAl2 O3 aC     RTK 1 ERev ¼ 1:898  ln ð2:5:43Þ nF ðaAl2 O3 Þ where aA is the activity concentration of aluminum in the melt, aAl2 O3 is the activity of Al2O3 in the melt, and rCO2 is the partial pressure of CO2. Substituting the appropriate values gives the dependent cell potential at reversible, nonstandard conditions:

879

Aluminum Production

ERev ¼ 1:191Vðat 960
CÞ

ð2:5:44Þ

2.5.2.6.3 Change in Enthalpy The change in the standard enthalpy per mole, or per kg of aluminum for transference of the reactants at 298 TK to the reaction products at the electrolyte temperature TK for reaction (2.5.10) is 550 kJ/mol Al at 1000
C. But the actual energy requirement for aluminum production depends on the cell current efficiency [26]. Substituting the various energy expressions for the specified conditions, 970
C and 8 wt.% alumina, gives the theoretical minimum energy requirements as a function of current efficiency, xAl.      139:180 143 ∘ þ 477:240 kJ=molAl ¼ DHtot ¼ þ 4:91 kWh=kgAl ð2:5:45Þ xAl xAl For a typical 5% loss in current efficiency in aluminum cells, the overall enthalpy is 6.65 kWh/kg (or 478.5 kJ/mol Al); in addition, for each 5% change in current efficiency, the change in the standard enthalpy is 0.08 kWh/kg. The electrolyte in industrial aluminum cells is not saturated with alumina, thus its chemical activity is less than unity. This necessitates a correction term in the above calculations: RT ln aAl2 O3 , where R is the universal gas constant, T ¼ 1250 TK, and aAl2 O3 ¼ 0.3. The energy correction term amounts to about 0.2 kWh/kg Al (or 12 kJ/mol Al). A small fraction of CO2 reacts with carbon to make CO by the Boudouard reaction, but in general, the reaction is small and thus ignored in these calculations.

2.5.2.7. Energy Efficiency of Aluminum Cells The aluminum industry average energy consumption is presently about 14.0 kWh/kg Al compared with the theoretical minimum energy consumption of 6.42 kWh/kg Al for cells operating at 95% current efficiency. The actual energy efficiency for industrial aluminum cells is only about 50% due to the large heat losses to the surroundings. The specific energy consumption of aluminum electrolysis cells can be calculated from the total cell voltage and % current efficiency. Energy consumption ðkWh=kgÞ ¼

298:06Ecell %CE

ð2:5:46Þ

The aluminum industry has made major progress in reducing the energy consumption as modern industrial aluminum cells operate close 4.2 V and 95% current efficiency and 13.2 kWh/kg Al. The improvement is due largely to changes in cell designs that reduce the cell voltage components by increasing the size of the conductors and using materials that have a higher electrical conductivity. A few aluminum plants operate at voltage as low as 3.85 V, which reduces the cell energy consumption to 12.2 kWh/kg Al by

880

Alton T. Tabereaux and Ray D. Peterson

reducing the metal instability by improving the magnetic compensation in cells and then decreasing the anode–cathode spacing lower than normal. The thermoelectric design of an aluminum cell is the aspect of the cell which has the most influence on the cell power consumption and it is also a key element affecting the life of the cell lining. The thermal balance of the cell is often the limiting factor which prevents the smelter to increase production by increasing the line amperage. Currently, the heat losses and corresponding ledge freeze profiles of one-quarter of the entire cell can be modeled using 3D thermoelectrical models. The relationship between the voltages corresponding to the enthalpy and cell heat losses was first introduced by Haupin in the voltage–energy diagram in Figure 2.5.22 for aluminum production with carbon [25]. The total cell voltage, 4.20 V, is shown on the left scale split up into individual components and it is compared with the total corresponding energy consumption (14.25 kW), comprising the process energy (6.4 kW) and the cell heat losses (7.1 kW), as shown on the right scale. Additional energy (0.25 kW) is required to heat the anode cover material and anodes to operating temperature in cells. Operating an aluminum cell with noncarbon anodes results in a 2.8 kWh/kg Al higher energy consumption compared with carbon anodes. Tests to date indicate that inert anodes will not result in a lower, or higher, current efficiency than cells using carbon anodes.

Cell-specific energy 4.20 0.15 Vext

4.00

0.40 Vca 0.30 Van

Cell voltage

0.05 Vbub 3.00 1.55 Vel+Vbub (Ohmic)

1.00

Cell heat loss = 7.1

Inert oxygen evolution anodes -9.2 kWh/kg

10.0

1.50 Vel DH f(CO2) = 2.8 DH = -0.25 Preheat cover and anodes

2.00

1.80 EBemf (Non-ohmic)

15.0

14.25

0.02 hca 0. 54 hsa 0.04 hac

E°rev= 1.20

DH Total = -6.4 Reaction and enthalpy

-DH f(AI2O3) -6.4 kWh/kg

Carbon – CO2 evolution anodes

0

Figure 2.5.22 Haupin’s voltage–energy diagram distribution for aluminum production.

5.0

0.0

Specific energy consumption (kWh/kg Al)

Total Eext = 0.5 Elnt = 13.75

Heat loss

Cell voltage

Aluminum process

5.00

881

Aluminum Production

2.5.2.7.1 Cell Voltage Components

Because the DG for the aluminum cell production reactions is positive, the E values are negative. Likewise, by thermodynamic convention, cell voltage and all components of it are negative. However, engineers and cell operators in the primary aluminum industry consider the cell voltage to be positive; hence in this work, the sign of all cell voltages will be made positive. But because some of the voltage components are considered to be positive or negative depending on the scientific principle used, the absolute value is indicated by brackets. The overall cell voltage Vcell is the sum of individual cell ohmic voltages VIR and the sum of the cell and electrode nonohmic polarization voltages |Epol|, X X



Epol

VIR þ Vcell ¼ ð2:5:47Þ The ohmic voltage drops in aluminum cells result from the potential difference when passing electrical current through the cell conductors (anode electrode, electrolyte, gas bubbles in the electrolyte, cathode electrode, and the connecting bus between cells) that have a measurable resistance. X ð2:5:48Þ IR ¼ ðIRbub þ IRan þ IRel þ IRca þ IRext Þ The nonohmic polarization voltage components |Epol| of an aluminum electrolysis cell include the reversible equilibrium potential |Erev| and the polarization concentration and reaction overvoltages  at the electrode interfaces. Typically, voltage values for all the individual voltage components in an industrial aluminum cell are illustrated in Table 2.5.3. The individual electrochemical potentials and the voltage drops within the ACD of aluminum cells are shown in the expanded drawing shown in Figure 2.5.23. The voltage Table 2.5.3 Typical Voltage Components in an Aluminum Electrolysis Cell Cell Component

V

Van

Anode voltage drop

0.30

Erev

Reversible equilibrium voltage

1.20

aa

Anode reaction overvoltage

0.50

ca

Anode concentration overvoltage

0.02

Vbub

Gas bubble voltage drop

0.05

Vel

Electrolyte voltage drop

1.50

ca

Cathode overvoltage

0.08

Vca

Cathode voltage drop

0.40

Vext

External busbar voltage drop

0.15

Vcell

Total cell voltage

4.20

882

Alton T. Tabereaux and Ray D. Peterson

Figure 2.5.23 Electrochemical and electrical resistance components of voltage within the ACD.

drops due to the electrical resistance of the electrolyte and gas bubble occur in the electrolyte and are a function of the electrical current in the cell. The electrochemical potentials occur at the boundary gradients at the electrode interfaces and are a function of chemical concentrations and temperature, and current intensity has an influence on the electrode overpotentials. The largest ohmic component in the cell is the voltage drop through the electrolyte, which is linearly dependent on the ACD. Under normal operating conditions, the ACD is between 4 and 5 cm. The ohmic voltage drop in the electrolyte can be calculated by the following equation: Vel ¼ ia =kðACD  db Þ where ia is the amperage, К is the electrical conductivity (S/cm), ACD is the anode– cathode distance (cm), and db is the average bubble size (cm). The resistance of the bath voltage drop can be reduced by decreasing the interelectrode distance or by improving the electrical conductivity of the melt by changing the individual additives. The combined effect of temperature and AlF3 can be large. An equation to calculate bath conductivity (К, per ohm/cm) with concentrations in wt.% and T in Kelvin is given by the following equation [27]. ln k ¼ 1:977  0:013ð%AlF3 Þ  0:0200ð%Al2 O3 Þ  0:0060ð%CaF2 Þ  0:0106ð%MgF2 Þ  0:0019ð%KFÞ þ 0:0121ð%LiFÞ  ð1204:3=Tbath Þ

ð2:5:49Þ

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Aluminum Production

2.5.2.7.2 Electrode Polarization 2.5.2.7.2.1 Polarization Voltage

Theoretically at the Nernst electrode potential, the rates of the reactions in the forward and reverse directions, expressed in terms of rate equations, are counterbalanced and result in no formation of product. Therefore, before the reaction can actually occur, the potential gradient at each of the electrodes must exceed the equilibrium Nernst potential. This extra voltage, which can be different at each electrode, is referred to as the electrode polarization or overvoltage. To get the reaction to go faster by carrying a high current density, the magnitude of the electrode potential gradient has to be increased by providing the required amount of polarization or overvoltage. The amount needed depends on the specific electrode reaction and the existing conditions. There is a very thin stagnant layer of electrolyte (boundary layer) at the surface of each electrode. Convection does not function in this layer. Ions must diffuse through it. This creates concentration gradients. Gas bubbling makes the layer much thinner at the anode than at the cathode. The total anodic overpotential is composed of two components: reaction overpotential, a, attributable to the energy necessary to drive the reaction controlling step at the given rate or prevailing current density (i), and the concentration polarization, c, due to any inhibition of transfer of oxyanions (diffusion, convection, geometrical or temperature constraints), leading to a concentration gradient across the boundary layer. The anode reaction overvoltage can be large, 0.40–0.56 V, depending upon the nature of the carbon, orientation of the anode, and current density, while concentration polarization is in the order of millivolts except near anode effect. 2.5.2.7.2.2 Back EMF Nonohmic Voltage

When current is passed through the electrolysis cell, the overall voltage is a combination of the ohmic voltage drops due to the resistance of the ionic and electronically conducting materials plus the sum of the operating anode and cathode potentials. Immediately on interruption of the current during normal operation of the cell, the ohmic component of the voltage drops disappears but there is a residual voltage given by the sum of the electrode potentials at the moment this current is interrupted—this is commonly referred to as the back EMF. At the instance of current interruption, each electrode has reaction products absorbed on the surface as well as reaction intermediates that may be performed before release of the final product. The electrolytic cell is therefore able to acts like a battery and that is why it is referred to as the back EMF. It is the sum of the reversible decomposition potential and the anode and cathode overpotentials that exist for the conditions that are based on the preceding conditions. This is a nonohmic cell voltage, VBemf, even though its establishment may be current dependent given by the following equation: VBemf ¼ jErev j þ jj

ð2:5:50Þ

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where |Erev| is the reversible equilibrium potential and || is the sum of the cell polarization overvoltages. The overpotentials are generally current density dependent as well as dependent on temperature, alumina concentration, and structure of the anode carbon. Where some of the overpotential is due to concentration gradients between the bulk of electrolyte and that existing on the electrode surface, this will slowly self decay. However, the “battery” can hold a voltage or potential greater than the Nernst potential due to the presence of intermediate products from the overall electrolysis reaction. Typically, this voltage is between the Nernst potential of 1.2 and 1.6 V. Because the cells are interconnected in series, there can be a very high back voltage of the potline. The theoretical relationships for establishing the electrode polarizations that exist are usually expressed in terms of current density of the cells. These theoretical relationships are complex functions of the current, even though the value is dependent on process conditions rather than current. Hence, it is a nonohmic voltage. Historically, cells used to be operated at anode and cathode densities in the range 0.7–0.8 A/cm2, and in this range, the sum of the polarizations typically were approximately 0.45. Hence, if voltage measurements within that current band were linearly extrapolated to zero current, the intercept voltage would be approximately 1.65 V. It assumes a higher values at higher current densities used in modern cell operation. In order to smooth out cell voltages when the line current fluctuates (and fluctuations occur through rectification and process changes), the simplification has been made to normalize the voltage by calculating a pseudoresistance based on the assumption that the back EMF or combined combination of Nernst potential and electrode polarization is 1.65 V. In addition, there is a safety issue regarding the nonohmic voltage potential. Each cell continues to have an electrode potential even after the power is switched off. These cells behave as a charged battery, slowly decaying with time, because the anodes still have partially discharged chemical intermediates in a thin region on its surface. The total voltage potential can be quite large as it is the sum of the individual decaying voltages on each cell in the potline, for example: 300 cells  1.5 V (average) ¼ 450 V. Mohammed and coworkers [28] reported that a slower decrease in potline amperage reduces the cell potential in cells when the potline power outages are planned, and also extra precautions are necessary when working on cells during power outages in potlines. 2.5.2.7.3 Cell Pseudoresistance and Cell Voltage Control In aluminum smelters, the computer process systems automatically control the cell voltage in aluminum electrolysis cells by moving the anodes up or down and thus change the ACD, depending on the target value and control bands. The control system also makes automatic adjustments in the cell voltage based on the variation in the actual voltage values or noise level. By Ohm’s law, the resistance of a circuit is equal to the voltage

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divided by current. However, the voltage in an aluminum cell changes whenever the potline amperage changes, for example, due to anode effects or other events. Thus, the computer process control system calculates a pseudoresistance that is used to actually control cells that corrects for errors in cell voltage control of cells during periods of current fluctuations by subtracting the total nonohmic voltage from the actual voltage values, thus making the change in cell pseudoresistance directly proportional to any change in potline amperage. The target set point for the pseudoresistance of an aluminum cell operating at 200,000 A and 4.20 V is calculated in the following manner: voltage Cell resistance ¼ amperes     ðVcell  VBemf Þ ð4:2  1:65Þ ¼ ¼ 12:7mO Cell pseudoresistance ¼ amperage 200,000

ð2:5:51Þ ð2:5:52Þ

2.5.2.7.4 Gas Bubbles Initially, nearly spherical CO2 bubbles nucleate and grow on the bottom horizontal surface of the carbon anodes. Buoyancy forces acting on the bubbles cause the small bubble to coalesce with nearby bubbles, and as they eventually start moving toward the edge of the anode, these bubbles engulf smaller bubbles adhering to the anode surface, sweeping the surface clean. These macrobubbles assume the shape of aerofoil in a cross section with the thicker front and a reduced thickness of the trailing end. Through instability caused by bubbles coalescing this generates a flow and a forward movement of the gas as it rolls along the horizontal surface prior to release from the edges of anodes (Figure 2.5.24). The presence of gas bubbles under the anode contributes to the overall bath resistance in the cross section of electrolyte in that zone because the gas is nonconducting and therefore the cross-sectional area of the electrolyte has been reduced. Accordingly, models have been developed for calculating the resistance effect based on the average layer thickness and fractional surface coverage of the surface. An expression for the additional voltage drop caused by gas bubbles that is applied in numerous cell models has been developed by Hyde and Welch [29] which expresses the

Anode CO2 Cryolite

Figure 2.5.24 Typical profile of a large CO2 gas bubble moving across the bottom surface of an anode.

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voltage drop in terms of the anode current density, electrolyte conductivity, average bubble layer thickness and fraction, is given by the following equation   ia db Vbubble ¼  ð2:5:53Þ  db k ð1  fÞ where ia is the anodic current density (A/cm2), k is the conductivity (1/Ocm), db is the bubble thickness (cm), and f is the fraction of anode covered by bubbles. Furthermore, Haupin [25] verified that the analysis and equations by Hyde and Welch can also be used to calculate the average bubble layer thickness under the anodes by the following equation:   ð0:5517 þ ia Þ db ¼ ð2:5:54Þ ð1 þ 2:167ia Þ The average data reported for gas coverage under the anodes are in the range 40–80%, the thickness of the gas bubble layer is 0.5–0.7 cm and the bubble size is 100–120 cm. These relationships have been developed for electrodes that are predominantly horizontally orientated but the rate of release of the bubbles is also dependent on the path they have to travel and also the angle of the surface from the horizontal. With the advent of slotted anodes, these parameters have therefore changed and the bubble layer resistance has been reduced. 2.5.2.7.5 Aluminum Cell Overvoltages The kinetics for the production of reaction products requires a hierarchy electrode potential beyond that of the Nernst potential for maintaining the imposed current density (which is the measure of the reaction rate). This is often referred to as overvoltages or more correctly overpotentials. These potentials resulting from the electrode kinetics can be due to some or all of the following: • Concentration gradients between the bulk of the electrolyte and those existing at the interfacial surface for the participating species. • Surface intermediate reactions at the electrode where rearrangements take place to form the final product. These can include variations in surface coverage. This requires additional voltage at both electrode (overvoltages) surfaces to cause the reaction to proceed at the desired rate to produce products (i.e., current density). These overvoltages result from electrode kinetics due to concentration gradients and surface reactions at the electrodes and are affected by the electrolyte chemistry, current density, and temperature. The components of the cell overvoltage || are ac ¼ anodic concentration overvoltage, ar ¼ anodic reaction overvoltage, cc ¼ cathode concentration overvoltage, and

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cr ¼ cathode reaction overvoltage. The cathode reaction overvoltage is negligible in aluminum cells and thus it is omitted in further considerations. jj ¼ ar þ ac þ cr þ cc

ð2:5:55Þ

The higher the current, the higher the steady state surface coverage, and this will depend on the rate constant, which in turn is dependent on the carbon structure and stability. The build-up of the surface immediately will introduce polarization by that concentration change, so the electrode potential has to increase to maintain the prevailing rate as high surface coverage is approached. 2.5.2.7.6 Environmental Issues Within the potroom, a combination of work practices such as anode covering and the automated alumina feeding system can generate alumina dust, because of the small particle size of the alumina and make-up electrolyte materials used. The emission of gaseous and particulate fluorides is of prime concern in aluminum smelters. In order to minimize the environmental impact of large capacity aluminum smelters to meet more stringent environmental requirements, the aluminum industry has made significant efforts to reduce fluoride emission on a kilogram (kg) per metric ton (t) of aluminum production basis. Aluminum cells are completely enclosed with removable covers to collect all the cell fluoride emissions of particulates and gas fumes. Aluminum smelters have highly efficient alumina dry scrubbing equipment, which removes up to 99% of all emissions from the pots. Dry scrubbers use the raw material alumina as the sorbent for removal of gaseous and particulate fluorides from the cell exhaust gases. The fluorides are chemisorbed on the surface of the alumina particles, which are then called secondary or reacted alumina. This material is stored in large silos and is later used as feed material to the cells. This means recycling of the captured fluorides and reduces the overall fluoride consumption significantly. As a result, current average levels of fluoride emission to atmosphere are as low as 0.5–0.6 kg of fluoride/t of aluminum for modern large capacity prebake smelters compared with 1 kg/t for older prebake smelters. It is even possible to reduce smelter fluoride emissions to the level of 0.3–0.4 kg/t with the development of new, innovative dual duct suction technology that increases the cell hooding/duct suction velocity by a factor greater than three, 5000–15,000 Nm3/h when covers are removed during anode changing. Also, the emission from hot spent anode butts accounts for 35% of the total emission in potrooms. The majority of HF emissions occur during the first 20 min of cooling. The solution was to put the hot anodes inside cooling boxes that retain the fluorides. Another major environmental impact of refining and smelting is the emissions of carbon tetrafluoride (CF4) and carbon hexafluoride (C2F6) which are strong greenhouse gases. Fluorides such as CF4 and C2F6 are of concern because of their high global warming potential due to their long lifetime in the atmosphere. The worldwide aluminum industry voluntarily reduced its PFC emission by 75% from 1990 to 2009 despite a

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90% increase in the primary production of aluminum. Since 1990, the industry has reduced PFC emission from 96 million tons, carbon dioxide equivalent, to 22 million tons while increasing primary aluminum production from 19.5 million tons to 37 million tons. A 23% reduction in the greenhouse emission intensity was achieved by aluminum companies, from a total of 10.40 metric tons of CO2 equivalents/ton of aluminum in 2005 to 7.95 metric tons of CO2 equivalents/ton of aluminum in 2011. PAHs result from the manufacture of anodes containing coal tar pitch as the binder and is removed from the exhaust gases of the anode plant and anode-baking furnaces by scrubbing with alumina. Furthermore, aluminum plants have discontinued the use of hot ramming paste for cathode construction that contained coal tar pitch binder due to the development of low-temperature ramming paste that does not contain coal tar pitch. PAH is emitted into the potroom atmosphere from the anodes of Søderberg cells during the electrolytic process; the average level of measured emissions to air of total PAH is 0.06 g/t of aluminum. Modernized Søderberg cells use dry anode paste technology with lower pitch content, cooler anodes, and some aluminum smelters have doors on top of the anodes that reduce the PAH emissions. However, Søderberg cells are being discontinued in the aluminum industry due to their lower production efficiency and inherent environmental problems, particularly the emission of PAHs from the large self-baking anodes. Sulfur is a typical impurity in the anode coke with concentrations ranging from 1 to 4 wt.%. This is ultimately emitted with the exhaust gases as the carbon anodes are consumed at the electrolytic process. While it is initially released as COS, as it mixes with the other hot gases, it is ultimately oxidized to sulfur dioxide as it enters the gas collection ducting system. Depending on the concentration of sulfur in the anodes, some aluminum smelters remove SO2 by scrubbing the exhaust gas with either sea water or chemicals. Spent pot lining (SPL) is a waste generated by the aluminum industry during the manufacture of aluminum metal in electrolytic cells. Aluminum cells have a bottom steel cathode shell that is lined with carbon materials and brick insulation. The cathode lining materials eventually become saturated with cryolitic bath, which causes expansion, cracking, and chemical deterioration, requiring that the spent potlining be replaced after 5–10 years. Over the useful life of the linings, the carbon becomes impregnated with aluminum and fluorides, averaging 34% of the spent carbon lining, and cyanide compounds, 400 ppm. Contaminant levels in the refractories portion of linings that have failed are generally low. Aluminum smelter produces 40–60 kg of mixed solid wastes per ton of product with spent cathodes linings being the major fraction. These linings consist of 50% refractory material and 50% carbon. The SPL material is considered a hazardous waste in many countries because it contains significant quantities of absorbed fluorides along with traces of cyanide. The initial disposal of SPL was primarily in landfills.

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However in 1993, the United States Environmental Protection Agency banned its disposal in landfills. Sophisticated plants were then established specifically for the treatment and disposal of SPL in the United States and Canada. Globally, many aluminum smelters today send their SPL to be consumed as fuel in both cement and ferrosilicon furnaces. The calorific value of SPL depends on the carbonaceous content, which varies with the percentage of refractory materials. The burnability of cement raw mix in presence of SPL showed reduction in free lime due to mineralizing effect of fluoride present in SPL. The carbonaceous content of SPL is well burnt at higher temperature and provides additional heat to the system. 2.5.2.7.7 The Development of Inert Anodes Extensive research has been conducted in the aluminum industry to develop a material that can replace the consumable carbon anode used in the electrolysis cells so that the electrolysis product is pure oxygen. These efforts are driven by the increasing emphasis in industry on reducing carbon dioxide emissions. The use of inert anodes in the electrowinning of aluminum would act as a catalyst to oxidize oxygen ions in the melt to oxygen gas and produce aluminum by the following equation: 3 Al2 O3 ¼ O2 þ 2Al 2

ð2:5:56Þ

The focus of the research has been the attempt to find conducting or semi-conducting materials that do not dissolve in the electrolyte under the operating conditions and are stable in the presence of pure oxygen atmosphere. Hitherto, no material has been found that can give long-term stability in conventional cells. The most extensively researched material for the electrode surface has been nickel ferrites and different methods of manufacturing these electrodes have been used. The two extremes being using an oxidized surface of a metallic alloy and the other extreme is to form a cermet material by blending and reacting ceramic oxides. Besides the challenge of developing such electrode materials, a disadvantage of inert anodes is the reversible decomposition potential of alumina to produce aluminum is about 1 V higher with inert anodes compared with the carbon anode process (2.10 versus 1.20 V). This means that the cell voltage would be about 1.0 V higher for inert anodes if the overvoltages were the same. However, it may be possible to offset the higher voltage by changes in cell design that reduces the ACD. 2.5.2.7.8 Aluminum Casting Molten aluminum siphoned from the electrolysis cells is transferred in a crucible to the casting shop where it is poured into large gas or fuel-fired holding furnaces, where the metal is cleaned by skimming, and fluxing gas and/or salt fluxing to remove bath

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and other impurities. The purified molten aluminum is then cast into a variety of forms such as T-ingots and pig ingots to be sold as primary aluminum for remelting. Small 25-kg ingots produced by a chain conveyor casting machine is a common item that are usually sold to smaller companies to remelt and manufacture a wide variety of aluminum products. Many producers manufacture semi-finished cast products such as rolling ingot (slab) and strip, which have been alloyed to meet different fabricating requirements. Large 15–25-ton aluminum slabs that are 1–2 m wide and up to 8 m long are made by a direct-chill vertical casting process (DC) using a water-cooled mold for producing slabs for aluminum-rolling mills to fabricate aluminum can stock. Aluminum slabs are also produced by advanced electromagnetic casting (EMC) technology that eliminates the surface ripples produced by direct-chill casting. The molten metal is levitated by generating an electromagnetic field inside the mold and solidified without contact to a solid surface yielding an ingot free of surface imperfections that do not require scalping to remove rough edges, thus saving time and money in the manufacturing process. Aluminum billets (10–50 cm diameter round and 6–8 m long) are cast to produce extrusion products. These billets are heated in furnaces to homogenizing temperatures about 450–520
C to eliminate the internal stress imposed on the aluminum bar, minimizing the deflection, improving the plasticity, and reducing the extruding resistance. Preheating or homogenizing reduces the chemical segregation of cast structures and improves their workability. Hot strip continuous casting is becoming common for manufacturing of products such as wire, foil, and pharmaceutical packaging. A continuous caster is a combination casting machine and rolling mill. Molten alloyed aluminum is continuously fed into the machine and emerges in solid form as a rod for a rod mill or as a roll stock for a sheet or foil mill. Aluminum metal produced by the Hall–He´roult electrolysis is 99% pure, but contains impurities of iron, silicon, etc. Aluminum can be further refined by using the three-layer electrolysis Hoopes cell. The three layers are separated from each other due to differences in density. The lower liquid layer consists of an alloy of impure aluminum with about 30% copper to increase the density to 3.4–3.7 g/cm2 that serves as anode. The middle liquid layer is the electrolyte consisting of cryolite and barium fluoride with a density of 2.7–2.8 g/cm2. The uppermost layer is the separated high-purity aluminum with a density of 2.3 g/cm2, which serves as cathode. During electrolysis, Al3þ ions from the middle layer migrate to the upper layer where they are reduced to aluminum. Pure aluminum tapped from the cell gives 99.99% pure aluminum. Higher purities of up to 99.9999%, or “six-nines” aluminum, can be obtained using additional zone-refining operations that traps impurities in a molten zone that moves gradually from one end to the other of specially prepared ingots.

Aluminum Production

2.5.3. ALUMINUM RECYCLING While “aluminum recycling” can include any collection and reuse of an existing aluminum structure, object, or piece, the term is generally meant to include a melting step to return the metal to a more common form that can be easily used in a subsequent manufacturing step. In the broadest sense, aluminum recycling deals with aluminum scrap and by-products. To understand aluminum recycling, it is first necessary to understand the uses of aluminum, the alloy families, and manufacturing processes as well as the thermodynamics and surface oxidation reactions that occur.

2.5.3.1. Aluminum Materials Recycled 2.5.3.1.1 Scrap 2.5.3.1.1.1 Major Classifications by Alloy Family

First, let’s take a quick look at the family of aluminum alloys. While aluminum is used in its industrially pure form for some applications such as foils and electrical conductors, it is most often alloyed with one or more other metallic elements to improve the physical properties of the resulting alloy. A few of these physical properties include: strength, ductility, fatigue resistance, formability, and molten fluidity. The primary alloying agents are: Cu, Mg, Mn, Si, and Zn. They can be added singly or in combinations to achieve the desired properties. Typically, additive rates are in the 0.3–12 wt.% range; consequently, alloying elements can be a significant fraction of the entire alloy. Many other minor elements are added intentionally to further modify the physical properties. Figure 2.5.25 shows the primary aluminum alloying elements, how they are combined to form common aluminum alloy families, and the general class of alloy they represent. This recycling discussion is not going to address the reasons for adding various elements (that can be found in various other texts [30,31]). Rather the discussion is to note that they occur and will impact our options when aluminum scrap and drosses are processed. 2.5.3.1.1.2 Sources and Material Flow

Aluminum scrap is generated in many steps of the manufacturing process and use by the end user. During manufacturing, aluminum material is lost from the process either during the melting step to form dross or in one of the many machining operations applied to the aluminum piece. Based on Figure 2.5.25, it makes the most sense to return a scrap material to the same alloy or at least the same family whenever possible. In a manufacturing operation, this is normally a simple operation. When the machining operations take place within the aluminum mill, the scrap is always returned to the cast house and remelted. When the scrap is generated at a subsequent manufacturing operation, it is highly likely that the scrap will return to a recycling operation and will normally be returned to the original alloy since the pedigree is known. This type of scrap has the highest value due to

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Zn

Mg

Al – Cu

2000

Al – Cu – Mg

2000

Al – Mg – Si

6000

Al – Zn – Mg

7000

Age hardening alloys

Al – Zn – Mg – Cu 7000

Al

Cu

Mn

Si

Al

1000

Al – Mn

3000

Al – Mg

5000

Al – Si

4XX

Al – Si – Cu

3XX

Al – Mg

5XX

Al – Cu

2XX

Work hardening alloys

Casting alloys

Figure 2.5.25 Principal aluminum alloy systems.

its known pedigree. Manufacturing scraps could include stamping skeletons, machine turnings, scrapped pieces due to imperfections, or any other form due to a multitude of reasons. Once the part is installed in a larger assemblage or is delivered to a customer, the traceability of the part decreases. Typically, postconsumer scrap is a mixture of alloys and sometimes even a mixture of metals. Some scrap types, like Used Beverage Containers (UBCs), are easily identified, but most postconsumer scrap is at best identified between one of several broad categories. At one time, the use of recycled or secondary scrap was a very small portion of the entire aluminum metal stream. However over the past 20 years in the United States, the contribution of aluminum recycling has remained steady, as domestic primary aluminum capacity has been replaced by imported metal. As an example, US usage of secondary Al in the past 20 years has accounted for 15–20% of the total metal input. The US domestic consumption for aluminum by source is presented in Figure 2.5.26 [32]. The material flow in the global aluminum market in 2010 is presented in Figure 2.5.27 [33]. The red lines and circles show the various recycle paths that occur within the larger aluminum metal flow. Note that the large size of the total aluminum products is still in an active role. As this old scrap comes out of service, it becomes “old scrap” or postconsumer scrap. It has been said that two-thirds of the aluminum ever created is still in use today [32]. Different aluminum market sectors have differing lengths of service. There is some disagreement on the average length of service life, but Table 2.5.4 gives a good representation. 2.5.3.1.2 Drosses The other major area of aluminum recycling is the processing of drosses. Many locations use various names for the mixtures of aluminum metal and aluminum oxide plus various

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10,000

Metal source (1000 mt)

9000 8000 7000 6000 5000 4000 3000

Import

2000

Secondary

1000

Primary

0 93

19

95

19

97

19

99

19

01

20

03

20

05

20

07

20

09

20

11

20

Year

Figure 2.5.26 Source of input aluminum for US consumption.

other components commonly lumped under the general classification of “dross.” To standardize the nomenclature, the Aluminum Association published a pamphlet [35] describing and defining the various forms of drosses. The pamphlet uses four major identifying characteristics: • Originating process source for the dross—type of furnace such as reverberatory furnace, rotary furnace, etc. • Salt content of the dross—how much and what type of salt flux is present in the dross? • Aluminum content of the dross—what is the aluminum content of the dross prior to processing? This value needs to be carefully defined because different methods of analyzing for the aluminum content will give significantly different results in the amount of available or free aluminum. In particular, very small aluminum droplets may be present and can be accounted for by a chemical analysis, but they may be too small to be recovered in a conventional high-temperature recycling process. • Form of the dross—what are the physical size and shapes of the dross, such as chunks, fines, skulls, or blocks? While there are many subcategories of aluminum by-products, there are three main forms of drosses. Each is described in more detail. 2.5.3.1.2.1 White Dross

White Dross is a mixture of aluminum oxide and aluminum metal with the metal content typically varying from 15% to 80%. It is produced during the melting of aluminum in a furnace hearth or the transporting of molten aluminum between furnaces or crucibles.

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Figure 2.5.27 Global aluminum metal flow—2010.

Table 2.5.4 Average Lifetimes of Aluminum Products in Years [34] Product

Years

Transport

12

General engineering

15

Electrical engineering

20

Building and construction

30

Packaging

<1

Home and office

10

Others

10

The dross is collected in a “skimming operation” where an operator either manually removes the floating mixture from the furnace, or in the case of modern large furnaces, an operator on a motorized piece of equipment skims the floating dross. In either case, the dross mixture is removed from the furnace and into a “skim pan” for containment and cooling. Diligence must be used at this point to prevent further oxidation of the hot aluminum metal once it is removed from the furnace. In addition to the two major components, aluminum metal and aluminum oxide, smaller amounts of other compounds can be present, depending upon the conditions

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in the furnace (as well as the subsequent dross pan), the metal source, and the alloy being processed. These other compounds include aluminum nitride (AlN), aluminum carbide (Al4C3), and possibly cryolite (Na3AlF6). The last compound is often associated with molten metal coming from electrolysis cells, while the two former compounds are associated with “thermite” reactions occurring in the furnace or dross skim pan. A thermite reaction is the uncontrolled chemical reaction between metallic aluminum and the gases in the air. A thermite reaction is exothermic and produces extremely high temperatures (1500
C or more) which allow rapid chemical reaction of oxygen, nitrogen, and carbon dioxide with aluminum to form the previously mentioned compounds. The chemical reactions involved in a thermite reaction are: 3 ð2:5:57Þ 2Al þ O2,gas ! Al2 O3,solid 2 ð2:5:58Þ 2Al þ N2,gas ! AlNsolid 8Al þ 3CO2,gas ! Al4 C3,solid þ 2Al2 O3,solid

ð2:5:59Þ

More details on the thermodynamics of these reactions will be given in the Section 2.5.2.7. The presence of these compounds in the dross is noticeable when the cooled dross reacts with moisture. The reaction with water liberates ammonia and acetylene gases which are easily detected by their characteristic smells. In special circumstances, the metallic aluminum can react with moisture to form hydrogen gas which has a high potential for flammability. White dross can include oxide compounds not directly formed in the furnace such as metallic skim removed from cooled transfer troughs, skulls from crucibles, and spills. Finally, white dross contains little or no salt flux. 2.5.3.1.2.2 Black Dross

Many aluminum scrap melters use reverberatory furnaces with external side charging bays. It is a common practice to charge the scrap to the furnace metal pool through this side bay. The bay often contains a salt flux layer to help protect the molten aluminum pool from oxidation and to improve the metal recovery of the scrap. This type of furnace is especially common when processing light gauge or high surface area scraps such as turnings or UBCs. The flux usually consists of sodium chloride (NaCl) and potassium chloride (KCl) with the possible addition of a fluoride salt compound. A mixture of chloride salts is used to lower the melting point of flux. The flux-containing dross resulting from this melting operation is called “black dross” because of its characteristic dark color. The black dross consists of a mixture of salt, oxides, and metal. At temperatures above the melting point of the salt flux mixture, the black dross consists of two immiscible liquid phases: aluminum metal and liquid salt containing oxide particles and films. As the oxide content increases in the salt flux, the mixture becomes less fluid and more viscous in nature. The flux is very efficient in trapping and holding the oxide compounds.

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The metal content of the black dross will vary depending on the scrap type being charged and the processing conditions, but varies from 7% to 35%. The oxide content of the black dross will be approximately equal to the salt flux content. 2.5.3.1.2.3 Salt Cake

Rotary furnace operations processing scraps and drosses use salt flux for the same reasons mentioned earlier, but the ratio of the salt flux to the contained oxide in the resulting byproduct is different. Two modes of operation are common: “wet” and “dry” with the difference being the amount of salt flux added and the rheological consistency of the molten black dross. In the wet mode, enough salt flux is added to make a liquid pool of molten salt with the approximate consistency of water. In the dry process, less flux is added and the salt cake has a much drier consistency. Metal content of salt cake is generally lower than that of black dross. The approximate compositional analysis of the three dross types is shown in Table 2.5.5.

2.5.3.2. Fundamental Thermodynamics of Aluminum Oxidation The element aluminum is one of the most common elements within the earth’s crust, yet unlike many common industrial metals, it is never found in the metallic state. As outlined in the section 2.5.2.7, a great deal of energy under very demanding conditions is required to make metallic aluminum. This is due to the thermodynamic driving force for metallic aluminum to recombine with oxygen. Reaction (2.5.57) can be rewritten in terms of a single mole of oxygen as: 4 2 ð2:5:60Þ Al þ O2,gas ! Al2 O3,solid 3 3 The Gibbs free energy of formation for aluminum per reaction (2.5.60) is 1050 kJ/ mol of oxygen. The high negative value shows that the reaction wants to proceed to the right and go to completion. This negative value is observed over the entire temperature range associated with metallurgical processing. 2.5.3.2.1 Oxide Formation Steps The crystallographic form of the aluminum oxide produced during the oxidation of aluminum metal is very dependent on the reaction temperature and the time at that Table 2.5.5 Approximate Compositions of Aluminum By-Products % Metallic Al

% Aluminum Oxide

% Salt

White dross

15–80

20–85

0–1

Black dross

7–35

30–50

30–50

Salt cake

3–10

20–60

20–80

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temperature. Typically, the oxide will start as an amorphous compound and then go through a sequence of crystallographic reorientations until the final and most thermodynamically stable phase, a-alumina, is formed. The typical progression is shown below: Amorphous Al2 O3 ! gAl2 O3 ! aAl2 O3

ð2:5:61Þ

Unfortunately, the alpha phase is considered “refractory” or inert to further processing. Because of this fact, it is not considered possible to feed the a-alumina back to an electrolysis cell or even to a Bayer plant. The reactivity and solubility is so limited that it is not possible to recycle the material in this manner. Consequently, the pressure to minimize oxidation during recycling operations is high. Once the aluminum is oxidized, it is lost from the metal stream and cannot be recaptured by any reasonable process today. 2.5.3.2.2 Ellingham Diagram Next, we will examine the oxidation thermodynamics for aluminum compared with other common metals. An easy graphical way to do this is through the use of an Ellingham Diagram (see Figure 2.5.28). All oxidation reactions are written in terms of one mole of oxygen reactant. This allows comparison of exchange reactions between different metals and metal oxides. With a few exceptions, most metal oxidation reactions are less negative than aluminum’s reaction. This implies that if metallic aluminum comes in contact with one of these metallic oxides, the aluminum will reduce the oxide and leave behind the other oxide as a metal. Additionally, the reaction will generate heat. This oxidation reaction is the basis of the classic “Thermite reaction” used to weld steel rails by igniting a mixture of aluminum metal with iron oxide. Only magnesium and the alkalis and alkaline earths are more reactive than aluminum. For recycling, this distinction is important because any other metal oxide present during the remelting operation will be reduced by the aluminum. This decreases aluminum recovery and increases the pickup of contamination [37]. Additionally, virtually all metals are soluble in molten aluminum. Once any alloying element is in the molten aluminum (with the exception of Mg), it is not possible to remove the metal by further chemical processing as is so common in other nonferrous systems. As will be discussed later, this phenomenon limits the opportunities to use some scraps in certain applications.

2.5.3.3. Oxidation Reactions and Kinetics 2.5.3.3.1 Solid Oxide Film Growth Aluminum metal will oxidize and form an aluminum oxide skin when exposed to oxygen when either in the solid or liquid form. In the solid phase, the oxide film is normally very tightly adhering which makes it more difficult for oxygen to reach the surface of the metal and to continue the oxidation reaction. Eventually, the oxidation ends (or at least becomes imperceptively slow) and the surface becomes passivated. The oxidation of aluminum on solid materials typically follows parabolic kinetics. Initially, the reaction is very fast and then

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Figure 2.5.28 Ellingham Diagram for metal oxides [36].

begins to slow down as the oxide film grows. This type of reaction kinetics occurs when one species has to diffuse through the film to react with another species and create more of the oxide film. This behavior can be defined by Equation (2.5.65): m2 ¼ Kt

ð2:5:62Þ

where m is the mass of the oxide film, t is the time, and K is the parabolic rate constant.

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Aluminum Production

At room temperature, a stable oxide thickness will soon form. As the temperature increases, the rate constant will increase and the thickness of the oxide layer can increase. In the remelting of scrap aluminum, the oxide layer will increase thickness during the heating and melting process. Longer melting times can lead to thicker oxide layers and lower metal recovery. High surface area scraps are particularly prone to melt loss due to oxidation. The film produced on pure solid aluminum slows down the kinetics of oxidation due to its protective nature. The tightly adhering film is unlike the oxide formation associated with iron which is nonprotective. Different alloying elements can impact the protective nature of the film. This is often explained through the application of the Pillings– Bedworth (PB) ratio. The PB ratio is defined as: PB ¼

ðvolume of oxide producedÞ ðvolume of metal consumedÞ

ð2:5:63Þ

When the PB ratio is below one, the film will shrink and crack. It is no longer protective, and linear oxidation kinetics will prevail. If the ratio is above one, the film has expanded and will be protective. In this case, parabolic kinetics will occur. As we have already pointed out, only a few elements are more reactive than aluminum and will react at the surface. Several of these more reactive elements are listed in Table 2.5.6 along with their PB ratio when forming oxide films. The g-Al2O3 has a PB ratio greater than one which implies that it forms a protective film. Meanwhile, MgO has a PB ratio of less than one and is nonprotective. Many Mg alloys are prone to higher oxidation rates due to the formation of MgO on the metal surface. It has also been known that additions of beryllium can help slow down the oxidation of high Mg alloys during melting and holding in reverberatory furnaces due to the formation of a protective film of B2O3 with a high PB ratio. Similar to Mg, both Li and Na have low PB ratios and can increase the oxidation of aluminum alloys. Also of interest to the practitioner of melting, aluminum scrap is the impact of particle size or sheet thickness on oxidation. For a given mass, thinner gauge sheet or finer particle size will result in more surface area. The entire surface will have a preexisting oxide film Table 2.5.6 Pillings–Bedworth Ratio of Metallic Oxides Oxide

PB Ratio

g-Al2O3

1.49

BeO

1.68

B2O3

3.10

MgO

0.79

Li2O

0.58

Na2O

0.55

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Alton T. Tabereaux and Ray D. Peterson

10 9 8 % Melt loss

7 6

200 mm 20 mm 2 mm 0.15 mm

5 4 3 2 1 0 650

700

750

800

850

900

950

Temperature (°C)

Figure 2.5.29 Impact of sheet thickness and final melting temperature on melt loss.

and this film can grow thicker at elevated temperatures associated with melting the solid scrap. Figure 2.5.29 shows the impact of sheet thickness and final melting temperature on melt loss [38]. For any give sheet thickness, a higher final processing temperature will result in more melt loss. Additionally as the gauge decreases, the amount of melt loss increases due to increased surface area. Another way of showing the impact of sheet thickness on melt loss [39] is given in Figure 2.5.30. Note the exponential increase in melt loss with decreasing gauge thickness. For a given mass of aluminum, the surface area will increase at the third power as the particle size decreases. The shape of the melt loss curve would tend to follow this pattern. Consequently, special techniques and precautions must be taken as the size of the aluminum scrap decreases. 2.5.3.3.2 Liquid Oxide Film Growth Thiele [40] measured the growth of oxide films on quiescent molten aluminum by monitoring the weight gain as a function of time. He also analyzed the forms of the oxide film and how the phases changed with time. For pure aluminum at 700
C, Thiele found that an amorphous Al2O3 formed first. It was not stable, and with sufficient time and temperature, it would convert to crystalline g-alumina. The g-alumina was a protective film which slowed down the oxidation reaction. Finally, the g-alumina would convert to a-alumina at nucleation sites within the existing film. There is a volumetric change associated with this phase change and the resulting stresses induced on the film cause ruptures. With the film ruptures come renewed oxidation termed “break away” oxidation. The

901

Aluminum Production

30

% Melt loss

25 20 15 10 5 0 0

5

10

15

20

Gauge thickness (mm)

Figure 2.5.30 Impact of sheet thickness on melt loss.

newly formed a-alumina particles act as nucleating sites and contribute to even more conversion of the existing g-alumina. Thiele found that the length of time that the protective g-alumina was stable was variable and led to results that were not always reproducible. The typical oxidation progression observed was: Molten Al þ O2,gas ! amorphous Al2 O3 ! gAl2 O3 ! aAl2 O3

ð2:5:64Þ

Examples of Thiele’s oxidation measurements are shown in Figures 2.5.31 and 2.5.32. As the temperature increases, the reaction kinetics increase for the formation and conversion of the oxide. The life of the stable amorphous phase is shortened, and the total amount of the reaction increases. Thiele also looked at the impact of various alloying elements on the oxidation kinetics. He found that most common alloying elements which are more thermodynamically noble than aluminum do not impact the oxidation kinetics at 700
C when added at the 1 atomic % level. Those elements that are more reactive than aluminum (Na, Ca, and Mg) increase the rate of oxidation. These results are presented in Figure 2.5.8. As Figure 2.5.32 shows, the addition of alloying elements can change the oxidation kinetics by changing the resulting oxide film. Of particular interest is the impact of magnesium additions. As noted earlier, Mg is one of the few elements that is more reactive than aluminum. Many investigators have examined the impact of Mg additions on oxidation kinetics of molten aluminum including Balicki [41]. It would appear that amorphous MgO forms rather than amorphous Al2O3 though this is surely impacted by the amount of Mg present in the alloy as shown in figure 34 from Balicki. The amorphous

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120 99.5 (700⬚C)

Weight gain (mg)

100

99.9 (700⬚C) 99.5 (800⬚C)

80

99.9 (800⬚C)

60 40 20 0 0

20

40

60

80

100

120

140

160

180

Time (h)

Figure 2.5.31 Oxidation of various aluminum species at 700 and 800
C.

280 240 Weight gain (mg)

Mg

Na

200

Ca

160 120 Se

80 40

Al 99.99, Fe, Zn, Cu, Be, Mn

Si

0 0

20

40

60

80

100

120

140

160

180

Time (h)

Figure 2.5.32 Effect of 1 at.% alloying additions on the oxidation of aluminum at 700
C.

MgO is not nearly as protective as the film formed from amorphous Al2O3. Consequently, these alloys are much more prone to continued oxidation formation. This can manifest itself as increased dross formation during the melting of Mg alloys. Additionally, the formation path for oxidation is different than in the pure Al case. Molten Al  Mg alloy þ O2,gas ! amorphous MgO ! crystalline MgO þ spinel

ð2:5:65Þ

903

Aluminum Production

After the formation of the amorphous MgO, crystalline MgO is formed as well as a spinel (MgOAl2O3 or Al2MgO4). This oxide layer is less protective and will cause the “break away” oxidation reaction to occur. This oxidation reaction tends to be much more pronounced than the reaction associated with pure Al. Additionally, the spinels that are formed are very detrimental to the physical properties of the final aluminum products and must be removed by filtration, settling, or flotation prior to casting (Figure 2.5.33). Similar oxidation patterns takes place on nonquiescent molten aluminum, but the presence of fluid movement can shear the surface oxide film and allow the oxidation reaction to continue without the benefit of passivation. This renewing of the surface oxide film on a molten surface is the basis for the formation of skim.

2.5.3.4. Dross Processing Options As mentioned earlier, white dross is an intimately mixed combination of aluminum metal and aluminum oxide. The many processes for handling aluminum dross [42] can be generally broken down into one of three categories: A. Melting, B. Crushing/grinding for liberation of Al metal, and C. Low temperature aqueous chemical processing. 100 Al to Al Mg 0.5 Al Mg 1

80

Al Mg 2.5

Weight gain (mg/cm2)

Al Mg 5 Al Mg 10

60

40

20

0 0

20

40 Time (min)

Figure 2.5.33 Oxidation of Al–Mg alloys at 700
C.

60

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Most dross is processed through melting operations so we will start our discussion with them. Most of the techniques we practice today were discovered and developed many years ago; we just refine the techniques to extract higher recovery.

2.5.3.4.1 Melting The earliest efforts of recovering aluminum metal from dross and skimmings typically centered on conventional reverberatory furnaces since they were the most common type of furnace in operation. One of the earliest patents (US 1,180,435) in 1917 centers on the use of a reverb furnace to heat and melt a salt flux mixture which in turn melts and cleans drosses charged into the salt flux to yield clean aluminum metal. Other patents involve intimate mixing of dross with molten metal and a flux layer (US 1,550,192 and US 1,729,631). It was found early on that salt fluxes helped to wet the oxide films and particles, and then remove them from molten aluminum metal. Later, batch reactors were used to treat hot drosses with reactive compounds (salts or gases) and then the whole mixture stirred to liberate the entrapped aluminum metal. Typical of these reactors was US Patent 2,754,199 in 1959 by Stroup and Dowd of Aluminum Company of America. Hot dross is processed in a closed inclined container into which chlorine or aluminum chloride gas can be injected. The closed container is rotated while on an incline. Today, the primary method for processing drosses is through the use of rotary furnaces. The first mention in the patent literature of a rotary furnace for processing aluminum is the Schmidt patent (GB 520,533) in 1940. A fixed-axis wet-salt flux rotary furnace system was described. A NaCl–KCl eutectic flux mixture was used. Currently, two major branches of rotary furnace technology exist: 1. Tilting rotary furnaces using what is termed dry salt flux practices and 2. Tilting rotary furnaces that are operated in a flux-free manner. The flux-free furnaces use heat sources provided by oxy-fuel burners or plasma torches. This type of operation is practiced where there are restrictions on generation or disposal of salt cake. Generally speaking, recovery from this mode of rotary furnace operation results in lower metal recovery. Advantages claimed are the ability to process dross without salt flux thereby avoiding disposal issues associated with salty wastes as well as a reduction in off-gases to clean. This type of operation is more common in Canada and Europe. The alternative rotary furnace technology being employed is often referred to as a “dry flux” tilting rotary furnace. A tilting rotary furnace allows charging and emptying through the mouth of the rotary barrel furnace. With this design, it was possible to charge large pieces of dross and to pour out materials that were more viscous. Rather than requiring a high ratio of salt flux to charge material to ensure slag fluidity as the earlier fixed-axis rotary furnace required, the tilting rotary furnaces were able to use other salt flux ratios and decrease flux consumption. The performance of this furnace type was very dependent upon operator skill and the type of support equipment used. The tilting rotary furnace is used to process almost every type of aluminum dross and scrap due to their processing versatility (Figure 2.5.34).

Aluminum Production

Figure 2.5.34 Example of a tilting rotary furnace.

The salt flux used plays a critical role in the conventional processing route for drosses and scraps. It serves three primary functions: 1. Strips away the oxide film from the aluminum droplets, 2. Traps the oxide films in the salt flux rather than in the molten aluminum, and 3. Protects the aluminum metal from further oxidation. The salt flux is typically a mixture of NaCl and KCl and a smaller amount of cryolite. A mixture of chloride salts is used to achieve a lower melting point of the flux and if cost were not a consideration, the eutectic blend would be used. Only those chloride salts more stable than aluminum chloride can be used as flux components. This limits the selection to those salts composed of alkali and alkaline earth elements. Furthermore, when cost considerations are taken into account, only a few chloride options exist. As is known from aluminum electrolysis, cryolite is an effective solvent for aluminum oxide. In the smaller amounts used in recycling of drosses and scrap, the cryolite can help to strip away the oxide film surrounding the molten aluminum droplet and allow coalescence of the aluminum droplets. To address the concern of salt cake disposal, several patents were developed in the late 1970s and early 1980s for the reprocessing of salt slags. With the increased use of wet-flux rotaries and their high salt flux demands, the need for a method of recapturing and reusing the salt flux became apparent. Patents by Papafingos and Lance (US 4,073,644) and Granthan and Johanson (US 4,379,718) are examples. Most development of closed loop recycling has been performed in Europe where governmental rules and societal pressures drove the demand.

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2.5.3.4.2 Mechanical Processing Following the lead of existing mineral processing systems, early technologists used crushing, grinding, and screening to liberate the metallic metal from the oxides. The first mention of crushing as a method to separate aluminum metal from dross was by Gills (GB 114,204) in 1918 as part of a larger process to fully treat white drosses. The next patent (US 3,037,711) devoted entirely to liberation of aluminum metal by crushing and grinding was issued in 1962 to Businger of Metallwerke Refonda, Weiderkehr & Co for a dry milling process which uses a mill to flatten aluminum pieces which are then separated from the friable oxides by screening or air classification. This basic process is still used today by some operations. Advantages for this processing method are the avoidance of thermal or chemical processing and the normally expensive equipment associated with these processes. Disadvantages associated with crushing and grinding drosses are the dusts created if dry crushing or the production of a reactive material in an aqueous slurry that can liberate troublesome gases. 2.5.3.4.3 Aqueous Chemical Processing The Gill patent of 1917 was the first patent to mention aqueous processing of the dross to form an aluminum compound. Later other patents were granted that focused on recovering the chemical value of aluminum through aqueous processing. In 1927, J.G.G. Frost of National Smelting Company (US 1,648,262) taught a method of processing white dross with water and sulfur dioxide gas to form aluminum sulfite which becomes a feed stock for other aluminum compounds. Other patents such as US 1,962,498 teach the method of creating aluminum sulfate from dross. In 1976, US Patent 3,955,969 discloses closed loop processing of salty wastes and the creation of calcined alumina from the oxide portions of the dross. Huckabay and Skiathas received a patent (US 4,252,776) in 1981 for an aqueous process operated at elevated temperatures to convert both white and black drosses to aluminum trihydrate. Advantages of aqueous processes are that they are generally not energy intensive and that all material is usually consumed in some manner. The disadvantages are that sophisticated processing systems are usually required to handle the off-gasses, quality of the final products can be impacted by the variable nature of dross sources, and the economics for processing in the methods are often disadvantageous compared with conventional thermal processing.

2.5.3.5. Scrap Processing The world of scrap processing includes many forms and types of scrap. The difficulties associated with processing and achieving high recovery for the scrap are impacted by three primary factors: particle size, contamination, and magnesium content. The impact of magnesium on oxidation formation and therefore melt loss has already been discussed.

907 Clean

Aluminum Production

Hot mill trim

Cleanliness

Bars coils Crushed cast scrap

Class II Borings and turnings

Oily foil Di ffi l cu

UBC

Coated coils

Painted siding

t

Dirty

y

Class I

Reject plate Cracked ingot and RSI sow

Extrusion scrap

s Ea

Clean Cold mill foil trim

Small

Size

Large

Figure 2.5.35 Relative comparison of various scrap types for ease of processing.

Figure 2.5.35 illustrates the impact of the other two variables. Large clean pieces of aluminum are relatively easy to melt. That is why we see T-bar ingot and Recycled Secondary Ingot as a common material entering aluminum melt shops. As the scrap size decreases or becomes dirtier, the difficulty of processing increases and more extraordinary measures must be taken. In the following sections, we will discuss scrap preprocessing and melting systems. It will become apparent that more effort is required to process the difficult materials than the “easy” scrap. 2.5.3.5.1 Preprocessing of Scrap Scrap is processed before melting to remove impurities that have deleterious impact downstream such as water, organic coatings, and tramp metals. Many of the unit operations used are based on standard mineral processing techniques.

2.5.3.5.1.1 Shredding

Postconsumer scrap often has other materials mixed in or attached that are not desired. A shredding operation in conjunction with other equipment is used to separate the unwanted materials. As an example, steel screws and rivets can be removed from aluminum sheet and tube by shredding to a sufficient size followed by a magnetic separator. Adhering dirt can be pounded off of scrap while other debris caught inside of aluminum containers can be released by breaking apart the containers and screening the resulting material. Several types of shredders are commonly used, but most involve large rotating hammers that hit the aluminum pieces at high speed. The pieces are repeatedly fractured until they are small enough to exit the shredder through appropriately sized grates.

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2.5.3.5.1.2 Screening

Screens are used to remove fines from aluminum scrap especially after shredding. Typical uses for screens include removing dirt and broken glass from shredded UBCs or shredded automotive parts. Screening can also remove lead and carbon from delacquered UBC. 2.5.3.5.1.3 Magnets

Magnets are useful for removing ferrous contamination from aluminum scrap. Two common pieces of equipment are: 1. The head pulley magnet used to separate magnetic materials from nonmagnetic aluminum scrap as it falls off the end of a conveyor belt and 2. The cross belt magnet which usually employs a strong electromagnet placed above a conveyor. The cross belt magnet has its own conveyor placed perpendicular to the belt. When a ferrous piece encounters the magnetic field, it is lifted off of the conveyor toward the electromagnet whose own inverted conveyor belt sweeps the piece away and into a collection bin. While magnets are effective for most ferrous materials, they do not remove some stainless steels which are weakly magnetic. When using magnets, it is important to correctly select the particle size from the shredder so that the iron is liberated from the aluminum and does not carry valuable aluminum into the reject pile. 2.5.3.5.1.4 Rare Earth Magnets and Eddy Current Machines

Stronger magnetic fields can be created with rare earth magnets. These are often capable of capturing and separating those stainless steels that are weakly magnetic. Eddy current machines contain a spinning cylinder composed of alternating permanent magnets. When this cylinder is spun quickly, the alternating magnetic field produced by the magnets will induce an electrical current in an electrically conductive material. When a conductive material like aluminum or copper passes over the rapidly spinning eddy current magnet placed at the end of conveyor belt, the conductive piece generates a counter magnetic field which will cause it to levitate away from the belt and follow an elevated trajectory off of the end of the conveyor belt. The nonconductive materials do not experience a levitating effect and will simply fall off of the belt in a normal trajectory. This equipment can separate out wood, plastic, foam, rubber, glass, and any other nonconducting material from the aluminum stream. It is critical to remove all ferrous materials from the scrap stream prior to processing with the eddy current machine. 2.5.3.5.1.5 Air Knife

An air knife directs a strong blast of air at a falling stream of scrap and separates the particles based on their density and cross-sectional diameter. Average particles will experience some deflection of their falling trajectories. Heavy materials like brass and lead will be

Aluminum Production

less affected by the air stream and can be removed from the shredded scrap since they will be deflected less. Very light materials like plastic films will report to the bag house. 2.5.3.5.1.6 Centrifuge

Machine turnings are usually contaminated with a cutting fluid that is a mixture of water and an oil emulsion. A centrifuge spins the turnings at high speed to extract the liquid. Depending upon the turnings’ shape and size, residual liquid will remain, but it is usually low enough that the materials can then be safely charged to a furnace. 2.5.3.5.1.7 Dryer/Delacquering

The thermal alternative to a centrifuge is a dryer. Typically, natural gas is combusted to generate hot air and the hot air is mixed with the turnings to remove both the water and oils. Various devices are used to mix and contain the turnings and the hot air. Most common is a rotary kiln, but screen decks and fluidized beds have also been used. A higher temperature version of the rotary kiln dryer is used to remove oils and painted coatings. In this case, the operating temperature is higher, and control of the kiln atmosphere and temperature are critical. This device is often used for removing coatings and lacquers from UBC prior to melting. While additional expense and complexity is added to the process, improved metal recovery will pay for the additional costs. 2.5.3.5.2 Scrap Melting Technologies Now we will look at various types of melting furnaces and their common uses. 2.5.3.5.2.1 Induction Furnace

When electrical current is passed through a conductor, it will produce a magnetic field. If another conductor is influenced by the magnetic field, an electric current will be induced in that material. When the applied current is alternating, the receiving material will begin to heat up due to internal electrical resistance heating. The induction furnace takes advantage of this phenomenon and can melt any conductive material and is especially useful for metals. These furnaces range in capacity from a few kilograms to thousands of kilograms. Their advantages include no hot flue gases, no burner systems required, and clean operation. On the negative side, the same induced current that supplies the heat also creates significant flow in the molten metal. In the case of aluminum, any existing oxides or oxides created during the melting process become entrained in the molten metal. This can lead to quality issues and also gives the appearance of higher recovery since the oxides are weighed up as part of the final metal. 2.5.3.5.2.2 Dry Hearth Furnace

A conventional dry hearth furnace is one of the most common furnaces used for melting aluminum scrap. This furnace essentially consists of a refractory lined steel box equipped

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with a combustion heating system. It is often called a reverberatory furnace due to the hot walls and ceiling radiating heat into the scrap. Scrap is placed in the box, and the burner system (usually natural gas or fuel oil) is used to heat the metal to the melting point and some bit beyond. The products of combustion exit from the furnace through the flue. When the metal is fully melted, a tap plug is removed and the metal leaves the furnace by gravity assisted drainage. Many variations on this theme exist: the shape of the furnace can vary, the roof can be removed for quick loading, the metal can be stirred with an electromagnetic stirrer, and the furnace may tilt to more closely control the pouring of metal to improve metal quality. The primary disadvantage of this style of furnace is that the scrap is in direct contact with the high-temperature products of combustion. This style of furnace is best for large pieces of aluminum. It is preferred practice to place the lighter gauge scrap on the bottom of the empty furnace and then place the heavier scraps on top to minimize oxidation. Another disadvantage is that only dry scrap can be placed in the furnace; even after the furnace is drained, the presence of minor amounts of aluminum could still lead to an explosion if the water was present in the scrap. 2.5.3.5.2.3 Side Bay Melter

If we take a conventional reverberatory furnace and add an external box to contain metal, we have created a side bay melter. In this configuration, the hearth part of the furnace contains the burner–flue configuration which supplies the heat to the metal, but in this case, the hearth is always filled with molten metal. The external bay is usually open to the atmosphere, and scrap is charged into the molten metal pool. Melting occurs by conduction of heat from the molten metal to the solid scrap. This avoids all flame impingement and results in a higher recovery for light gauge scraps. Often a quick submergence technique is used to provide even faster melting and less oxidation. Since light gauge scraps from postconsumer sources are often processed in this type of furnace, salt flux is used in the side bay to help metal coalescence and to improve metal recovery. The side bay is normally connected to a fume capture and environmental system to handle any fumes or smoke that might be generated during the melting process. Additionally, it is necessary to use some form of molten metal pumping to bring the hotter metal from the hearth into the side bay so that it can come in contact with the scrap. A circulation system is established by the use of a mechanical pump constructed of graphite or “high tech” ceramics, an electromagnetic pump, or a permanent magnet pump. The expectation is to have enough volumetric flow to create the equivalent of 5–10 complete exchanges of the metal per hour. This maximizes energy and melting efficiency of the furnace. This style of furnace is one of the most common in the scrap processing industry due to its high melting capacity and low operating cost in comparison with other systems. In the processing of UBCs, this furnace is often coupled with a decoating or delacquering kiln which dries and removes the organic coatings of the shredded UBC. The hot UBCs are directly charged into the side well along with the correct amount of salt flux.

Aluminum Production

The same side bay furnace is used for processing scrap used to produce foundry alloys. Various scraps are selected to produce the correct final chemistry and are charged through the side bay until the full charge is created. Once the furnace is full, approximately one-third of the metal will be removed from the furnace. A minimum amount of metal is required to maintain the necessary head for the molten metal pump and for a thermal mass. 2.5.3.5.2.4 Sweat Furnace

A special subset of dry hearth reverb furnaces is the sweat furnace. Some aluminum scraps, especially from the postconsumer processing of automobiles, contain significant amounts of iron and steel. If these scraps were melted in a conventional furnace, the iron would dissolve and raise the iron levels. In a sweat furnace, the scrap is placed across the inclined floor of the hearth. As the furnace heats up, the aluminum begins to melt and drips away from the iron pieces. After a sufficient amount of time has passed, the aluminum has melted and drained away from the remaining steel pieces which are now raked or removed from the furnace. The cycle is then repeated. 2.5.3.5.2.5 Rotary Furnace

For the processing of aluminum drosses and low recovery materials, the rotary furnace is the common work horse of the aluminum industry. While maybe not the optimal melting device for high recovery scraps, they can also be used with only a small recovery penalty. The good mixing and high energy efficiency allow the processing of many materials. Rotary furnaces for processing scraps come in several styles. Originally, rotary furnaces used in scrap processing rotated about a fixed horizontal axis. The scrap and flux were charged through the furnace mouth and then the burner was swung into position. The material would tumble in the barrel and be directly exposed to the heat from the flame well as indirectly by the heated walls. After the charge had melted and reached the correct temperature, a tap plug would be opened and the metal would be drained out from below the salt slag. In order to remove the salt slag, a pouring door would be opened on the furnace end or wall and the viscous slag would be emptied out. The amount of flux required was dictated by the amount need to keep the slag fluid as it became “loaded” with oxides. The second design of rotary furnaces using the dry flux process has already been discussed in Section 2.5.3.4. The tilting rotary furnace has replaced most of the fixed-axis furnaces due to its higher energy efficiency, lower operating cost, and superior recovery. 2.5.3.5.2.6 Multichamber Furnaces

Some furnace manufacturers are designing and building furnaces that can handle dirty or heavily contaminated scrap within a single unit while still being energy efficient. Typically, the furnace consists of two or more chambers where different reactions take place. Typically, a first chamber is utilized where the scrap is heated and the organics are

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pyrolyzed. The organic rich gases with heat value are transported to a second chamber where they are combusted along with natural gas to provide heat to molten metal and the hearth area. The heat in this second chamber is used to provide heat to the first chamber. At some point, the now hot scrap is mixed with the hot molten metal either by a molten metal pump or by simply pushing it forward into the pool. The combustion gases are passed through some sort of regenerative or recuperative burner to capture a portion of the heat leaving the furnace that would normally be lost. These furnaces are very complex, but offer high energy efficiency within a single piece of equipment. 2.5.3.5.3 Scrap Melting and Blending As mentioned earlier, the many aluminum alloys tend to be divided into two major categories: • Wrought alloys and • Casting alloys The major compositional distinction is the amount of silicon present. Within the wrought category, there are six major families, and within the cast category, there are four major families. There is overlap within the families as material scientists change compositions and thermomechanical processing paths to achieve the desired physical properties. 2.5.3.5.3.1 Recycling by Alloy and Family

When processing scrap materials, the preference is to recycle the material back into the same alloy and product. For some scrap products, this is easier than others because more is known about the alloy chemistry. Manufacturing scrap is generally easy to return to the same alloy family because the scraps are typically segregated by alloy at the manufacturing operation. If more information is known about a scrap’s chemistry, a higher price will normally be paid because it is easier to blend. Mixed scraps are always sold at a discount, so manufacturers have an economic incentive to segregate. In the case of postconsumer scrap, usually less is known about the chemistry because the scrap is usually mixed in nature. There are a few exceptions to this rule like UBC which is composed of the same two alloys and is easily identifiable. Postconsumer scrap is often sorted to some extent by scrap collection facilities. Wrought and cast alloys are separated and sometimes easily identifiable scraps like extrusion scraps (6XXX) may be separated from the wrought. If enough information is known, then typically the scrap will be returned to the same general family. This type of recycling is illustrated by the inner recycle loops from end-of-life product in Figure 2.5.36 back to the same alloy family. The figure also shows how the scrap moves through the aluminum industry. Prime metal is alloyed with other elements to make wrought alloys. At the end of the life, the metal is returned to the same wrought alloy family. A similar recycling loop exists for the cast alloy family. Some cast alloys are made with prime metal and silicon. However, we see that some wrought alloy is removed from its loop and enters the cast family loop. This can occur for several

Aluminum Production

Figure 2.5.36 Progression of alloys in recycling.

reasons—the need for scrap where the foundry is located pulls in other materials, or the wrought scrap may be too contaminated to allow return to the original family. There is also a portion of both loops that is lost due to nonrecycling. When an alloy becomes too contaminated to be used for a conventional alloy, the material can be used for deoxidant in the steel industry. At this point, the alloy is being used for the chemical content of the aluminum to attract oxygen in the molten steel. This is the end of the line for the aluminum. 2.5.3.5.3.2 Blending

For aluminum recycling to be successful, we must be able to use the recovered metal into a desirable alloy. Manufacturing scrap is generally well segregated and consists of a known chemistry. This scrap is easily recycled back to the original alloy. With postconsumer scrap, the challenge to place the recovered metal into a standard alloy is more difficult due to the variability in chemistry of the scrap. Various scrap types are described by ISRI [43], but the classifications are broad enough that the chemical variability can be significant. An example of variation in Fe content for old painted siding [44] is presented in Figure 2.5.37. The range of Fe values shown in the histogram is significant. Some of the variability is due to varying amounts of different alloys used to make a common product. Another portion of the variability is due to the presence of varying amounts of tramp iron as an impurity in the scrap. Similar concentration variation is seen for other elements as well. Typical sampling methods are of dubious value. Small samples will not represent the truck load sized chemistry due to the heterogeneity of the scrap. Larger sampling system can become cost prohibitive.

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Number of observations

20

15

10

5

0 0.5

0.53

0.56

0.59

0.62

0.65

0.68

0.71

0.74

0.77

0.8

0.83

Fe lower limit interval

Figure 2.5.37 Fe distribution in old painted siding scrap [14].

When the chemistry of the scrap is known to a high degree such as for manufacturing scrap, blending calculations are mathematically straightforward. When knowledge is lacking, the problem becomes more difficult. Statistical models can be used to predict final chemistry windows with a certain degree of confidence. Alternatively, blending can be performed in a stepwise fashion where scraps with the largest degree of uncertainty are added first in the chargemaking process followed by better quantified scrap later as the entire load nears its target weight and chemistry.

2.5.3.6. Energy Considerations As detailed in the Section 2.5.2.7 on primary production of aluminum, significant quantities of energy are required to produce aluminum metal from alumina and even greater quantities of energy if the starting reference is bauxite ore in the ground. Conversion of alumina to aluminum metal requires approximately 13 kWh/kg (46,800 MJ/t) of electrical power, while the entire process for the North America Aluminum Industry to produce aluminum metal from bauxite in the ground requires 186,262 MJ/t [45]. The recycling process requires much less energy to convert the aluminum scrap back into a reusable form. An often quoted industry standard of 2000 Btu/lb (2110 MJ/t) typically applies to just the melting and casting of the recovered aluminum. This value is very dependent upon the melting technology used. On the low end of the energy consumption spectrum, rotary furnaces with oxy-fuel burners can operate as low as 580 MJ/t, while inefficient reverb furnaces might require as much as 4220 MJ/t.

Aluminum Production

As a comparison, the recycling and recovery of aluminum from scrap require approximately 5% of the energy that the original process used. Obviously, this value depends on which technology is being used to process the scrap and on which basis the original production is used. To be fair, the energy requirement for converting scrap and drosses back to useable aluminum is really more than just the melting process. Energy is also used in the collection and transportation of the feed materials to a recycling processor. Additionally, various preprocessing steps may also be employed which will require energy input. In the same Life Cycle Analysis study, quoted above, the total energy consumed in the recycling process in North America was stated to be 11,690 MJ/t. This basis of comparison shows that only 6.3% of the total energy required to produce aluminum ingot from bauxite is required to convert scraps back to ingot form. As in any Life Cycle Analysis study, the setting of boundaries for the analysis can be critical. Finally, the industry continues to look at methods to reduce energy consumption during the processing of the aluminum scraps and by-products. This desire to lower energy usage is tempered by the desire to maximize aluminum metal recovery. The oxidation of aluminum releases heat and in theory oxidizing just 4–5% of the aluminum could provide sufficient heat to melt the entire load. The value of the metal is much greater than the value of the heat generated through oxidation, so every effort is made to ensure that lower energy consumption for melting aluminum comes from good thermal practices rather than at the expense of recovery. As shown earlier in the chapter, aluminum metal requires significant amounts of electrical energy in its production. As with any resource, it behooves us to practice sustainability and recover and recycle all aluminum scraps and by-products. Recycling aluminum makes economic and environmental sense.

GLOSSARY Alumina Aluminum oxide (Al2O3) Anode effect The decomposition of ionic fluoride containing species at the anode carbon surface generating perfluorocarbon (PFC) gases, CF4 and C2F6. Bath A molten electrolyte mixture typically containing 80–85% cryolite, 1–4% alumina, 10–12% aluminum fluoride, and 5–7% calcium fluoride. Boudouard reaction The equilibrium reaction of CO2 with carbon to form CO at a specified temperature. Dross The general term covering various types of aluminum metal and aluminum oxide mixtures produced during the melting of aluminum. Liquidus temperature The temperature at which the electrolyte starts to freeze (solidified). Magnetohydrodynamics (MHD) Magnetic forces generated in conducting liquids (aluminum pool) due to the interaction of magnetic fields and DC electrical current. Pillings–Bedworth Ratio The ratio of a metal’s volume divided into the metal’s oxide volume. Point feeder A mechanic device that breaks a hole in the cell’s alumina crust and add from 1 to 2 kg of alumina powder to the molten electrolyte.

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Potline A long row of aluminum electrolysis cells, typically 240–330, located in one or two rooms that are connected in an electrical DC series circuit with the transformer and rectification systems situated at one end. Søderberg cell An aluminum electrolysis cell that has one large continuous self-baking anode. Superheat The difference between the operating temperature and the liquidus temperature of the electrolyte.

REFERENCES [1] A.T. Tabereaux, Light Metals Age, 68 (1) (2010) 22–24. [2] K. Grjotheim, H. Kvande, Introduction to Aluminium Electrolysis, second ed., TMS, Warrendale, PA, USA, 1993, p. 42. [3] A. Solheim, A˚. Sterten, in: J. Thonstad (Ed.), Proceedings of the Ninth International Symposium on Light Metals Production, NTNU, Trondheim, Norway, 1997, pp. 225–234. ˚ . Sterten, I. Mæland, Acta Chem. Scand. 39 (1985) 241–257. [4] A [5] E. Skybakmoen, A. Solheim, A˚. Sterten, Light Metals (1990) 317–323. [6] P.A. Foster, J. Am. Ceram. Soc. 58 (1975) 7–8, 288–291. [7] H. Kvande, Electrochim. Acta 25 (2) (1980) 237–240. [8] G.S. Picard, F.C. Bouyer, M. Leroy, Y. Bertaud, S. Bouvet, J. Mol. Struct. 67 (1996) 368. [9] A˚. Sterten, Electrochim. Acta 25 (1980) 1673–1677. [10] R.G. Haverkamp, B.J. Welch, Chem. Eng. Process. 37 (1998) 177–187. [11] N. Dando, X. Wang, J. Sorensen, X. Weizong, Light Metals (2010) 541–546. [12] D.S. Severo, A.F. Schneider, E.C. Pinto, V. Gusberti, V. Potocnik, Light Metals (2005) 475–480. [13] N.E. Richards, B.J. Welch, Proceedings of the First Australasian Conference on Electrochemistry, Sydney, Australia, 1964, pp. 901–922. [14] B.J. Welch, N.E. Richards, in: AIME International Symposium on the Extractive Metallurgy of Aluminium, New York, 1962. [15] S.M. Hume, M. Utley, B.J. Welch, R.C. Perruchoud, Light Metals (1992) 687–692. [16] J. Zoric, I. Rousar, J. Thonstad, Light Metals (1997) 449–456. [17] H. Guldbrandsen, N. Richards, S. Rolseth, J. Thonstad, Light Metals (2003) 323–328. [18] A. Tabereaux, N. Richards, C. Satchel, Light Metals (1995) 325–333. [19] M.M.R. Doreen, Cell performance and anodic processes studied by gas analysis, Ph.D. Thesis, University of Auckland, New Zealand, 2000. [20] M.M.R. Dorreen, D.L. Chin, J.K.C. Lee, M. Hyland, B.J. Welch, Light Metals (1998) 311–316. [21] B.J. Welch, in: 10th Australasian Aluminium Smelting Technology Conference, Launceston, Australia, October 9–14, 2011, 2011. [22] J. Marks, C. Bayliss, Light Metals (2010) 259–263. [23] W.E. Haupin, W.B. Frank, in: Electrometallurgy of Aluminum, Electrochemical Processing, vol. 2, Plenum Press, New York, 1981, pp. 301–325. [24] R.D. Peterson, A.T. Tabereaux, Light Metals (1986) 491–499. [25] W. Haupin, Light Metals (1998) 531–537. [26] K. Grjotheim, C. Krohn, M. Malinovsky, K. Matiasˇovsky, J. Thonstad, Aluminium Electrolysis— Fundamentals of the Hall-He´roult Process, second ed., Aluminium-Verlag, Dusseldorf, Germany, 1982, p. 114. [27] J. Hives, J. Thonstad, A˚. Sterten, P. Fellner, Light Metals (1994) 187–193. [28] A.H. Mohammed, M.M. Al-Jallaf, A. Kumar, Light Metals (2009) 509–513. [29] T.M. Hyde, B.J. Welch, Light Metals (1997) 333–339. [30] J.R. Davis (Ed.), ASM Specialty Handbook: Aluminum and Aluminum Alloys, ASM International, Ohio, 1993. [31] D. Altenpohl, Aluminum: Technology, Applications and Environment, Wiley, New York, 1998. [32] Aluminum Association Website. http://www.Aluminum.org.

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[33] Global Aluminium Industry Sustainability Performance 2010, International Aluminium Institute, 2011. [34] G. Rombach, Light Metals (2002) 1011. [35] Guidelines and Definitions, By-Products of Aluminum Melting Processes, pamphlet by Aluminum Association. [36] University of Cambridge Website. http://www.doitpoms.ac.uk/tlplib/ellingham_diagrams/ ellingham.php. [37] R.D. Peterson, Light Metals (2013) 941–946. [38] H. Rossel, Light Metals (1990) 721–729. [39] L. Stewart, J.G. McCubbin, Melting Aluminum and Aluminum Alloys, Proceedings of the First International Aluminum Extrusion Technology Seminar—1969, March 3–5, 1969, New Orleans, LA, Aluminum Company of Canada, Ltd, Ontario, Canada, 1977. [40] W. Thiele, Aluminium 38 (1962) 707. [41] S. Balicki, Prace Inst. Hutn. 10 (4) (1958) 208. [42] R.D. Peterson, L. Newton, Light Metals (2002) 1029–1037. [43] Scrap Specifications Circular 2012, ISRI, p. 7. [44] R.D. Peterson, Light Metals (1999) 855–859. [45] A.S. Green, Aluminum Recycling and Processing for Energy Conservation and Sustainability, ASM International, Ohio, 2007, pp. 33–66.

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