5,078,786 Process for recovering metal values from jarosite solids

Patent Number/Link:

United States Patent [19]

Peters et al.

11111111111101111

USOO5078786A

[11] Patent Number: 5,078,786

[45] Date of Patent: Jan. 7, 1992

20 Claims, 5 Drawing Sheets

This invention concerns a process for recovering metal

values from jarosite-eontaining materials by leaching

with a calcium chloride solution at a temperature above

about the atmospheric boiling point of the solution and

under at least the autogeno~s pressure. .

OTHER PUBLICATIONS

"The Encyclopedia of Chemical Technology", Kirk-

Dthmer, vol. 24, pp. 812-824, 3rd Edition.

"The Jarosite Process-Past, Present & Future", V. Arregui

et al. Lead-Zinc-Tin, TMS-AIME World Symposium

on Metallurgy and Environmental Control,

1980, J. M. Cigon, T. S. Mackey and T. J. O'Keefe, Eds.

pp.97-123.

Primary Examiner-Peter D. Rosenberg

Attorney, Agent, or Firm-Sheridan, Ross & McIntosh

3,969,107 7/1976 Lippert et al. 75/101

4,054,637 10/1977 DreuUe et aI 423/39

4,070,437 1/1978 Van Ceulen 423/1

4,128,617 1211978 DeGuire et al. 423/106

4,182,852 3/1980 Pammenter et aI 423/42

4,244,735 1/1981 Reynolds et aI 75/101

4,266,972 5/1981 Redondo-Abad et aI 75/101

4,305,914 1211981 Pammenter et aI 423/146

4,342,592 8/1982 Lamb 75/114

4,355,005 10/1982 Rastas et aI 423/41

4,366,127 1211982 Rastas et al. 423/26

4,383,979 5/1983 Rastas et al. 423/36

4,410,496 10/1983 Smyres et al. 423/1

4,415,540 11/1983 Wilkomirsky et al. 423/99

4,505,744 3/1985 Weir et aI 75/120

[57] ABSTRACT

[54] PROCESS FOR RECOVERING METAL

VALUES FROM JAROSITE SOLIDS

[75] Inventors: Mark A. Peters, Northglenn; Wayne

W. Hazen, Wheat Ridge; James E.

Reynolds, Lakewood, all of Colo.

[73] Assignee: Resource Teclmology Associates,

Tulsa, Okla.

[21] Appl. No.: 438,502

[22] PCT Filed: Nov. 26,1986

[86] PCT No.: PCTIUS86/02476

§ 371 Date: May 26, 1989

§ 102(e) Date: May 26, 1989

[87] PCT Pub. No.: W088/03911

PCT Pub. Date: Jun. 2, 1988

[5 I ] Int. ct.S C22B 3/00

[52] U.S. Cl. 75/432; 75/726;

75/733; 75/736; 75/961

[58] Field of Search 75/432, 726, 733, 736,

75/961

[56] References Cited

U.S. PATENT DOCUMENTS

820.000 5/1906 Just.

915.705 3/1909 Seigle 423/22

3.143.486 8/1964 Pickering et a!. 204/119

3,493.365 211970 Pickering et a!. .

3,684.490 8/1972 Steintveit 75/101

3,691,038 9/1978 von Roepenack et al. 204/119

3,910,784 10/1975 Rastas 75/1

3,959,437 5/1976 Rastas et al. 423/36

SPENT

EL£CTROLYTE

CaO

14)

Pb", 2"

IoIETIoJ.

CaO

STRONC NH4OH

u.s. Patent Jan. 7, 1992 Sheet 1 of 5 5,078,786

~Pb/A9

RESIDUE

CALCINE

JAROSITE

RESIDUE

JAROSITE

PRECIPITATION

FIG.1

CALCINE

NEUTRAL LEACH .

5.0

FIG.2

'V---------~Pb

-----t......----------Zn

1.0 2.0 3.0 .4.0

POTASSIUM IN INITIAL LEACH SOLUTION

G/L K

O~--..L----I------'---.......II--..-.-..I o

100

~. z

0::

~>w<

U.S. Patent Jan. 7, 1992 Sheet 2 of 5 5,078,786

100

80 • 180° C,10 GIL K

,,,- • 95°C,

70 I

t( ,, o G/L K . , z 60 I

0 ,

F ,

~ 50

,,,

~ ,

w 40

t1' <-

30 10 GIL K

95°C,

10 5 G/L K

0 1 2 3 4 5

LEACH TIME, HOURS FIG.3

JAROSI~5

CaO-----.

CaCI2 LEACH

___ ~IO

: RECYCLE

! /9

SOLIDILiQUID I--_--'--~ LEACHATE

SEPARAnON LIQUID

FIG.4 SOUD-8

TAILS

u.s. Patent Jan. 7, 1992 Sheet 3 of 5 5,078,786

JAROSITE/""5

0 MAKEUP

,t ,~ t CaCI2

v-6

LS~ - CoC'2 LEACH -

1"'" 9

13 (15

L -- -.. CEMENTATION Ag CEMENT

Pb)

, 17

Pb REMOVAL r -.. 'Pb

t r 19

Zn SEPARATIONY - Zn RECOVERY

0 23r - I NHJ l' 25",- LIME BOIL

CoC'2

RECYCLE

FIG.5

{ 14

META

(Zn or

CaC03

CaD

FERRITE/ 31

H 2 804 -. r-----NH

,..--1_..&.---1.-_ 35

FERRITE t--~ Zn RECOVERY

LEACH

u.s. Patent Jan. 7, 1992 Sheet 4 of 5 5,078,786

71"\ 49) ,I •

SPENT

: Zn PLANT ~~

ELECTROLYTE •

31 FERRITE I

33\ ~ :35

It .. L1aUID1.. \ I LEACH: ~

41"- JAROSITE

43 , SOUD

JARO(ITE .. I REPULP TANK 45

MAKEUP

47...... CaCl2

6, •• I .~

CaD ~ CaCI2 LEACH --

•

14) ~Ir /"13 r l5 Pb or Zn ..~! CEMENTATION! .. Ag CEMENT

METAL

'r /51 (53

H2S - I SULFIDING I • - PbS CAKE /' It 59.......

., r

CaC03 _IZ • I Z STRIP~ ..

I n EXTRACTION. \ - l n -

56 57)

'25\ '

CoCIZ RECYCLE

CoO :--: LIME BOIq ..

f1G.7 , STRONG NH40H -

u.s. Patent Jan. 7, 1992 Sheet 5 of 5 5,078,786

A9/Cu~EMENT .

MAKEUP 59

CeCI 2

WASH

ZnO

or Pb 0

31 SPENT

~ ELECTROLYTE

~ ...--=------.w.----l-----..:........t

Pb(OH)2 87

cea2 RECYCLt'

Zn STRIPPING I--;----J

PARTIAL NEUlRALlZATION 85

CeC03

CeO

FIG. 8

5,078,786

conditions allow extraction of up to 95 percent of the

lead and silver content of the waste. However, Applicants

have found that unpredictably with some jarosite

wastes this process provides recoveries of less than 20

5 percent of the silver present.

U.S. Pat. No. 4,054,638 of Dreulle et al. (1977) is

directed to a process for recovering metals from sulfated

residues from electrolytic zinc plants. The residue

is digested preferably at a temperature between 95· and

10 lIS· C. with hydrochloric acid in the presence of calcium

chloride. This leaching process dissolves the metals

present, including the iron, by forming the corresponding

metal chlorides. Consequently, the process

requires that the iron chloride be removed by extraction

by an organic solvent. This process has a disadvantage

of solubilizing the iron and requiring a separate separa~

tion step. There is no suggestion or disclosure of using

superatmospheric pressure for this leach.

U.S. Pat. No. 4,070,437 of Van Ceulen (1978), discloses

a process for recovery of metals from jarosite

sludges. The process involves leaching the jarosite with

an acidic calcium chloride solution, preferably formed

by mixing hydrochloric acid and calcium hydroxide or

calcium carbonate. The leaching is preferably carried

out close to the boiling point of the leaching medium.

Insoluble calcium sulfate is formed and is separated by

filtration. This process has the disadvantage ofsolubilizing

essentially all of the iron in the jarosite.

Another waste which contains metal values is zinc

ferrite-containing materials. Modern electrolytic zinc

processes commonly use a two-step leaching process as

depicted in FIG. 1. The second leaching step involves a

hot acid leach to dissolve zinc ferrite present. However,

prior to the development of the two-step leach, a single

neutral leach was used which caused much ofthe zinc

ferrite and associated metal values to be discarded as

wastes. Therefore, there are existing waste lagoons

which contain substantial quantities of zinc ferrite and

other metal values. The term "ferrite" is used herein to

refer to a combined metal oxide-ferric oxide ml,lterial,

e.g. zinc ferrite (ZnO.Fe203).

A number of processes have been developed for the

purpose of recovering this zinc. One such process is

disclosed by Rastas et al. in U.S. Pat. No. 3,959,437

(1976). Rastas et al. disclose a process in which the

ferrite of a non-ferrous metal, as well as the oxide of the

non-ferrous metal, is subjected to a neutral leach which

dissolves most of the oxide but leaves the ferrite substantially

unaffected. The non-ferrous values in the

solution are recovered and the undissolved ferrite material

is further treated in a "conversion" stage with sulfuric

acid-bearing solution at atmospheric pressure and at

a temperature of about 80· C. to about lOS· C. in the

presence of alkali or ammonium ions. Under these conditions,

the non-ferrous metals dissolve as sulfates,

while iron is simultaneously precipitated as an insoluble

complex sulfate, i.e., jarosite. U.S. Pat. No. 4,355,005 of

Rastas et al. (1982), U.S. Pat. No. 4,366,127 ofRastas et

al. (1982), as well as U.S. Pat. No. 4,383,979 ofRastas et

al. (1983) each disclose modifications to the process

disclosed in the '437 patent.

Steintveit in U.S. Pat. No. 3,684,490 (1972) discloses

a method for treating jarosite residue in which the residue

is subjected to leaching with sulfuric acid at a temperature

of SO· to 95· C. and an acid concentration of 10

to 70 grams per liter (hereinafter gil). These leaching

FIELD OF THE INVENTION

PROCESS FOR RECOVERING METAL VALUES

FROM JAROSITE SOLIDS

This invention relates to a process for recovering

metal values from the jarosite containing wastes from

electrolytic zinc recovery plants.

BACKGROUND OF THE INVENTION

Electrolytic zinc processes are used to treat complex

zinc-containing ores that cannot readily be treated by

pyrometallurgical recovery. The usual steps in such an

electrolytic process include: (a) concentrating the zinc

ores; (b) roasting the zinc concentrate to eliminate sui- 15

fur and produce zinc calcine; (c) leaching the zinc calcine

to provic!e an impure zinc sulfate solution; (d) separating

the iron present usually by forming a jarosite

precipitate; (e) purifying the zinc sulfate solution; and

(f) subjecting the zinc sulfate solution to electrolysis to 20

recover the zinc metal. Such a process is described in

U.S. Pat. No.4, 128,617 (1978) of DeGuire et al. which

is incorporated herein by reference. A simplified process

flow sheet for an electrolytic zinc plant is shown in

FIG. 1. 25

Additional details of various process modifications

can be found in "The Encyclopedia of Chemical Technology",

Kirk-Othmer, Vol. 24, pp. 812-824, 3rd Ed,

incorporated herein by reference.

As indicated above, a common method of removing 30

the iron present in the leachate is through the formation

of a "jarosite" precipitate. Jarosite, MFe3(S04h(OH)6

where M is a monovalent ion, usually an alkali metal

(general sodium or potassium) or ammonium, is commonly

formed by adding a source of ammonium or 35

sodium ions to the leach solution and maintaining the

solution at an appropriate pH by the addition of base.

This process is also shown in FIG. 1. Various modifications

to the so-called "jarosite process" are discussed in

the article entitled "The Jarosite Process-Past, Present 40

and Future", V. Arregui et aI., Lead-Zinc-Tin, TMSAIME

World Symposium on Metallurgy and Environmental

Control, 1980, J. M. Cigon, T. S. Mackey and T.

J. O'Keefe, Ed., pp. 97-123, incorporated herein by

reference. 45

There are a number of problems associated with the

formation of jarosite waste material. The jarosite can

contain valuable metals such as silver, zinc, copper,

lead, indium, etc. which require numerous expensive

process steps to recover. The jarosite can also contain 50

toxic species which can be leached into the environment

by rain and groundwater. Therefore to avoid

environmental contamination it is usually ·necessary to

store the jarosite wastes in sealed lagoons which are

expensive to build. 55

A number of methods for treating such wastes have

been disclosed. Steintveit et al. in Norwegian Patent

142,406 (1980) disclose a process for leaching iron-con!

aining waste with chloride-containing, acidic solution

at a temperature between SO· C. and the boiling point of 60

the solution. An alkali metal chloride or an alkaline

earth metal chloride is used as the source of the chloride

with calcium chloride being disclosed as the preferred

material. The pH of the solution is adjusted during the

leaching step so that the iron remains as a precipitate 65

while the valuable metals are leached into the hot solution.

The pH is adjusted to between about 2 and 4 preferably

with calcium hydroxide. It is disclosed that these

5,078,786

conditions are intended to decompose any zinc ferrites

present and provide for a greater recovery of the zinc.

U.S. Pat. No. 3,691,038 of Von Roepenack et al.

(1972) discloses a method for recovering zinc from

oxides containing zinc and iron. The oxide is leached 5

with sulfuric acid at a temperature of95· to 100· C. with

an excess of sulfuric acid to solubilize the zinc and iron.

Alkali metal or ammonium ions are added to the liquid

phase along with a zinc-containing oxidic material at a

temperature of 95· to 100· C. to precipitate jarosite.

U.S. Pat. No. 4,192,852 of Pammenter et al. (1980)

discloses a process for treating zinc plant residues containing

zinc ferrite and precipitating the iron as a jarosite.

The sulfate solution containing ferric iron, free acid

and non-ferrous metals is cooled, partially neutralized 15

and then heated to a temperature not exceeding the

boiling point at atmospheric pressure in the presence of

sodium, potassium or ammonium ions. U.S. Pat. No.

4,305,914 of Pammenter et al. (1981) discloses a process

similar to that in the '852 patent. 20

U.S. Pat. No. 4,128,617, of DeGuire et al. (1978),

describes a three-step process for the treatment of zinc

calcine containing zinc oxide, zinc sulfates, and zinc

ferrites. The first step involves the neutral leaching of

the zinc calcine with an effective amount of aqueous 25

sulfuric acid containing solution. The leach residue is

subjected to hot acid leaching with sulfuric acid followed

by jarosite precipitation by alkali with the subsequent

recycling of the jarosite-containing pulp. The

preferred temperature range for the hot acid leaching is 30

from about 80· C. to the boiling point and preferably the

temperature is greater than 90· C. There is no suggestion

of the use of pressure.

The processes described hereinabove have one or

more of the disadvantages of (1) having low rates of 35

extraction of metal values, (2) providing low levels of

recovery of certain metal values, (3) having poor filterability

of the iron-containing residue, and/or (4) solubilizing

large amounts of iron.

Several processes disclosed in the art have used ele- 40

vated temperature and pressure leaching steps in treating

zinc plant residues.

U.S. Pat. No. 3,143,486 of Pickering et al. (1964)

discloses a process for the extraction of zinc from zinc

ferrite containing residue. The process involves subject- 45

ing the residue to a first-stage leaching treatment under

non-oxidizing conditions in a closed vessel in the presence

of excess sulfuric acid at a temperature between

140· C. and 260· C. Zinc is dissolved as well as ferrous

sulfate which is stable at the temperatures and acidities 50

used. Ferric iron is precipitated as a basic sulfate. The

leachate is then subjected to a second-stage leaching

treatment at 140· C. to 260· C. under oxidizing conditions

to oxidize the ferrous sulfate to ferric and precipitate

the ferric material. Similarly, U.S. Pat. No. 55

3,493,365 of-Pickering et al. (1970) discloses a two-step

high temperature method of treating zinc plant residue

containing zinc ferrite. This process differs from that of

the '486 patent in that in the second step a source of a

cation selected from the group consisting of sodium, 60

potassium and ammonium is added in order to precipitate

the iron from the liquor as a jarosite material.

A process for treating sulfide ores which involves a

two-step leach is disclosed by U.S. Pat. No. 4,266,972 of

Redodno-Abad et aI., (1981). The first leach uses sulfu- 65

ric acid under an oxygen atmosphere at 150· to 250· C.

Zinc and copper are solubilized with lead, the noble

metals, and iron oxide remaining as a residue. After a

solid liquid separation, the filtrate is adjusted to a pH of

1.5 to 2. Sodium chloride, calcium chloride and ferric

chloride are added to precipitate calcium sulfate. The

leach is conducted at a temperature between 60· C. and

90· C. with the silver, lead, and gold being solubilized as

the chlorides. The iron oxide remains as a residue. After

a solidlliquid separation, the silver, lead, and gold are

recovered by cementation with zinc, with the liquid

being subjected to an extraction to recover the zinc.

None of these processes which use high temperature

and pressurized leaches discloses or suggests that jarosite-

containing wastes can be advantageously treated in

such a manner. In fact, most disclose the use of an oxidizing

atmosphere to form ferric iron which will precipitate.

These patents also disclose that potassium ions can

be added to a zinc ferrite leach solution in order to

precipitate potassium jarosite. As discussed in detail

hereinbelow, it has been found that the recovery of

metal values can be unpredictably affected by the presence

of potassium ions.

Accordingly, there is a need for a process to treat

jarosite and ferrite containing wastes from zinc recovery

processes in order to recover metal values which

are contained in the waste materials and render the

residue suitable for disposal as a nonhazardous waste.

There is also the need for a process which will not be

subject to the unpredictable effect of potassium ions.

SUMMARY OF THE INVENTION

It has now been found that the above described disadvantages

of known processes can be minimized or eliminated

by the instant invention. According to the present

invention, a process is provided for recovering metal

values from jarosite wastes from a zinc recovery plant

said process comprising leaching the waste with an

acidic solution of calcium chloride at or above the atmospheric

boiling point of the solution and under at

least the atmospheric pressure.

One of the embodiments of the instant invention comprises

a process for recovering metal values from jarosite-

containing wastes from a zinc plant using an acidic

solution of calcium chloride at a temperature above the

boiling point of the solution and under super-atmospheric

pressure.

Another embodiment of the instant invention comprises

a process for recovering metal values from a

jarosite-containing waste wherein the process comprises

contacting the waste at a temperature of between

about 110· C. and 300· C. and a pressure of at least the

auto genous pressure with a solution containing between

1.0 and 5.0 molar calcium chloride. The contacting

solution has a pH of between about 2.0 and 3.5. This

pH is maintained by the addition of a calcium compound

selected from the group consisting of calcium

oxide, Calcium hydroxide, calcium carbonate and mixtures

thereof.

In a further embodiment, the instant invention comprises

a process in which jarosite-containing wastes are

leached with an acidic solution of calcium chloride at

greater than atmospheric pressure and a temperature

greater than the atmospheric boiling point of the solution

to form a liquid leachate and a solid residue. The

liquid leachate is separated from the solid residue and

contacted with a reducing metal to reduce silver cations

contained in the leachate to metallic silver. The metallic

silver is then separated from the liquid solution. The

liquid solution is sulfided by mixing the solution with a

sulfide compound to precipitate the lead contained in

5,078,786

The overall leach process, including regeneration of

eaCh, may be represented by combining the above

reaction with the reaction occurring during the· "lime

boil" operation,

(I)

2NH4Fe3(SO.h(01l)6+CaCI2+3Ca(01l)2-+

6Ft(.OHh+2NH.C/+4CaSO.

some jarosite wastes the presence of potassium ions can

result in significantly lower recoveries of metal vaJues

such as silver and lead from the leach solution. FIG. 2

is a graphic presentation of experimental data showing

the effect of potassium on the amount of metal extracted.

This effect is particularly evident at lower

leaching temperatures. Thus, using higher leaching

temperatures than disclosed in the prior art appears to

unexpectedly compensate for the presence of potassium

in the leaching solution. Graphic evidence of this previously

unrecognized phenomenon is provided in FIG. 3

where the effect of increasing the extraction temperature

from 130° C. to 180' C. is shown. The results also

show that only 10 percent of the silver was extracted at

95° C. when 5 grams per liter ofpotassium were present.

With 10 grams per liter of potassium, 80 percent of the

silver was extracted at 180' C.

The detrimental effect of ionic potassium is not observed

with every jarosite material. Differences in the

effect of potassium concentration on the recovery of

silver have even been observed on samples of jarosite

obtained from different locations in the same waste

lagoon. Although extensive work has been done to

determine the basis of this phenomenon, the reason

remains unknown. A method for predicting the magnitude-

ofthis effect has not been discovered as yet. Therefore,

practicing the present invention assures that the

recovery of metal values can be maximized even with

variations in the magnitude of the potassium effect.

The leaching process for ammonium jarosite can be

represented by the following reaction scheme.

BRIEF DESCRIPTION OF THE DRAWING

the solution as lead sulfide. The solid lead sulfide is

separated from the liquid phase. Substantially all of the

zinc in the remaining liquid phase is recovered using a

zinc recovery process to provide a liquid solution substantially

free of zinc. The resulting liquid solution is 5

heated with calcium oxide to provide a vapor containing

ammonium hydroxide and bottoms which contain

calcium chloride.

In another embodiment, the instant invention comprises

contacting a ferrite which contains metal values 10

with a leach solution of sulfuric acid and ammonium

ions. The leach solution containing the ferrite is heated

at a temperature above about 90° C. for up to about 12

hours to form a solid containing ammonium jarosite and

a liquid phase. The solid phase is mixed with an acidic IS

leaching solution containing calcium chloride. This

mixture is heated at superatmospheric pressure and at a

temperature above the atmospheric boiling point of the

solution to solubilize a substantial portion of the metal

values. 20

FIG. 1 shows a typical process flowsheet for an electrolytic

zinc plant;

FIG. 2 shows the effect of potassium on metals ex- 25

traction;

FIG. 3 shows the effect of temperature and potassium

concentration on silver extraction;

FIG. 4 shows the flowsheet for a high temperature

calcium chloride leach; 30

FIG. 5 shows a process flowsheet for the recovery of

metal values from jarosite;

FIG. 6 shows a process flowsheet for a ferrite leach

followed by a high temperature calcium chloride leach;

FIG. 7 shows a process flowsheet for metals recovery 35

from ferrite and jarosite in association with a zinc plant;

and

FIG. 8 shows a process flowsheet of a preferred embodiment.

DESCRIPTION OF THE PREFERRED

EMBODIMENTS

The process of this invention comprises a method for

recovering metal values from jarosite containing wastes

from electrolytic metal recovery plants. The jarosite 45

waste is leached at or above the boiling point of the

leach solution under pressure with an acidic calcium

chloride solution. Optionally, non-jarosite wastes such

as zinc ferrite can be subjected to a preleach step in

which the iron is converted to jarosite material and the SO

zinc is substantially solubilized.

The use of a high temperature, pressurized leaching

process for leaching jarosite has been found to provide

several unexpected advantages over the lower temperature,

atmospheric pressure processes described in the 55

prior art. It has been found that the final leach slurry

unexpectedly filters two to three times faster than comparable

slurries obtained with a 95' C. to 100' C. leach

such as that described by Steintveit in Norwegian Patent

No. 142,406 (Supra). Additionally, the leach times 60

are significantly less than those required when leach

temperatures below 100' C. are used. This procedure

has also been found to provide higher extractions of

metal values from certain types of wastes than obtained

with the lower temperature leaches. 65

It has also been found that certain materials unexpectedly

interfere with the recovery of metal values from

jarosite. In particular, it has been observed that with

(2)

to get the following overall reaction,

2I1'H4Fe3(SO.h(OH)6+ 4Ca(OH)2--6Fe(OH)3+-

4CaSO.+2NH3+2H20. (3)

The calcium chloride leaching process of the instant

invention is normally conducted at or in excess of the

atmospheric boiling point of the leach solution. As used

herein, the terms "atmospheric boiling point" and "boiling

point" are used interchangeably to refer to the temperature

at which the solution boils under the particular

atmospheric pressure to which it is being subjected.

Thus the boiling point for a solution of the same composition

can vary depending upon the geographical elevation

at which it is heated. Ordinarily, the boiling point

of the leach solution is in excess of 110° C. It has been

found that leaching at a temperature of at least the boiling

point of the solution and preferably above 110' C.

provides improved results in extracting metal values

from the jarosite-containing waste materials. The boiling

point of the leaching solution can be increased by

increasing the concentration of calcium chloride in the

solution. Therefore, it is contemplated that the instant

invention encompasses the embodiment of heating a

solution at or above I 10° C. even though the boiling

point of the solution is in excess of 110° C. Leaching

temperatures from the boiling point to about 300' C. can

5,078,786

7 8

be used with temperatures ofbetween about 120· C. and material called jarosite has the formula MFe3(S04)-

225· C. being preferred and temperatures of between 2(OH)6. There are a variety of jarosite type compounds

about 130· C. and 200· C. being most preferred. The which contain different ions in place of the M, for exampressure

on the leaching solution has not been found to pIe, potassium, lead, silver, sodium, rubidium or ammobe

critical as long as the pressure is sufficient to prevent, 5 nium. Ammonium jarosite is a principal effiuent from an

the solution from boiling. Ordinarily, autogenous pres- electrolytic zinc process.

sure is maintained in the system, i.e. greater than 1atmo- Other iron containing materials can be used as feeds

sphere and preferably at least about 1.1 atmospheres. As to produce jarosite e.g. zinc ferrite (ZnFe204). The zinc

used herein, the term "autogenous pressure" means that ferrite used can be any such material containing metal

pressure which develops in a closed system when the IO values. The source of zinc ferrite is ordinarily waste

system is heated. lagoons from zinc plants.

The concentration of calcium chloride should be An alternative source of zinc ferrite material is the

maintained above about 0.5 molar in the leaching solu- dust formed during steelmaking in an electric arc furtion

to achieve the best extraction and solubilization of nace (EAF). The dust primarily comprises zinc ferrite,

metal values. The concentration can range from about 15 zinc oxide, and various forms of ferric oxide. The dust

0.5 molar up to the saturation point for calcium chloride can also contain lead, cadmium and chromium and is

for the particular temperature and solution. It is pre- therefore usually classified as a hazardous waste. The

ferred that the concentration of calcium chloride be dust can additionally contain metal values such as silbetween

about 1 molar and about 4 molar in the leach- ver. Thus, the dust can be fed into the process of the

ing solution with the most preferred concentration of 20 instant invention to allow recovery of metal values

calcium chloride between about 2 and 3.5 molar. and/or removal of the toxic metals.

Calcium chloride can be added to the leaching solu- In order to be used in the present process, the EAF

tion directly or can be formed in situ by using a chloride ferrite is subjected to a leach operation in which the

source and a calcium source. For example, materials ferrite is contacted with sulfuric acid and an ammonium

such as hydrogen chloride, sodium chloride and ammo- 25 source to precipitate ammoniumjarcsite which contains

nium chloride can be used as the chloride source. Useful some of the metal values. The ammonium jarosite can

sources of calcium include lime, hydrated lime, and then be conveyed to the calcium chloride leaching procalcium

carbonate. The particular materials used de- cess.

pends upon the economics. The use of the calcium oxide In the practice of the instant leaching process, the

or calcium hydroxide as the calcium source has the 30 jarosite 5 is added to the leaching vessel 6 along with

additional advantage that the pH of the solution can be calcium chloride as represented in FIG. 4. The calcium

adjusted to the desired range by their addition. To avoid chloride can be makeup or be recycled from downthe

buildup of certain cations, e.g. sodium, in a continu- stream process steps or a mixture of the two sources.

ous system, it may be necessary to bleed a small stream During the high temperature leaching process, the pH

out of the recycle or to remove the cation by other 35 of the solution tends to decrease due to generation of

means such as ion exchange. acid. The pH is adjusted to the preferred leaching range

For effective leaching with calcium chloride, it is by adding a basic material, preferably calcium oxide,

important that the pH of the solution be less than about calcium carbonate, and/or calcium hydroxide. The

5. It is preferred that the pH be less than about 4 and solution is heated to the leaching temperature, preferamost

preferred that the pH be less than about 3. For best 40 bly by steam, with agitation of the mixture. After the

results, the extraction should be carried in a pH range of appropriate leaching time, the resulting slurry is subabout

2.0 to 3.0 During the calcium chloride leach of jected to a liquid-solid separation 7 to provide solid tails

the jarosite, the pH of the solution can decrease to a pH 8 and a liquid leachate 9. The solid tails contain iron

of less than 1.0 unless adjusted by the addition of a base. oxides, gypsum, and/or unreacted gangue while the

Although any base, such as sodium hydroxide or so- 45 liquid leachate contains solubilized zinc and other metal

dium carbonate can be used, it is advantageous as dis- values. The tails are discarded while the leachate is

cussed hereinabove to use a calcium containing base subjected to further downstream treatment as discussed

such as calcium oxide, calcium hydroxide or calcium in more detail hereinbelow. Optionally, a portion of the

carbonate. leachate is recycled 10 to the calcium chloride leach

Other things being equal, the higher the leaching 50 step to effect a build-up of metals in the pregnant solutemperature,

the shorter the leaching time required to tion.

achieve a particular level of extraction of metal values. The leachate from the calcium chloride leaching

The time required for the extraction depends upon the process can be treated by any suitable method to reconcentration

of calcium chloride in the leach solution, move the metal values such as copper, silver, or gold

the temperature at which the leaching is conducted, the 5S which might be present. When silver and copper are

pH of the solution and the particular waste feed being present, it is preferred that a metal cementation process

treated. Ordinarily, the jarosite leaching requires less be used to remove the silver and/or copper. As reprethan

about 6 hours with leach times in the range of sented in FIG. 5, this can be accomplished by treating

about 5 minutes up to about 2 hours being preferred. At the leachate with any material which is a suitable reducthe

higher leaching temperatures, it is expected that in 60 tant for the metal values. Metals which have higher

excess of about 75 percent and preferably in excess of 85 reduction potentials than silver, such as lead or zinc, can

percent of the silver and lead present in the jarosite is be used to form a silver cement. Presently for economic

solubilized during the preferred leaching time. reasons, it is preferred to use zinc dust. Zinc dust can be

. The jarosite-containing waste materials suitable for added to precipitate a solid consisting of a mixture of

use as feed to the calcium chloride leaching process 65 silver, lead, copper (depending upon the presence of the

comprise materials which contain the metal values such particular metal in the solution) and zinc. The solid

as zinc, silver, lead, indium. The preferred feeds are product which is separated from the liquid phase can

jarosite residues from an electrolytic zinc process. The then be subjected to additional standard metallurgical

5,078,786

build-up of undesired materials such as potassium ions in

the system if recycle is used. Preferably a "lime boil"

operation can be used in which lime is added to the

raffinate to increase the pH to a basic range, preferably

above about 9, and the mixture heated to form ammonia

which can be removed by volatilization. Preferably, a

temperature of at least the boiling point of the solution

is used. The ammonia formed is removed, usually in the

vapor, and can be used to form jarosite in the ferrite

treating process described hereinbelow or can be transferred

to the zinc plant. The brine solution from the

lime boil process contains calcium chloride as well as

residual metal values. This brine solution can be recycled

for use in the calcium chloride leaching process.

An alternative method for recovering the zinc is to

treat the liquor from the sulfiding procedure with a base

such as calcium oxide to precipitate zinc hydroxide. In

this operation (not shown in FIG. 5), sufficient base is

added to provide a solution pH of between about 8 and

10. Bases which are useful in this process include calcium

hydroxide, calcium oxide, sodium hydroxide, etc.;

provided there is no detrimental build-up of cations in

the system. The zinc hydroxide precipitate is separated

from the liquor by standard liquid-solid separation techniques

such as fJltration or centrifugation. The zinc

hydroxide solid is preferably transferred to a zinc plant

for recovery of the zinc. The liquor from this precipitation

process is transferred to a lime boil process as described

hereinabove.

Depending upon the level oflead and/or zinc present

in the liquor from the silver cementation process, the

basic precipitation step can be optionally performed on

the liquor from the silver cementation step. This will .

result in a mixture oflead hydroxide and zinc hydroxide

35 solids being formed. After a solidlliquid separation,

these solids can be disposec! of in any manner appropriate.

The liquor from this precipitation step is transferred

to a lime bOil process as described hereinabove.

When zinc ferrite (ZnFe204) is a significant component

of the waste, a sulfuric acid (sulfuric) leach step is

used. In this leach there is a simultaneous leach of the

ferrite and a precipitation of the iron in the form of

jarosite. A method for this simultaneous leach and precipitation

is described by Rastas et al. in U.S. Pat. No.

3,959,437 (1976). Rastas, however, is only concerned

with solubilizing the zinc in the ferrite and does not

treat the jarosite to recover metal values.

In the present invention, the ferrite material is combined

with sulfuric acid and an ammonium source at a

temperature above about 50· C. and below the decomposition

temperature of ammonium jarosite. Preferably,

the temperature is between about 80· C. and about 170·

C. The sulfuric acid used is preferably spent electrolyte

from an electrolytic zinc plant and is in the ferrite leach

solution to the extent of between about 10 and about 60

grams per liter, preferably between about 30 and about

50 gil. The ammonium compound is preferably recovered

from the lime boil process step described hereinabove.

This leach is ordinarily conducted at a temperature

of about 80· C. to 100· C. However, as discussed

hereinabove, the leach can be accomplished at a temperature

above the boiling point of the solution under

pressure. It is contemplated that the leach is conducted

under autogenous pressure at the elevated leaching

temperatures.

The ferrite leach is conducted for a period sufficient

to precipitate jarosite and solubilize zinc. Preferably the

leach time is about 12 to 36 hours. Commonly, at least

processes to recover the pure metal values. The separation

of the solid product from the liquid phase can be

accomplished by standard solid-liquid separation techniques

such as filtration or centrifugation.

The liquid from the metal cementation process con- 5

tains dissolved reducing metal, calcium chloride, lead,

zinc and other noncementable impurities. The liquid

can then be subjected to additional processing steps to

separate the reducing metal and other metals such as

lead and zinc (if not used as the reducing metal) from 10

the liquid. Optionally, the liquid can be subjected to a

sulfide precipitation procedure to recover the lead and

any trace of heavy metals which precipitate as sulfides

under these conditions. The sulfide precipitation can be

accomplished by adding a sulfide-containing material 15

such as hydrogen sulfide to the liquid to form solid lead

sulfide. The liquid can then be separated from the solid

lead sulfide cake by conventional solid-liquid separation

techniques. Alternatively, the pH of the liquid can be

adjusted to precipitate metals such as lead and zinc as 20

their hydroxides. For example, calcium oxide can be

added to precipitate lead hydroxide and zinc hydroxide.

In another embodiment, the liquid from the metal cementation

process is subjected to an extraction process

to separate zinc as described herein below. The raffinate 25

can be partially neutralized to a pH of about 8-9 to form

a lead hydroxide precipitate which is separated from the

liquid. The liquid can then be subjected to a lime boil as

discussed below. The particular choice of procedures to

be used to treat the liquid will depend upon the metals 30

present, their concentrations and the economics of the

various alternatives. The solids recovered can be· disposed

of in an appropriate manner or can be subjected

to additional processing steps to recover any metal

values present.

The liquid from the lead removal process can be

subjected to zinc recovery procedures. Any standard

method of recovery suitable for removing zinc from

such a solution can be used. A preferred method is

solvent extraction in which the aqueous solution con- 40

taining zinc is contacted with an extractant which transfers

the zinc from the aqueous solution to the organic

phase. The second phase can be a solvent which is not

miscible with the aqueous solution or it can comprise

the extractant. Alternatively, the liquor containing the 45

zinc can be passed through an ion exchange resin to

remove the zinc. A preferred method involves the use

of extractants such as di-2-ethylhexyl phosphoric acid

dissolved in a hydrocarbon solvent such as kerosene. In

the use of such materials, the pH is adjusted to deter- 50

mine the phase in which the zinc ultimately resides. In

a preferred method of operation, the zinc species is

selectively extracted into the organic phase by maintaining

the pH in an appropriate range depending on the

extractant used. The pH can be adjusted with a base 55

such as calcium oxide, calcium hydroxide, calcium carbonate,

etc. Zinc is then stripped from the loaded organic

with a strong acid solution such as spent electrolyte

from a zinc plant. Strip liquor containing the zinc

can be transferred to the zinc plant for electrowinning. 60

The raffinate from which the zinc has been removed

can be subjected to additional processing to recover the

calcium chloride and ammonia. A base can be added to

increase the pH of the solution to a basic range, preferably

above about 9. Although bases such as NaOH or 6S

KOH can be used, it is preferred that a basic calcium

compound such as calcium oxide, calcium hydroxide,

calcium carbonate or mixtures thereof be used to avoid

5,078,786

11 12

about 70 percent of the zinc present in the ferrite is carded. Optionally, a slip stream from the leachate is

solublized with less than five percent of the iron and recycled 10 to the calcium chloride leaching process to

essentially no silver or lead solubilized. The solid jaro- allow further concentration of metal values. The resite

containing metal values and any remaining un- maining leachate liquid is subjected to further processtreated

ferrite is separated from the liquid phase by 5 ing for recovery of metal values.

standard solid-liquid separation techniques· such as The recovery of the metal values can be accomthickening,

filtration or centrifugation. The liquor from plished by any known method useful for separating such

the ferrite leach step which contains zinc sulfate is pref- materials. A particular separation scheme is set forth in

erably conveyed to a zinc plant for recovery ofthe zinc. FIG. 5. The leachate 9 from the calcium chloride leach

The solid jarosite is subjected to the calcium chloride 10 6 is contacted 13 with a metal 14 capable of reducing

leach process described hereinabove. silver. Preferably zinc metal is used although lead is also

Referring now to FIG. 4, a process involving a cal- suitable and the choice is generally a matter of economcium

chloride leach ofjarosite in a system at a tempera- ics. The metal, ordinarily in the form of fine particuture

above the boiling point of the solution is repre- lates, is contacted with the leachate preferably at a

sented. Preferred embodiments cf this process are here- 1S temperature between about 40· C. and about 80· C. The

inafter described. Sufficient jarosite 5 is added to the silver cement 15 is separated 17 from the liquid by filtraleaching

vessel 6 to provide a slurry containing about 10 tion although centrifugation could also be used. The

to about 40 weight percent jarosite. A chloride source silver is recovered from the silver cement by convensuch

as calcium chloride is added to the slurry. Alterna- tional means such as smelting.

tively, other chloride sources can be used depending 20 As shown in FIG. 7, lead can be separated from the

upon the economics of the process. The pH is main- liquor from the cementation process by sulfiding 51

tained in the desired range of about 1.5 to about 3.5 by with hydrogen sulfide or any appropriate sulfide source

the addition of lime, i.e. calcium oxide, and/or calcium to precipitate lead sulfide 53. Ordinarily, greater than

hydroxide, calcium carbonate. The slurry is heated to a about 95 weight percent ofthe lead is precipitated. The

temperature ordinarily above the boiling point of the 25 lead sulfide is separated from the liquid by ftItration

solution, normally above 110· C. and preferably above although centrifugation can also be used.

about 120· C. The leaching zone or leaching vessel must The liquid from the sulfiding step is treated 19 to

be capable of withstanding the equilibrium pressure at recover zinc. Solvent extraction and precipitation are

the temperature selected. While the vessel can be pres- preferred methods of separation although any method

surized with an inert gas, in ordinary operation it is 30 suitable for recovering zinc 21 from such a stream can

maintained under autogenous pressure, i.e. the pressure be used. The zinc-solvent extraction 55 process, deestablished

in the vessel by the vaporization of volatile picted in FIG. 7 and FIG. 8, involves contacting the

components such as water at the leaching temperature. aqueous solution containing the zinc with an extractant

Ordinarily, the pressure in the leaching zone will be which will bind the zinc and allow its extraction into an

between about 5 to about 210 psig with the preferred 35 organic phase. A preferred extractant is di-2-ethylhexyl

range being from about 15 to about 130 psig. phosphoric acid dissolved in an appropri.ate diluent. In

The leaching operation can be carried out in either a operation, the extractant is mixed with the zinc feed

batch or a continuous mode. The particular choice of solution. Since the extraction ofzinc lowers the aqueous

operation will depend upon the leaching time necessary pH, a base such as NaOH, Cao, or CaC03 can be added

to extract the desired amount of the metal values. In a 40 to maintain the desired pH. At a pH of between about 2

batch operation, the necessary quantity of calcium ox- and 3 the zinc is selectively extracted into an organic

ide, calcium carbonate and/or calcium hydroxide can phase 56 in contact with the aqueous solution. This

be introduced initially instead of being added incremen- organic phase containing the zinc is separated from the

tally during the leaching process. The appropriate aqueous phase and the zinc is then stripped 57 into a

amount of base can be readily calculated by one skilled 45 strong sulfuric acid solution. The solution 59 containing

in the art based upon the jarosite content of the feed the zinc can be transported to a zinc plant for recovery

material added to the leach. The initial quantitative of the zinc.

addition of the calcium oxide is operationally advanta- In another process sequence (shown as a part of the

geous since it eliminates the equipment required to add scheme in FIG. 8), the liquid from the cementation step

the material incrementally throughout the leaching 50 is subjected to solvent extraction to remove the zinc as

process as well as the need to monitor the pH of the described hereinabove. The raffinate liquor from the

system. We have found that there is no difference in the zinc solvent extraction is partially neutralized, preferafmal

level of the metal values extracted when the cal- bly with calcium oxide, to a pH of between about 8 and

cium oxide is added initially as opposed to incremen- 9 to precipitate lead hydroxide. The lead hydroxide

tally during the leaching operation as long as the fmal 55 solid is removed, for example by filtration 65 or centrifpH

of the leaching solution is within the desired range. ugation and disposed of or the lead can be recovered by

At the conclusion of the leaching process, the liquid- known procedures.

solid slurry is separated 7 into solid tails 8 and a liquid The liquid 23 recovered from this separation is subleachate

9. This can be accomplished by any known jected to a lime boil 25 by adding lime to provide a pH

method for solid-liquid separations, although with this 60 above about 9 and heating to about the boiling point of

type of slurry, it is preferred that the separation be the mixture. Ammonia is vaporized and passed through

accomplished by filtration accompanied by appropriate a condensor to provide an ammonium hydroxide stream

washings. Surprisingly, it has been found that slurries which can be recycled to the ferrite leach or to the zinc

from the high temperature leaches filter two to three plant to produce ammonium jarosite. The liquid contimes

faster than the slurries from a leach conducted at 65 densate can be neutralized with sulfuric acid to produce

95· C. to ]00· C. This is an unexpected advantage of ammonium sulfate suitable for fertilizer.

operating at the higher leach temperatures. The solid Alternatively (not shown), the zinc can be precipitails

which contain gypsum and iron oxides are dis- tated as zinc hydroxide by adding calcium oxide to the

5,078,786

liquor from the sulfiding process. Sufficient calcium

oxide is added to provide a solution pH in the range of

about 8 to to. The precipitated zinc hydroxide is separated

by liquid-solid separation means preferably filtration.

The recovered zinc hydroxide can be conveyed to 5

a zinc plant for recovery of the zinc.

As shown in FIG. 8, the liquor from the zinc separation

process is subjected to a lime boil treatment with

calcium oxide. This lime boil step is conducted at a

temperature effective to form ammonia and calcium 10

chloride commonly above room temperature. Sufficient

calcium oxide and/or calcium hydroxide is added to

provide a pH of above about 9. This process step provides

an ammonia overhead vapor stream and a brine

stream containing calcium chloride and/or a precipitate 15

of lead hydroxide. The brine stream is preferably recycled

to the calcium chloride leach step to allow reuse of

the chloride ·values. The ammonia, which can be dissolved

in water or used directly, is preferably recycled

to a zinc plant or to a ferrite preleach process. 20

In an alternative procedure (not shown), the liquor

from the cementation step can be combined with sufficient

calcium oxide to cause the combined precipitation

of zinc hydroxide and lead hydroxide. Calcium oxide is

added to provide a pH in the range of about 8 to 10. The 25

metal hydroxides are separated by usual solid-liquid

separation techniques preferably by filtration. These

hydroxides can then be returned to a zinc plant for

recovery of the zinc. The liquor from the precipitation

process is then subjected to a lime boil as discussed 30

hereinabove. The resulting brine solution is preferably

recycled to the calcium chloride leach with the ammonium

hydroxide formed preferably being recycled to a

zinc plant or sold as a by-product.

When zinc ferrite is treated, a sulfuric acid ferrite 35

leach operation is used as discussed hereinabove. As

shown in FIG. 6, an aqueous slurry of the ferrite solids

31 is combined with sulfuric acid and ammonium hydroxide.

Sufficient ferrite solids are added to provide a

slurry containing about 30 weight percent ferrite. Sulfu- 40

ric acid is added to the extent necessary to provide a

final acid level of about to to about 60 grams per liter.

Preferably, the sulfuric acid used is spent electrolyte

from a zinc plant. Sufficient ammonia is added to accomplish

precipitation of the ammonium jarosite. Suit- 45

able ammonium sources include without limitation ammonia,

ammonia water, ammonium hydroxide and ammonium

salts such as ammonium sulfate. Preferably, the

ammonia is received from the lime boil operation described

hereinabove. Preferably, ammonium ions are SO

present at the end of the leach to the extent of about to

gil to about 20 g/1 of the solution. About 15 grams of

ammonia per liter of slurry is used in this operation. The

ferrite leach 33 is accomplished at a temperature of

about 95' C. for a period of about 12 to 24 hours. This 55

resulting slurry is subjected to a thickening liquid-solid

separation procedure. The addition of a small amount of

"seed" jarosite can be added to aid in the formation of

solid jarosite. The liquid 35 from this ferrite leach is rich

in zinc and can be conveyed to a zinc recovery process. 60

The solid material 37 is predominantly jarosite which

contains zinc and other metal values. This material is

conveyed to a calcium chloride leach 6 as described

hereinabove.

In FIG. 7 there is shown an embodiment of the in- 65

stant invention in which the product 41 from a ferrite

leach 33 is combined with additional jarosite 43 in a

repulp tank 45 and the resulting pulp 47 is fed to a cal-

cium chloride leach 6. The subsequent metal recovery

steps have been described in detail hereinabove. Also

shown in this embodiment is the movement of various

streams to and from a zinc plant 49. The zinc-rich liquid

35 from the ferrite leach is transferred to the zinc plant

for recovery of the zinc. Also, the strip liquor from the

zinc extraction is transferred to the zinc plant. Ammonia

from the lime boil is used in the zinc plant and/or the

ferrite leach step to form jarosite. The spent sulfuric

acid electrolyte from the zinc plant is used in the ferrite

leach and in zinc stripping. As discussed hereinabove,

an alternative method of zinc separation which is not

shown is the precipitation of zinc hydroxide. This zinc

hydroxide can also be conveyed to the zinc plant for

recovery of the zinc metal.

In FIG. 8 is shown a preferred process scheme for the

treatment of ferrite and/or jarosite wastes. A ferrite

containing feed 31 is subjected to a leach 33 by contacting

it with sulfuric acid and an ammonium source as

described hereinabove. The source of the sulfuric acid is

spent electrolyte 71 from a zinc plant 49. The ammonium

source is ammonia or ammonium hydroxide 73

recycled from the lime boil 25. A flocculant 75 is added

to the slurry and the solids are separated from the liquid.

The liquid 77, containing zinc, is conveyed to a zinc

plant for zinc recovery. The solids, containing ammonium

jarosite and metal values, are transferred to a

repulp tank 45 where these solids can be mixed with

additional jarosite feed 43 and slurried with a calcium

chloride source. Theleach mixture 6 is heated to above

Ito' C. under pressure and stirred. The pH of the leach

is adjusted to the desired level by the addition of calcium

oxide, calcium carbonate and/or calcium hydroxide.

After leaching, the slurry is filtered 79 with washing

to separate the solid tails 8 and the leachate 9.

The leachate is contacted 13 with metallic zinc to

form a silverlcopper cement 15. The cement is removed,

preferably by filtration 81, and the silver and

copper are purified by standard metalIurigical methods.

The liquid is mixed 55 with an extractant with the pH of

the solution adjusted by adding calcium carbonate. The

extractant _binds the zinc and allows its extraction into

an organic phase. The phases are separated and the

zinc-containing organic phase 56 is contacted 57 with

spent electrolyte 71 from a zinc plant to strip the zinc

from the organic phase into an aqueous phase. The

aqueous phase 59 containing the zinc is transferred to a

zinc plant for recovery of the zinc. The organic phase

83 containing the barren extractant is recycled to

contact fresh zinc-containing solution.

The aqueous phase from which substantially alI ofthe

zinc has been removed is contacted with lime 85 to

increase the pH to between about 8 and 9 to precipitate

lead hydroxide. The solid lead hydroxide is separated

from the liquid by f1ltration 65. The liquid phase is

mixed with lime and heated to boiling. The liquid from

the lime boil 25 is a calcium chloride-brine solution and

is recycled 87 to the calcium chloride leaching step. The

vapor containing ammonia is condensed 89 with cold

water and the resulting ammonium hydroxide solution

is recycled 73 to the zinc plant and/or to the ferrite

leach. Any uncondensed vapor can be passed into an

acid scrubber using, for example, a sulfuric acid wash to

form a solution of ammonium salt such as ammonium

sulfate.

The following examples are given for illustrative

purposes only and are not to be a limitation on the subject

invention.

TABLE 2

>103' C. tests:

95-103' C. tests: 330 gil CaCI2; 4-5 hours;

pH 1.8-3.5.

autoclave; 330 gil CaCI2;

1 hour; pH 1.8-3.5.

Conditions:

allowed recovery of the metal values even in the presence

of potassium. Sample 3 feed showed no potassium

effect.

5,078,786

EXAMPLE 1

Three jarosite feed materials were used in the Examples.

Sample 1 was taken from a first zinc plant waste

lagoon. Samples 2 and 3 were taken from different loca- 5

tions in a second zinc plant waste lagoon. The assay

results are given in Table lA.

TABLEIA

A series of runs were made in which the jarosite feed 20

material was contacted with a leaching solution containing

330 grams per liter calcium chloride, for 5 hours

unless otherwise noted. The pH was maintained in the

range of 1.8-3.5 by the addition of calcium hydroxide.

The amount of solids in the leach was about 16.8 weight 25

percent.

The leaching process was conducted at different temperatures

as indicated in Table 18. When the temperature

of the leach solution was at or below the boiling

point of the solution, the leach was conducted in conventional

glassware. When the leaching temperature

was above the boiling point of the solution, the leach

was conducted in a 2-liter Parr agitated autoclave.

The leachate was separated from solids by filtration

through a Buchner funnel using low vacuum.

EXAMPLE 3

Element. o/c Sample I Sample 2 Sample 3

Zn 4.95 7.53 7.46

Ag ozrT 7.53 12.3 11.2

Fe 27.1 22.6 19.9

Pb 2.13 6.98 7.10

Nl4 2.04 1.01 1.20

K 0.054 0.364 0.337

Na 0.122 0.162 0.156

Cu 0.32 0.3 0.347

Mn 0.362 5.17 4.47

In 0.010 0.008

Run

No.

Leach

Temp

'c.

103 (boiling)

110

120

130

95-100

180

95-100

180

K in Leach

Solution

gil

5-10

5-10

Extraction, %

Ag Pb Zn

76 81 49

78 77 47

93 89 44

96 85 42

99 94 48

72 84 41

13 35 27

79 88 (60)

73 86 34

72 84 28

76 86 37

t.r>R,un In the nid:eJ bomb reactor for approximately 5 minut~. Bo~b leach eXUac· SS

lions are gener.U)' lower than comparable autoclave values. Descnption of nickel

bomb reaclor in E>ample 3.

EXAMPLE 2

The effect of potassium on the recovery of metal 60

values from certain feeds is shown in Table 2. The

leaching conditions and apparatus of Example 1 were

used. The conventional glassware was used for the 95°

C. to 103° C. leaches with the Parr autoclave used for

the leaches conducted at temperatures greater than 103°65

C. With Sample 2 feed, the recovery of metal values

was significantly reduced when potassium was present.

Increasing the leach temperature in Run 18 however,

TABLE3A

Sample 3

Run(o) Leaching Time, Extraction, %

No. hours Ag Pb Zn

6(h) 0.5 54 44 21

6 1.0 58 56 22

6 3.0 61 79 29

19«) 3.0 57 76 30

6 5.0 72 84 28

19 5.0 73 86 34

19 8 73 82 32

19 12 75 84 36

19 24 74 85 33

(')330 gil Caell' pH 1.8-3.5

(blRun 6 had 25 gil K

(<)Run t9 had 0 gil K

minutes Ag Pb

Extraction. 9(

TABLE3B

Leaching Time,

Sample I Jarosite (130' C.)

5 25 9.4

10 53 24

Run

No.

TABLEIB

Leach

Run Temperature. Extraction. 9{

No. 'c. Ag Pb Zn

Sample I Jarosite (5-10 gil K)

1 95 76 81 49

2 103 (boiling) 78 77 47

3 110 93 89 44

4 120 96 85 42

5 130 99 94 48

Sample 3 Jarosite (25 gil K)

6 95 72 84 28

7 110 72 (62) 28

8 120 73 (62) 29

9 130 73 88 29

10 140 72 89 31

II 150 74 80 30

12 180 (I hr) 76 86 37

13 200 (2 hr) 75 84 34

14 225 (I hr) 75 87 35

15(0) 325 74 80 31

TABLE 3B-continued

5,078,786

TABLE 4C-continued

EXAMPLE 4

A series of runs was made using Sample I jarosite to

determine the effect of temperature on the amount of·

solids or pulp density which can be effectively leached. 20

The leach was conducted using 330 g eaCI2/1, at 95 to

100' c., for 6 hours, and 1.8 to 3.5 pH. The results are

given in Table 4A.

Sample I and Sample 3 jarosite materials were

leached in the nickel bomb reactor of Example 3 with 2S

330 g eaCb/l, pH 1.8-3.5 at temperature for 10 minutes.

The leach of Sample I also contained 25 gil K.

These residues were washed with hot CaCI2 brine solution

(330 g eaCh/I). The effect of a single wash is

shown in Table 4B. The effect of cumulative washings 30

on two residues is given in Table 4C.

TABLE4A

EXAMPLE 5

Closed-cycle process simulation runs were made.

Sample 3 feed was used in Runs 40 and 41 and Sample

I feed was used in Run 42. The same eaCI21each procedure

was followed as in Example I except the leach

time for each was 30 minutes. The temperature for Run

40 was 180' C. and was 130' C. for Runs 41 and 42.

Each leachate from the eaCh leaches was separated

from solid residue by filtration. In Run 40, a stainless

steel autoclave was used which might have resulted in

some cementation of silver during the leach. The

"Cycle No.", e.g. C·I, refers to the number of times

spent brine was recycled to the eaCb leach step. Results

are given in Table SA.

The filtrates were cemented with zinc to recover

silver and copper. The cementation was accomplished

by bringing the flltrate to temperature in a beaker. Zinc

dust was added and the slurry was agitated for one

hour. The resulting slurry was filtered and washed with

three 5Q.ml portions of deionized water. The effect of

reaction time and zinc requirement on the recovery of

metals by cementation with zinc powder is shown in

Table 58.

Run Leaching Time, Extraction. '*

No. minutes Ag Pb Zn

22(0) 10 68 69 42

23 20 77 76 46

5 30 (autoclave) (84) (SO) (45)

Sample 3Jarosite (180' Cfb)

24(T) 0 72 57 28

25 2.5 69 69 31

26 5 72 62 31

27 10 69 78 32

28 30 71 77 33

(0'20<;, solid. leach. Other leaches: 16.8'1< solid•.

(·lNickel bomb reactor. 330 gil Caell. pH 1.8-3.5. 25 gil K.

(elM"" to temperature. cool immediatel)·. Approximately 3min betl up.

Filter Cake

Sample 3 residue

Wash

Cumulative

Solution

Displacements

1.2

1.9

1.6

3.25

4.9

6.6

Cumulative

Washing.

Efficiency. o/c

Ag Zn

78.5 83.6

89.3 91.2

o 0

66.0 60.0

96.9 94.1

(100) 98.6

(100) 99.4

TABLE5A

TABLE 4B

(.)~ IOhds IS the' weight % ofjarosite feed in the initial leach slurry:. Ca(OHb added

so maintain the leach pH i> not included.

Run Liquor Stoichiometric Barren

De- From Zn Addition for Solution,

scrip- Run No. Ag and Cu (II) Cementation. % gil

tion (Cycle) Cementation Ag Cu Pb Ag Cu

A 40 5.9 84 42 0.004 0.042

B 41, (C-I) 2.8 6()0 98 18 0.004 0.001

C 41, (C-2) 14.2 90 660 3 0.001 0.002

D 41, (C·3) 0.7 95 24 0.1 0.001 0.002

E 42, (C-l) 1.2 94 99.4 0.8 0.002 0.002

F 42, (C·2) 1.3 98 99 21 0.001 0.004

G 42, (C-3) 7.3 98 97 99.6b 0.001 0.002

Other conditions: Temperature 60- C.

SO Time 1 bour

pH 2.6-S0

4rLow value due to Jow initial concentration, in feed to cementation.

•Anomalous value.

55 TABLE5B

Reaction Cementa- Barren

Stoichiometric Zn Time, tion.% Solution. gil

for Ag &< Cu minutes Ag Cu Ag Cu

1.0 15 62 73 0.008 0.042

60 30 71 70 0.006 0.046

60 86 66 0.003 0.051

1.5 15 67 68 0.007 0.048

30 76 68 0.005 0.049

60 90 74 0.002 0.040

2.0 15 71 70 0.006 0.045

30 81 77 0.004 0.035

60 90 85 0.002 0.023

Conditions: Temperature: 6O'C.

Feed 0.021 gil Ag. 0.152 g/l Cu.

Solution: 3.15 gil Pb, 6.3 gil Zn. 0.012 gil Fe

Cumulative

Washing.

Efficienc\'. o/c

Ag Zn

o 0

42.8 47.1

Extraction. 'i(

Wash

Cumulative

Solution

Displacements

0.6

TABLE4C

Filter Cake

Initial

Leach'ic

Solidslol

Sample 1residue

AI! Pb

Original Rewash Ori!;!ina! Rewash Zn

Sample J jarosite (130' C.)

68 83 69 79

68 84 60 78

66 82 52 72

69 82 57 77

Slurry too thick to agitate

Slurry too thiCK to agitate

Sample 3 jarosite (ISO' C.)

27 16.8 69 78

34 20 71 84

35 30 70 74 29 70

36 35 66 72 18 83

37 40 70 76 19 83

38 45 Slurr}' too thick to agitate

39 SO Slurr)' too thick to agitate

Run

. No.

5,078,786

TABLE 5B-continued TABLE 7A-continued

(from autoclave leach of Sample 3. Analysis. gil Precipitation. o/c

330 g CaC1211. 180' C. for I hour. pH Zinc Lead Zinc Lead

25 gil K. 25. I wt. o/c solids. pH 1.8-

3.5). 5 8.0 2.06 .125 76.3 96.2

Zinc 1.0 stoich. = O. 163 g Znll 9.0 1.89 .055 78.2 98.3

Requirement: (0.0064 gil for Ag. 0.157 gil for CUI. 9.5 1.47 .441 83.1 86.7

Final pH: 5.1-5.2 10.0 1.47 2.15 83.1 35.0

10.7 2.06 2.11 76.3 36.3

EXAMPLE 7

A calcium chloride leach of Sample 3 feed was con- 35

ducted with a leach solution of 330 gil CaCh at an

initial solids of 25.1 weight percent at a temperature of

about 180' C (±3' C) for one hour. The target pH was

1.8-3.5 with Ib/ton offeed of Ca(OHh added initially to

maintain the pH within the target range. The leach 40

mixture was filtered and the leachate solution, maintained

at 22' C, was mixed with hydrated lime, Ca-

EXAMPLE 8

Filtrate from Example 7 was mixed with calcium

hydroxide and heated at boil. The ammonia evolved

was recovered in the distillate or by condensing or

scrubbing the ofT-gas. The pH was maintained between

.about 8.8 and 10.5 with Ca(OHh. Results showing essentially

complete NH3 recovery is possible by boiling

the solution at the indicated pH are provided in Table 8.

Some of the zinc or lead present in the feed solution

precipitated during the lime boil step and those values

are given in Table 8 as residuals precipitated.

TABLE 8

The filtrates from Example 6 were each mixed with

hydrated lime to adjust the pH to a target final pH of

9.5. The temperature was maintained at about 60' C. for

15 minutes. The solid zinc hydroxide was separated by

filtration. using a Buchner funnel. The solids were

15 washed with three 5D-ml portions of deionized water

and dried overnight at 100' C. prior to assay. The results

are given in Table 7B.

TABLE 7B

20 Test Ca(OH12 Zinc Zinc

Desig· From Required Feed Assay. gil Precipitation

nation Test No. gil pH Feed Final o/c

A 40 19.8 0.55 5.39 (1.I5)a (80)

B 41. (C·I) 7.2 1.0 4.77 1.47 69

25 C 41. (C·2) 18.0 0.55 5.16 1.00 80

0 41 (C·3) 0.40 4.98 0.88 82

E 42, (C·I) 30.0 0.30 6.24 1.18 81

F 42. (C·2) 0.30 7.47 2.21 70

G 42. (C·3) 19.7 1.37 3.35 1.89 62

avalue calculated from solids assays.

EXAMPLE 6 10

From Run Solution Pb

Desig· Run No. Time. emf. mv Final Precipitation

nation (Cycle) minutes Initial Final pH %

A 40 IS +327 -74 1.1 38

B 41. (C·I) IS +297 -26 0.65 48

C 41. (C·2) 30 +290 -25 0.20 72

0 41. (C·3) 30 +348 -30 99.6

E 42. (C·l) 30 +246 +0 42

F 42. (C·2) IS +304 -3 0.28 46

G 42. (C·3) Data not applicable.

Other condillom Temperature: bO' C.

Feed solutIOn pH, 3.5-5.0

Filtrates from the zinc cementations of Example 5

were each brought to temperature and contacted with

hydrogen sulfide gas sufficient to provide a solution emf

below 0.0 millivolts versus a standard calomel electrode

(SCE). This quantity of H2S resulted in final solution

pH's of 0.0 to 0.2. After the indicated reaction time, the

slurry was filtered and the solid was washed with three

5D-ml portions of deionized water.

TABLE 6

From Volume NH4 Residuals

Run No. Target Ca(OH12 Reduction NH4 Assay. gil Volatilized Precipitated. o/c

Designation (Cycle) pH Required. gil o/c Feed Final Distillate % Zn Pb

A 40 10.5 33 1.76 om 3.02 99.8 48 5

B 41. (C·l) 10.5 3.2 27 1.47 3.94 100 83 10-12

C 41. (C·2) 8.75 2.4 6.8 11.4 (lOll) 4 I

0 41. (C·3) 8.75 not assayed

E 42, (C·l) 8.75 0.0 8.8 2.94 1.06 21.1 72 48 I

F 42. (C·2) 10.5 1.9 10 not assayed 81 7

G 42. (C·3) 8.60 2.0 4.3 not assayed 74 4

(OHh, to adjust the pH to the indicated values. The 55

amounts of zinc and lead hydroxides precipitated at the

particular pH was determined. Results are given in

Table 7A.

EXAMPLE 9

Analvsis. gil

pH Zinc Lead

Feed 2.58

3.0

4.0

5.0

6.0

7.0

7.5

8.68

9.33

9.64

9.58

10.2

9.7

5.4

3.31

2.83

2.98

3.28

3.13

1.4

0.0 0.0

0.0 14.5

o 10.0

o .9

o 5.4

37.8 57.7

5,078,786

the distillate (75% of the theoretical recovery based on

the Ca(OHh addition). Initial distillate contained over

300 gil NH3, demonstrating that the lime boil step can

recover a high concentration NH3 product. The usable

NH3 concentration should be in the 200 gil range as

shown by the Distillate No. 1+ 2+ 3.

TABLE lOB-continued

Run No.

43 44 45 46

Element Time. hr % Extracted

Pb I 0 4.4 0 0

4 \.5 4.7 \.9 0

TABLE 9

Time Volume NH3 CaCI2 NH3

Product minute,; ml gil gil Amount. g

Feed to lime boil 0 SOO (SO) (450) 25.0

Distillate No. I 7 13.2 305 4.02

Distillate No.2 14 17.2 173 2.97

Distillate No.3 20 17.5 146 2.55

Distillate No.4 29 27.0 70 \.88

Distillate No.5 30 29.0 37 \.08

Distillate No.6 43 30.5 20 0.62

Distillate total 134.4 13.12"

Depleted feed 363 (737)b

Overall total 497

Distillate Volume. NH3 Evolved.

% of Feed % of Theoretical

Increment Cumulative Increment Cumulative

0 0

2.6 2.6 23.0 23.0

3.4 6.\ 17.0 40.0

3.5 9.6 14.6 54.5

5.4 15.0 10.7 65.2

5.8 20.8 6.2 7\.4

6.1 26.9 3.5 75.0

·SufTlcienl Ca(OH)~ \Il,'ti added to vol.tiliz~ 17.5, }'>o;H) (JOOW efTkicncy); therefore. net efficiency after 43 minutes Wti 75.0% (13.12/75 X 100).

~culated value includes CaCI2 fonned from Ca(OH)2 in lime boil.

ooo

0.5

3.3

ooo

8.5

11.5

Run No.:

45a 45b

Element Time. hr Extractions. %

Zn I 11 52

3 7.4 36

5 8.3 38

Ag I 22 99

3 19 98

5 24 98

Fe I 0.1 0.5

3 0 0

5 2.2 0.4

Pb I 18 54

3 4.2 SO

5 0 35

EXAMPLE II

Residues from Runs 45 and 45 of Example 10 were

leached with Cach at 95°-100° C. (4Sa and 46a) and

105° C. (4Sb and 46b). The procedure of Example I was

used to leach the residue from the ferrite leach. The .

results are presented in Tables IIA and IIC. .

TABLE IlA

The recovery for the overall extraction, i.e. the ferrite

extraction (given in Table lOB) and the Cach extraction

(given in Table IIA at two temperatures) for

Run 45 are given in Table lIB.

47 55 TABLE lIB

5 hour Extractions Leach(45a) Leach(45b)

33 51 Element Zn Ag Fe Pb Zn Ag Fe Pb

59 Extractions %

64 Ferrite Leach 89 \.8 5 0 89 \.8 5 0

75 60 CaCI2 Leach 8.3 24 2.2 0 38 98 0.4 35

11 Overall extraction. % 90 25 7 0 93 98 5 35 444I

TABLE IIC

18 65 Test No. 46a 46b 8 Element Time. hr Extractions. % 6

5 Zn 1 9.5 3.9

2 3 23 37

TABLE lOB

Run No.

43 44 45 46

Time. hr % Extracted

I 43 28 58 38

4 57 45 73 58

8.5 60 53 81 76

11.5 66 53 82 83

24 80 65 89 89

I 2.9 6.2 5.2 5.2

4 4.4 9 4.4 4.8

8.5 5.1 6.6 4.7 5.8

1\.5 6.2 5.1 3.3 6

24 \.8 2 \.8 2.3

I 17 17 34 18

4 6.8 8.8 20 12

8.5 2.9 3.8 12 16

11.5 3.5 3 9.2 16

24 2.7 2.6 5 15

EXAMPLE 10

Element

Run No.

Vanable conditio", 43 44 45 46 47

Leach solution

Initial H2SO4. gil 100 100 100 150 100

Ib/ton 750 750 750 1125 750

Initial (Nt4hS04. gil SO ISO 84 84 84

Ib/ton 375 1125 625 625 625

Target H2SO4. gil 30 30 22 58 30 SO

A zinc ferrite waste was subjected to simultaneous

leach-jarosite precipitation. The ferrite was assayed and

found to contain the following components in weight

percent: Zn, 13.0; Fe, 33.4; Pb, 0.325; NH., 1.12; K,

0.095; and Na, 0.104. Silver was present in the amount 30

of 5.06 ounces per ton of waste.

The following procedure was used to leach the solid

ferrite waste. The ferrite was mixed with H2S04 (150

gil H2S04) and (NH4hS04 at 20 percent weight/weight

solids and heated at 90°-95° C. The H2S04 was 35

maintained at the target level indicated in Table lOA by

the periodic' addition of H2S04 solution (150 gil

H2S04). The mixture was agitated for 24 hours with

samples taken periodically as indicated in Table lOB.

The slurry was filtered to remove solids and the leach- 40

ate was analyzed with the results given in Table lOB.

TABLE lOA

5,078,786

TABLE II C-continued

Test No. 46a 46b

Element Time. hr Extractions.. %

5 23 43

Ag 1 8.4 14

3 41 92

5 71 92

Fe 1.4 0.8

1.2 0

0.2 0.2

Pb 18 6.2

4.2 54

0 39

80 to 230 Ibs/ft2/hour, filter two to three times faster

than the slurries from 95°_100° C.leaching. Two of the

leach slurries were flocculated with Percol 351 prior to

filtration. Flocculant doses of approximately 200 ppm

5 (solids basis) were required to coagulate the slurry

solids. The flocculated filtration Runs 51 and 52, produced

inconsistent results as shown in Table 12.

TABLE 12

Filtration Rates

Cake

Brineb H20b Thickness

Run No. Solida Filtrateb Wash Wash (Inches)

95-100' C. Leaches

53 9.6 4.0 !

54 SO 24 7.0 2.1 ,

51 99 30 10.6 9.0 3/16

55 31 9.3 3.2 6.0 ,

52 31 9.S 3.5 3.6 !

1 44 18 4.6 4.5 5/16

50 _81_' _4_2__ .illL ~ .2L!L

Average 56 22 5.8 5.0 3/16

(excluding Run 53)

130' C. Autoclave Leaches

56 228 95 16 16 5/16

48 83 43 31 20 ,

57 91 35 6.8 5.9 !

58 120 42 13 15 3/16

59 130 54 5.8 5.8 !

5 ~ 83 -...2:2.- ~ .2L!L

Average 142 59 9.9 9.9 3/1H

180 and 225° C. Leaches

49 III 44 21 27

60 75 30 7.3 II

·Sollds fillratlon rale unitS: Ibs/ftl/hour.

br.iquid filtration rate unitS: gal/ftl/hour.

The recovery for the overall extraction, Le. the fer- 40

rite extraction (given in Table lOB) and the CaCI2 extraction

(given in Table IIC at two temperatures) for

Run 46 are given in Table liD.

TABLE lID

5 hour Extractions Leach\46a) Leach(46b)

Element Zn Ag Fe Pb Zn Ag Fe Pb

Extractions t;(

Ferrite Leach 89 2.3 IS 0 89 2.3 15 0

CaC12 Leach 23 71 0.2 0 43 92 0.2 39 50

Overall extraction. o/r 92 72 15 0 94 92 15 39

EXAMPLE 12

EXAMPLE 14

Several leaches were conducted at different chloride

levels. Leach solutions of 1.8 normal, 6 normal and 10

normal chloride were used. The conditions were similar

to those of Example 1. The results are given in Table 14.

-metal extractions were calculated from feed and residue assay~.

Filtration rate determinations were performed on 55

final slurries from 15 of the jarosite leach tests. The

filtration data are summarized in Table 12. Rates were

determined using an Eimco 0.I-ft2 vacuum filter leaf

apparatus, fitted with a medium-weave polypropylene

filter cloth. The apparatus was top-loaded with leach 60

slurry and 18 to 22 inches Hg of vacuum was applied.

The filter cakes were washed first with hot CaCl2 brine

followed by water, and the washing rates determined.

The anomalously high wash rates of Runs 48, 49, and 50

probably are due to cake cracking (channeling) and the 65

values are not included in the group averages.

The filtration rate averages in Table 12 show that the

elevated temperature leach slurries, with solids rates of

25.6

Extraction·. o/r

Ag Fe

53.2 0.2

43.6

-COrrected value .fter rewash of residue. Extraction calculated (rom original

residue (incompletely washed) was 12.1Yrc:.

bC;C of specie in inlet liquor to particular process step.

It will be understood that the above description of the

present invent~on is susceptible to various modifications,

changes and adaptations and the same are intended

to be comprehended within the meaning and

range of equivalents of the appended claims.

What is claimed is:

1. In a process for recovering metal values from

waste containing MFe3(S04h(OH)6, where M is a

monovalent ion, by leaching said waste with an acidic

Example 16

A process simulation using closed-cycle steps was

performed using Sample I as feed. The following process

steps were included: CaCh leaching, Ag/Cu cementation

with zinc, Pb precipitation as sulfide, Zn

precipitation as hydroxide, NH3 evolution by lime boil,

and recycle of the processed CaCh solution to the next

stage of leaching. The values (Ag, Pb, Zn) were removed

from the leach solutions prior to recycle to the

next leach. The results are given in Table 16.

TABLE 16

Lime Boil 1 48 0.6

2 81 7

3 74 4

Average 68 4

Highest 81 7 72

22 46 84

Precipitation. ere :b

81 SS

70 0.4

62 9

71 22

81 5S

Precipitation. ere: Evolution, ere

Element

Zn Ag Pb Cu Fe Nf4

Extraction. ere:

46 94 94 47 13 (7)

48 97 92 ~1 0.01

38 96 61° 18 om

44 96 82 39 4

48 97 94 51 0.01

Precipitation. ere:b

94 0.8 99

98 21 99

98 99.6 97

97 41 98

98 99.6 99

Precipitation. ere:b

22 42 84

~ 46 53

0 0 0

13 44 68

Cycle

Process Step No.

CaC12 Leach 1

Average

Highest

Ag/Cu 1

Cementation 2

Average

Highest

Lead 1

Precipitation 2

Average

(excluding

Cycle 3)

.Highest

Zinc I

Precipitation 2

Average

Highest

5,078,786

TABLE 14

Leach Solution

gil CaCI2 100 (1.8 ~Cl-) 330 (6 ~CI-) ~SO (10 NCI-)

Feed Sample I Sample 2 Sample 2

Extraction. % 5

Ag 15 70 74

Pb 8 84 89

Zn 38 41 42

EXAMPLE 15

A direct recycle of leach filtrate to the next stage of

leaching was conducted to determine the effect of recycle

and impurities build-up on the recovery of values.

Five leach cycles were conducted using Sample 3 as 15

feed. The first three cycles used the strongly agitated

2-liter autoclave, while the final two stages were performed

in the less well agitated nickel bomb reactor.

The chloride concentration of the solution was determined

before and after each cycle and was maintained 20

at 6 normal CI (330 gil CaCh) by addition of CaCh, if

required. The 16.8% solids leaches were performed for

one hour at 180· C. and a pH of 1.8 to 4.0 by initial

addition of Ca(OHh. The leach filtrate from one cycle

was advanced to the next cycle with no intervening 25

solution treatments other than reestablishing 6 N CI

concentration, if required. -

The results in Table 15 show a steady decrease in the

apparent Pb extractions with leach cycling. From cycle

I to 4, the concentration of Pb in solution increased 30

from 15 to 45 gil (calculated), which approaches Pb

saturation in 6 N CaCh brine at 50· to 70· C. Leach

slurries'were cooled to this temperature range prior to

removal from the reactor and filtration. It was found

that the reduced Pb extractions in cycles two through 35

five were due to saturation of the cooled solutions,

causing PbCh to crystallize in the residue solids. The

standard residue washing procedure, two to three cake

displacements with 80· C. CaCh brine followed by an

equal amount of water, was not effective in totally re- 40

moving the PbCh from the residues, thus producing

low extraction values.

The true leach cycle extractions for Pb ("Pb-corr"

column in Table 15) were determined by repulping the

residues in 80· to 90· C. CaCh to dissolve the residual 45

PbCb prior to reassay of the solids. The corrected Pb

extractions show that leach cycling does not affect the

efficiency of the CaCh leach significantly, provided

thatthe leach slurry is filtered and washed under conditions

(temperature and wash volume) which assure SO

complete dissolution of marginally soluble species such

as PbCh.

TABLE 15

Run Cycle Reactor Extraction. 'iC Filtrate Assay. gil

No. No. Type Zn Ag Ph Pb-corr(l) Zn Ag Ph

12 1 A 37 76 86 86 7.0 0.071 (I~)

12a 2 A 31 76 80 87 11.6 0.135 (26)

12b 3 A 35 75 76 84 18.9 0.191 (36)

12c(2) 4 B 32 72 72 84 (25.4) (0.264) (4S)

12d 5 B 31 72 4S 81 (21.1) (0.253) (35)

\'al~ In parenth~ Ire calculated

tI"tExtractJons calculated from rc'tll:ashed residues. The Pb-corr (corrected) column presents the

actual extraction achieved in the leach.

C2tFihr.Jle was diluted 10 670/( sITength prior 10 advancin~ 10 12d leach. Dilution occurred when

washing slurry from bomb reactor.

5,078,786

solution of metal chloride in a closed system, the im·

provement comprises leaching said waste with a solu·

tion comprising calcium chloride at a temperature

above the atmospheric boiling point of the solution and

under a pressure of at least the superatmospheric autog- 5

enous pressure which develops as the system is heated.

2. The process of claim 1 wherein said temperature is

in excess of about 110· C.

3. The process of claim 1 wherein said solution has a

pH of about 1.5 to about 3.5.

4. The process of claim 1 wherein said temperature is

between about 120· C. and about 300· C.

5. The process of claim 1 wherein said calcium chloride

concentration is between about 1.0 molar and the

saturation point ef the solution.

6. The process of claim 1 wherein said temperature is

between about 150· C. and about 220· C.

7. The process of claim 1 wherein potassium is pres·

ent at a concentration greater than about 0.5 grams per

liter of said solution. 20

8. The process of claim 3 wherein said pH is maintained

by adding a compound selected from the group

consisting ofcalcium oxide, calcium hydroxide, calcium

carbonate and mixtures thereof.

9. The process of claim 1 wherein chloride is pro- 25

vided by adding calcium chloride to said solution.

10. A process for recovering metal values from waste

containing MFe3(S04h(OH)6 where M is a monovalent

ion, wherein said process comprises contacting, in a

closed system, said waste at a temperature of between 30

about 120· C. and about 200· C. and a pressure of at

least the superatmospheric autogenous pressure which

develops as the system is heated with a solution containing

between about 2.0 and about 4.0 molar calcium

chloride wherein said solution has a pH of between 35

about 1.5 and about 3.5 which is maintained by the

addition of a calcium compound selected from the

group consisting of calcium oxide, calcium hydroxide,

calcium carbonate and mixtures thereof.

11. The process of claim 10 wherein said waste com· 40

prises ammonium jarosie formed by subjecting a material

containing zinc ferrite to a leach said process consisting

essentially of:

(a) mixing an aqueous slurry of said material with

sulfuric acid and a source of ammonium ions to 45

form a sulfuric acid leach mixture;

(b) heating said sulfuric acid leach mixture to form a

solid containing ammonium jarosite and a liquid

containing zinc sulfate; and

tion of calcium chloride.

12. The process of claim 10 wherein said leaching

provides a liquid leachate and a solid residue and

wherein said process further comprises:

(a) separating said liquid leachate from said solid 55

residue -wherein said solid residue comprises an

iron oxide;

(b) contacting said liquid leachate with a reducing

metal to reduce silver cations contained in said

leachate to metallic silver and then recovering said 60

metallic silver from the liquid phase;

using a zinc recovery process to provide a liquid

solution substantially free of zinc; and

(d) adjusting the pH of the liquid solution from step 65

then heating the solution to provide a vapor con·

taining ammonia.

13. The process of claim 12 wherein said zinc is recovered

in step (c) by a process comprising extracting

said zinc by contacting the liquid phase containing said

zinc with an extractant to remove said zinc from the

liquid phase to a second phase.

14. The process of claim 12 wherein the basic material

from step (d) is selected from the group consisting

of calcium oxide, calcium hydroxide, calcium carbonate

or mixtures thereof.

15. The process of claim 14 wherein said basic material

consists essentially of calcium oxide.

16. The process of claim 12 wherein said reducing

metal is selected from the group consisting of lead and

zinc.

17. The process of claim 16 wherein said reducing

metal is lead and wherein the process further comprises

the steps (i) mixing the liquid phase remaining in step (b)

after removing said metallic silver with a sulfide compound

to precipitate the lead as lead sulfide, and (ii)

removing substantially all of said lead sulfide from the

solution and then conveying said substantially lead·free

solution to the zinc recovery process.

18. The process of claim 16 wherein said reducing

metal is lead and wherein the process further comprises

the steps of:

(a) contacting the substantially zinc·free liquid phase

from step (c) with sufficient calcium oxide, calcium

hydroxide, calcium carbonate or mixtures thereof

to provide a solution pH of between about 8 and

about 9 to precipitate lead hydroxide;

(b) removing the precipitated lead hydroxide from

the liquid phase;

calcium carbonate or mixtures thereof to the

liquid phase from step (ii) to increase the pH to

above about 10; and

(d) inc'reasing the temperature of the resulting solution

to about the boiling point of said solution.

19. The process of claim 12 wherein said zinc is recovered

by contacting said liquid phase from step (b)

with calcium oxide, calcium hydroxide, calcium carbonate

or mixtures thereof to precipitate said zinc as

zinc hydroxide and separating said zinc hydroxide precipitate

from the liquid phase.

20. A process for recovering metal values from waste

containing MFe3(S04h(OH)6, where M is a monovalent

ion, said process comprising:

(a) mixing an aqueous slurry containing a material

comprising zinc ferrite, sulfuric acid and a source

of ammonium ions to form a sulfuric acid leach

mixture;

(b) heating said leach mixture to form a solid phase

containing ammonium jarosite and a liquid phase

containing zinc sulfate;

(d) contacting said separated solid phase with a solution

comprising between 1.0 and 5.0 molar calcium

chloride in a closed system at a temperature between

about 110· C. and about 300· C. and a pressure

of at least the superatmospheric autogenous

pressure which develops as the system is heated

wherein said solution has a pH of between about

1.5 and about 3.5 to form a liquid leachate and a

solid residue;

(e) separating said liquid leachate from said solid

residue;

(f) contacting said liquid leachate with a reducing

metal selected from the group consisting of lead

5,078,786

and zinc to reduce silver cations contained in said

leachate to metallic silver;

(g) separating said metallic silver from the liquid

phase as a cement;

(h) recovering zinc contained in said liquid phase by 5

contacting said liquid phase with an extractant

which selectively removes said zinc from said liquid

phase to a second phase;

(i) contacting the substantially zinc-free liquid phase 10

from step (h) with sufficient calcium oxide, calcium

hydroxide, calcium carbonate or mixtures thereof

to provide a solution pH of between about 8 and 9

to precipitate lead hydroxide;

(j) removing the precipitated lead hydroxide from the

liquid phase;

(k) adding sufficient calcium oxide, calcium hydroxide,

calcium carbonate or mixtures thereof to the

liquid phase from (j) to increase the pH to above

about 10;

(I) increasing the temperature ofthe resulting solution

to form a vapor containing ammonia and a liquid

residue containing calcium chloride brine;

(m) condensing the vapor from step (I) to provide a

solution containing ammonium hydroxide; and

(n) recycling the liquid residue from step (1) to the

calcium chloride leaching of step (d). • • • • •

UNITED STATES PATENT AND TRADEMARK OFFICE

CERTIFICATE OF CORRECTION

PATENT NO.

DATED

INVENTOR(S) :

5,078,786

January 7. 1992

Peters et al.

It is certified that error appean in the above-identified patent and that said Letters Patent is hereby

corrected as shown below:

Column 19, line 39, insert --200-- before "lb/ton"

Column 22, line 30, delete the second listed "45" and

insert --46-- therefor

Column 27, line 41, delete "jarosie" and insert

--jarosite-- therefor

Attest:

Attesting Officer

Signed and Sealed this

Seventh Day of September, 1993

BRUCE LEHMAN

Commissioner of Patenrs and Trademarks