Byproduct Processing And Utilization

Ion Wast slag produce! is around 100 million tons per year (about 10% vrt of the global

FeiOj

7-9

3

Si02

10-18

22

AljOj

3.3-3.6

5.5

CaO

35-37

64.1

MgO

7-11.5

1.4

MnO

4-6.5

-

VjOs

0.87

-

TiOz

0.76

-

PjOs

1.71

-

SO,

-

2.1

Table 9,6. Types of

Table 9,6. Types of

Type of oxidizing

Method of cooling

Types of

Abbreviation

Ai"gran°la'edS,ag

to about 200 °C

Only fine

AGSF

Water-granulated

coarse aggregate6

WGSF WGSC

Acid-cooled slag

Slowly cooled slag (at the

Coarse aggregate

ACSC

Improved slag

C^rsTaSÎatc

IWGSF 1WGSC

of EF

Type of slag

Major mineral composition

Minor mineral composition

Iron chromite (FeCr204)

Magnesioferrite (MgFe204)

Rankinite (Ca3Si207) Gehlenite (Ca2Al2SiO,)

Water granulated slag

Wustite (FeO) Gehlenite (CajA^SiO?)

(B-Ca2Si04)

oxide (MgFeA104) (Ca,Mg)Si04

(B-Ca2Si04)

Gehlenite (Ct2Al2Si07)

Magnetite (Fe304)

Wollastonite (CaSi03)+

Wustite (FeO) Gehlenite (CajAljSiO,)

(B-Ca2Si04)

Kirschsteinite (CaFeSi04)

oxide (MgFeA104) (Mn3Cr2Si3012)

Tabk 8.3), pig iron (1.6-1J % P) results in a slag rich in

P2Os content of the slag exceeds 12 high in the first slag and it can be used in tent is less and it may be

Ca0.7Al20j + 12 H20 CaO. A1203,6H20 + 6 A120,.H20 (9.4)

  • SiOCa + M2+ =SiOM + CaJ+ (9.5)
  • SiOCa + M2+ =SiOM + CaJ+ (9.5)

2 A1 (in the dross) + 2 NaOH + 2 H20 -» 2 NaAlOj+3 H2 (9.6) NaAlCh + 2 HjO -> Al(OH), + NaOH (9.7)

Processing of Dross 353

Percipiteted Aluminum OxideSterile Processing Process Flow Chart
Figure 9.9 Generalized flow diagram for the processing of salt cake produced in the treatment of aluminum dross (Pickens 2000)

processes (leaching and evaporation) allowing for more flexible operating schedules. Clarified brine is next pumped into a submerged combustion evaporator. Excess combustion air is used. The evaporator operates at 99 0C. Salt precipitated is removed by constant circulation through crystallizer tanks, which provides cooling and lower velocities to allow the salt formed to accumulate in their conical bottoms. The salt slurry is then sent to a centrifuge for dewatering and recovery. The dewatered salt retains approximately 165 % moisture. It is mixed with potassium chloride (about 10 % of the

Zn Fe Pb Mg A1 Na K Ca 0.805 1.37 0.12 1.64 3.92 5.97 4.90 42.11 40.18 2.12 9.7 0.06 1.85 11.48 7.95 0.23

rce in Japan

2000). Acid ngin 10%

rce in Japan

2000). Acid ngin 10%

Recovery Ferric Chloride Flow
Figure 9.14. Flow sheet of the fly ash

2 NaCl + 6 Si02 + A1203 + H20- 2 HCl + 2 Na2AlSi30, (9.8) Pb(Zn)0 + 2 HCl - Pb(Zn)Cl2 + H20 (93)

6 NH4Fe3(S04)j(OH)2 - 6 NH3 + 2 H20 + 5 Fe203 + 4 Fe2(S04) (9,10)
Metallurgical Integration

Chapter 10

RESOURCE RECOVERY FROM PROCESS WASTES

1), This could be a feed stock for preparing ferric of use in industry (Rao et al, 1991).

  • Dry) tons per tout
  • Dry) tons per year
  • Dry) tons per tout
  • Dry) tons per year

11.4

Rutile

68,400

26.7

Leucoxene

160,000

7.9

llmenite

59,000

10.3

Zircon

61,800

1.3

Rare earths

Silicon 2.4 2.6 18.7

Titanium 36.7 40.7 2.4

Silicon 2.4 2.6 18.7

Titanium 36.7 40.7 2.4

M RECOVERY/FECUNI9

M RECOVERY/IM

X M RECOVERY/STERILE CONTROL

„QU NlfilWnO/FECUNlS

--0-

KftMDIH

Hft RATIO 1 STERILE CONTROL

Figure 10.7. Leaching of nickel from pyrrhotite tailings with and without Thiobacillus ferrooxidans; kinetics of nickel recovery and selectivity of solubilization (Tackaberry et al., 1998)

Figure 10.7. Leaching of nickel from pyrrhotite tailings with and without Thiobacillus ferrooxidans; kinetics of nickel recovery and selectivity of solubilization (Tackaberry et al., 1998)

pH VALUE/FECUMS pH VALUE/M -A-pH VALUE/STERILE CONTBOL ..D- pH AFTER AODFN./FECUNB ~Q.. pH AFTER ACIOFN. /M pH AFTER ACIDFN./STERILE CONTOOL _pH AFTER ACIDFN. - pH IMMEDIATELY AFTER STANDARO H2S04 ADDTHOM_

Figure 10.8. Leaching of nickel from pyrrhotite tailings with and without Thiobacillus ferrooxidans (Tackaberry et al, 1998)

Ferrooxidans
sofvent system (Cole and Nagel, \99lf 8 P ° 8

Equilibrium pH

NaBH4 + 4Me2+ + 2h20- NaB02 + 4Me0 + 8tf" (10.12)
Nstse 2015 Class Questions
CuJ+ + CsHioOs + HjO Cu(s) + C3H10O6 + 2lT (10.14)
  • m-fe-Cl-S-W-Al-si-S-ÂT~
  • 60 0.3-0.7 4-9 30-48 3-8 0.3 3-5 1 0.1-0.2

controlled by addition of caustic soda to prevent formation of chlorine. Final cobalt concentration in the filtrate is less than 1 mg/L. The cobalt is precipitated in a different part of the plant more suitable for the use of chlorides and chlorine formation. Zinc co-precipitation is significant at this stage, therefore a purifying stage is introduced to selectively re-leach zinc while upgrading the precipitate in cobalt (typically more than 47% Co). The re-leached zinc units (with some re-leached cobalt units) are precipitated at pH 9 and recycled within the flowsheet.

Internal

Internal

Flow Diagram Cement Waste

effluent NajZnOj solution treatment

Figure 10.20. Schematic flow diagram of zinc cement treatment (Henry et al,2004)

effluent NajZnOj solution treatment

Figure 10.20. Schematic flow diagram of zinc cement treatment (Henry et al,2004)

simplifie!» process flowsheet

3 Mn2+ + Cr2Oj+ 2 H+ -> 3 Mn02 + 2 Cr3+ + H20 (10.23)

Zinc Cemcentation Flowsheet
03 + 2 Ni*4 + 4 OH" -> 2 NiOOH + 02 + H20 (10.25)

Zn2+ + 2 e —> Zn, at the catho Mnz+ -> MnJ" + e, at the anc 2 Mn3+ + 2 H20 Mn2+ + Mn02 + 4 IT

(OH)8(M&FC)4AUSi2O10 + 5Si02->(Mg.Fe)2AUSiJ01o+ ICM&FeJSiQ, + 4H20 (10.26)
3 Fe3+ + 2 SO42" + M* + 6H20MFe3(S04)2(0H)6 + 6lT (10.33)
4 Fe2+ + 02 + 6 HjO -» 4 FeOOH + 8 H+ (10.34)

2 NaFejCSO^fOH)« + 6 H2S04 3 Fe^SO«), + Na2S04 + 12 H20 3 Fe2(S04)2(0H)6 —> 3 Fe203 + 9 H2S04

2 NaFej(S04)2(0H)g 3 Fe203 + Na2S04 + 3 H20 + 3 H2S04 (10.35)

3 Fe2(S04)j + 6 H20-> 6 Fe(S04XOH) + 3 H2S04 (10.36)

3 Fe2(S04)j + 6 H20-> 6 Fe(S04XOH) + 3 H2S04 (10.36)

H2Te04 + 4 Cu + 3 H2S04 -> CuTe + 3 CuS04 + 4 H20 (10.53)

TeOj+ 3 H20 + 4 e Te + 6 OH" E° = - 0.247 V (10.62)
Se02(js) + 2 H20(D + 2 SOacg) Se w + 2 H2S04ro (10.63)

Recovery from Waste Sludges 429

separator to separate quartz, alumina and alkalis in nonmagnetic fraction. The final concentrate on spirals is dried, heated and screened (at 0.315 mm) and sent to electrostatic separator. This separates conductive minerals, ilmenite, magnetite and gold. The ncmconductive fraction consists of zircon, garnets and quartz. Further gravity separation of nonconductive fraction by table concentration leads to production of zircon concentrate with 52.6 % ZrOj. This is useful in ceramic industry. The components of the conductive fraction are sent to magnetic separator (H — 800-1,000 Oersted) in four stages to separate magnetite, ilmenite and gold concentrates. The final products are magnetite (20.2 % Ti02), ilmenite (45 % TiCy and a gold product (50-100 g/t). (For details of gravity separation techniques, see Chapter 3).

STEP ONE: STEP TWO; STEP THREE: STEP FOUR:

Contaminated soil is The coniairtnated sdt Is The vapors are cooled and The heated soil is then

STEP ONE: STEP TWO; STEP THREE: STEP FOUR:

Contaminated soil is The coniairtnated sdt Is The vapors are cooled and The heated soil is then

Figure 10.31. Process flow diagram to recover mercury from contaminated soil (Harris and Baum, 1996)
Figure 10.32. Flow sheet for recovering valuable minerals from overburden rock (Georgescu et al, 2001)

Si02 (s) + 2 (CH3)2CO (g) -> (CH30)4Si (g) + 2 C02 (g) (10.67) Si (s) + 2 (CH30)2C0 (g) ->(CH30)4Si (g) + 2 CO (g) (10.68)

2 Na + 2 H20 —» 2 NaOH + H2 2 Al+ 2 NaOH + 2 H20 2 NaA102 + 3 H2

ind HCI

FeClj strip hquor

NICIj CrClj strip solution liquor

FeClj strip hquor

NICIj CrClj strip solution liquor

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