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
71
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 .
Ag
5.0
FIG.2
'V---------~Pb
-----t......----------Zn
1.0 2.0 3.0 .4.0
POTASSIUM IN INITIAL LEACH SOLUTION
G/L K
10
O~--..L----I------'---.......II--..-.-..I o
100
90
80
70
~. z
0
~
u
c(
0::
~>w<
U.S. Patent Jan. 7, 1992 Sheet 2 of 5 5,078,786
100
90
80 • 180° C,10 GIL K
,,,- • 95°C,
70 I
t( ,, o G/L K . , z 60 I
0 ,
F ,
~ 50
,,,
~ ,
X
,
w 40
t1' <-
30 10 GIL K
20
95°C,
10 5 G/L K
o.
0 1 2 3 4 5
LEACH TIME, HOURS FIG.3
JAROSI~5
CaO-----.
CaCI2 LEACH
6
7
___ ~IO
: RECYCLE
II
! /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
or
t r 19
1
21
Zn SEPARATIONY - Zn RECOVERY
0 23r - I NHJ l' 25",- LIME BOIL
CoC'2
RECYCLE
Co
FIG.5
Co
{ 14
META
(Zn or
CaC03
CaD
FERRITE/ 31
33
H 2 804 -. r-----NH
4
0H
,..--1_..&.---1.-_ 35
FERRITE t--~ Zn RECOVERY
LEACH
u.s. Patent Jan. 7, 1992 Sheet 4 of 5 5,078,786
.~
71"\ 49) ,I •
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.......
55
., 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
15
A9/Cu~EMENT .
MAKEUP 59
CeCI 2
WASH
ZnO
or Pb 0
43
71
31 SPENT
~ ELECTROLYTE
~ ...--=------.w.----l-----..:........t
73
57
Pb(OH)2 87
cea2 RECYCLt'
Zn STRIPPING I--;----J
56
55
PARTIAL NEUlRALlZATION 85
CeC03
CeO
89
FIG. 8
5,078,786
2
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
1
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
10
3
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
4
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.
6
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
5
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.
40
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
10
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
9
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
13
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-
14
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:
16
allowed recovery of the metal values even in the presence
of potassium. Sample 3 feed showed no potassium
effect.
5,078,786
15
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
15
2
3
Run
No.
1
2
3
4
5
16
17
18
19
6
12
Leach
Temp
'c.
95
103 (boiling)
110
120
130
95-100
95-100
180
95-100
95-100
180
K in Leach
Solution
gil
5-10
5-10
5-10
5-10
5-10
o5
10
o
25
25
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
32
33
Zn
Extraction. 9(
TABLE3B
Leaching Time,
Sample I Jarosite (130' C.)
5 25 9.4
10 53 24
20
21
Run
No.
50
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
17
TABLE 3B-continued
5,078,786
18
TABLE 4C-continued
IS
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.
S
10
Filter Cake
Sample 3 residue
Wash
Cumulative
Solution
Displacements
1.2
1.9
o
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
6S
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
42
37
38
40
32
33
30
31
31
Cumulative
Washing.
Efficienc\'. o/c
Ag Zn
o 0
42.8 47.1
Extraction. 'i(
Wash
Cumulative
Solution
Displacements
o
0.6
TABLE4C
20
30
35
40
45
SO
Filter Cake
Initial
Leach'ic
Solidslol
Sample 1residue
22
29
30
31
32
33
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.
19
5,078,786
20
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.
30
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
2.98
3.28
3.13
1.4
0.0 0.0
0.0 14.5
o 10.0
o 10.0
o .9
o 5.4
37.8 57.7
5,078,786
21
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.
22
TABLE lOB-continued
Run No.
43 44 45 46
Element Time. hr % Extracted
5
Pb I 0 4.4 0 0
4 \.5 4.7 \.9 0
47
4
2
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.
2
22
ooo
0.5
oo
o
3.3
o
ooo
8.5
11.5
24
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
Fe
Zn
Ag
Run No.
Vanable conditio", 43 44 45 46 47
45
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
25
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
23
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
24
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.
35
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
Zn
25.6
Extraction·. o/r
Ag Fe
53.2 0.2
Ph
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
26
22 46 84
Precipitation. ere :b
81 SS
70 0.4
62 9
71 22
81 5S
Precipitation. ere: Evolution, ere
72
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
2
3
Average
Highest
Ag/Cu 1
Cementation 2
3
Average
Highest
Lead 1
Precipitation 2
3
Average
(excluding
Cycle 3)
.Highest
Zinc I
Precipitation 2
3
Average
Highest
5,078,786
2S
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
10
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
10
15
27
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
(c) conveying said solid into contact with said solu· 50
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;
(c) recovering zinc from the liquid phase of step (b)
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
(c) to above about 9 by adding a basic material and
then heating the solution to provide a vapor con·
taining ammonia.
28
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;
(c) adding sufficient calcium oxide, calcium hydroxide,
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;
(c) separating said solid phase and said liquid phase;
(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
29
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;
30
(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). • • • • •
15
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65
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