Note: Descriptions are shown in the official language in which they were submitted.
11799~i~
-- 2 --
This invention relates to a process for leaching
zinc containing sulphide minerals in order to recover
metal values contained therein and at least some of the
sulphide values in the form of elemental sulphur.
In our South African Patent No. 80/0224 issued
on March 26, 1980, we described a process wherein zinc
sulphide containing material may be subjected to a
two-stage leach prccess. In the first stage the ground
material is leached, under pressure, with an aqueous
solution having a high iron content and a low acid
content. In the second stage the partly leached material
together with precipitated iron, is leached with spent
electrolyte which has a high acid content. Both leaching
stages are carried out at an elevated temperature which
is below the melting point of sulphur (i.e. 119C), and
pressure, the partial pressure of oxygen being one of the
important factors.
',,~
1179~6
The process as described in South African
Patent No. 80/0224 has the disadvantage that an excess
of zinc bearing sulphide mineral must be present to
achieve acceptable leaching rates under elevated
pressure. It is thus necessary to separate the un-
leached sulphide mineral from the final residue which
contains a mixture of sulphides, gangue, elemental
sulphur and any iron precipitate present, in order
to recycle the unleached zinc sulphide mineral back to
the first stage leach. It is also necessary to separate
the elemental sulphur from the final multi-component
residue mixture.
The present invention represents an improvement
of the process described in South African Patent No.
15 80/0224, and enables metal values to be recovered from
zinc bearing sulphide minerals in such a way that the
necessity to separate unleached sulphide mineral from
a multi-component residue, in order to recycle unleached
mineral may be eliminated.
Thus an object of this invention is to provide
a process of the general type described wherein an
excess of sulphide mineral is not necessary in order
/ . . .
11'7~ i6
-- 4 --
to achieve high dissolutions whilst also retaining the
advantages of the process of our said earlier patent.
In accordance with this invention there is
provided a process for the treatment of zinc sulphide
containing materials comprising the steps of: (i)
contacting in a first leaching stage at a temperature of
from 70C to 119C at an elevated oxygen partial pressure
of between 100 and 1000 kPa and under oxidising conditions,
a sub-divided zinc sulphide containing material with an
aqueous leach solution which has dissolved therein zinc
and impurities, from 5 to 50g/Q iron, together with ar
maximum of between 1 g/Q and 20 g/Q sulphuric acid to
dissolve zinc and other soluble metals in the zinc sul-
phide containing material by the oxidative action of
ferric ions formed by the contemporaneous oxidation of
ferrous ions to their ferric state and cause precipita-
tion of iron; (ii) separating the solids and leach liquor
resulting from such first leaching stage; (iii) subjecting
the leach liquor obtained in step (ii) to any required
purification steps follows by electrowinning of zinc
therefrom to leave a spent electrolyte having regenerated
sulphuric acid therein; (iv) contacting, in a second
leaching stage, and at elevated temperature of from 70C
to ll9~C and at an elevated oxygen partial pressure of
between 100 and 1000 kPa and under oxidising conditions,
1 31 `7~9~
-- 5 --
spent electrolyte resulting from step (iii~ above with the
solids resulting from step (ii) above to cause redissolu-
tion of iron precipitates and the further dissolution of
zinc and metal impurities from the residue of the zinc
S sulphide containing materials by the oxidative action of
ferric ions resulting from the dissolution of iron pre-
cipitate and also formed by the co-temporaneous oxidation
of ferrous ions to the ferric state, said second leaching
stage being terminated only when at least sufficient
ferric ions are present to complete the leaching of the
zinc sulphide containing material; (v) carrying out a
third leaching stage with the solids and liquids of the
second leaching stages still in contact but under substan-
tially ambient pressure, in the absence of free oxygen and
at temperatures of up to the boiling point of the solution;
(vi) separating the solids and leach solution resulting
from step (v) and; (vii) utilising the leach solution
resulting from step (vi) as at least the major portion
of the aqueous leach solution employed in step (i).
Further features of the invention provide for
conditions in step (i) to be controlled such that the iron
is precipitated predominantly in the form of goethite
which is easily acid soluble in step (iv), for iron to be
removed at an overall rate at which it is introduced into
117~9~6
the system by way of the zinc containing material; and for
the zinc containing material to be a finely ground zinc
sulphide containing mineral or ore such as sphalerite, for
example.
1 1'75~9~i~
-- 7 --
Still further features of the invention provide
for the iron added to the system by way of the zinc sul-
phide containing material to be removed from the liquid
resulting from the solid-liquid separation (step 2)
following the first leaching stage and for such removal
to be effected by the addition of low grade oxidic
manganese ore as an oxidant and neutralising agent or,
alternatively, by the addition of conventional oxidising
and neutralising agents.
Where such a manganese agent is employed the
manganese and zinc can then be removed simultaneously
in the form of manganese dioxide and electrolyte zinc
by a process as is fully described in the Complete
Specification of our issued South African Patent No.
80/1590 issued on May 28, 1980. In such a process the
manganese added to a system is electrolytically recovered
as manganese dioxide simultaneously with some zinc in a
irst electrolytic cell in which conditions are partic-
ularly favourable for the recovery of manganese dioxide.
A second electrolytic recovery follows for the main bulk
of the zinc. Further details of the process may be
obtained by referring to the said Complete Specification.
11799~6
The advantages of the improvements brought
about by the present invention will become more apparent
in the following description of the process. In this
description reference will be made to the accompanying
S drawing which is a process flow diagram.
The basic steps of the process, which are dis-
cussed more fully below, are as follows :-
The concentrate or other material may require
grinding and this is carried out in a grinding step1. The material i5 then mixed with the leaching
solution at step 2 wherein ferric ions tend to be
reduced to the ferrous state prior to being subjected
to oxidising conditions at elevated temperature and
pressure at step 3 to conduct the first leaching
stage.
The solids and liquid are separated at step 4
and the solids contacted with regenerated sulphuric acid
emanating from the zinc recovery circuit. This takes
place at elevated temperature and pressure under oxi-
dising conditions and constitutes the second leaching
/ . . .
11799~
g
stage indicated as step 5. The following non-
oxidative third leaching stage is conducted at ambient
pressure and is indicated as step 6.
The subsequent liquid/solid separation is known
as step 7 followed by treatment at step 8 of the solids
to recover elemental sulphur from the gangue. The
liquid, now low in acid is recycled to the first leach-
ing stage (step 3) optionally after partial oxidation
at step 9 to remove or consume acid if this is
necessary.
The liquid from the first liquid/solid separa-
tion of step 4 is subjected at step 10 to the removal
of iro~ added to the circuit by way of the incoming
concentrate and precipitated iron is removed at step
11. The resultant liquid is purified at step 12
such as by the addition of zinc dust and solids are
removed at step 13.
The zinc containing solution is then ready
for electrowinning which, where manganese is added
to the circuit, takes place in a ~irst electrolytic step
14 in which manganese dioxide and zinc are removed
simultaneously followed by the main zinc removal and
. . .
~:1799~
--1 o
second electrolytic step 15 whence regenerated
acid (spent electroyte) is fed to the second leaching
stage.
The starting material may be any suitable zinc
sulphide containing material but in this specification
a zinc manganese and iron containing sulphidic flo-
tation concentrate will be used by way of an example.
The concentrate enters the process at step 1. Finer
grinding of the mineral may prove to be advantageous
as it could result in acceleration of leaching rates,
and a decrease in leaching retention times and hence
in a decrease in the size and cost of the leaching
equipment. On the other hand step 1 may not be
necessary and the flotation concentrate may possibly
be used as received. The concentrate is then
pulped with leaching solution usually at between
70C and 90C in step 3 to produce a slurry with
conveniently having a 5:1 to 7:1 liquid : solid
ratio. The amount of concentrate actually slurried
` 20 per unit volume of leach solution will be such as to
permit the desired amount of zinc to be dissolved per
unit volume of solution.
. .
The leach solution entering step 2 will
typically contain between 5 and 50g/~ dissolved iron;
between 1g/~ and 15glR sulphuric acid but preferably
as low as possible and less than about 5g/~ acid;
between 70g/~ and 120g/~ dissolved zinc, but pre-
ferably about 90g/~ zinc, and between 5g/~ and 20g/~
dissolved manganese but preferably about 15g/~ to
20g/~ manganese.
In step 2, ferric ions present in the leach
solution will rapidly reduce to their ferrous state.
This is an advantage since it has been found, and it
is known by those skilled in the art that the
flocculation, settling and filtration characteristics
of the iron precipitate produced in step 3 are improved
when the ferric ion concentration is maintained at
low levels.
The slurry from step 2 proceeds to step 3,
where the first leaching stage takes place at a
temperature of between say 70C and 119C, and
preferably at 90C to 100C with an oxygen overpressure
of between 100kPa and 1000kPa, but preferably between
300kPa and 500kPa. In step 3, soluble sulphide metals
such as zinc, manganese or iron dissolve, and iron is
/-
~.,
~ ~'7~6
-12-
precipitated simultaneously until the dissolved iron
remaining in solution equals that which is capable of
being dissolved from the concentrate throughout the
entire process so that removal of this residual iron
in step 10 will result in a mass balance of the dissolved
iron throughout the circuit.
The solution at the end of the first leach in
step 3 will contain typically less than 5g/~ of free
sulphuric acid, typically 120 to 160g/~ dissolved
zinc, 5 to 15g/~ dissolved iron (depending on the
amount of soluble iron originally present in the
concentrate); and typically 10 to 25g/~ dissolved
manganese, (depending on the amount of manganese
dissolved in step 3).
The slurry from step 3 is dewatered in step 4.
In batch laboratory experimentation it ~Jas found that
flocculation of the slurry was assisted by prior
mixing of the pulp with say two parts of previously
prepared filtered leach liquor with same solution
composition as the slurry solution composition, to
one part of fresh slurry. (In a continuous operation
such dilution would probably occur automatically in
117~
the feed well of a thickening apparatus). The
flocculated solids are permitted to settle, after
which the supernatant liquor is decanted and the
flocculated, settled solids are filtered.
The leach li~uor filtrate from step 4 proceeds
to step 10 where the re~aining dissolved iron is re-
moved at between 70 to 100C by simultaneously treating
the solution with a neutralising agent ~such as zinc
oxide calcine,zinc oxide dross or zinc dust), and an
oxidant (such as air or oxygen). It has been found
that a ground oxidic manganese ore can be used to
simultaneously oxidise and behave as a neutralising
agent and so enable the dissolved iron to be precipitated
from solution. The removal of part or all of the
residual dissolved iron by such an oxidic manganese
ore, could result in the necessity to use significantly
less zinc oxide, zinc dross or zinc dust in step 10,
and result in the recovery of the e~tra manganese ions
introduced into the solution, as valuable battery grade
manganese dioxide in step 14.
The solution from step 10 is flocculated and
filtered in step 11, and the filter cake is disposed
of after suitable washing to recover soluble values
. . .
1~79~6
-14-
therefrom. The filtrate from step 11 containing
up to 180g/~ zinc, and up to 30g/~ manganese and
at a pH greater than 2,0 is subjected to several
further purification steps at between 70C and 90~C
as typified by step 12; to reduce contaminants such
as iron, cobalt, copper, cadmium, etc. to very low
levels by use of processes well known to those skilled
in the art. The purified solution from step 12 is
filtered as denoted by step 13 and proceeds to the
optional electrolytic step 14 for simultaneous electro
recovery of part of the zinc at a cathode, and
sufficient manganese as manganese dioxide at an
anode to maintain a manganese ion balance throughout
the process. The solution enters the first electro-
winning step 14 direct from the purification steps and
thus initially contains a very low acid concentration
and high zinc concentration.
During the first electrolysis step 14 sulphuric
acid is regenerated, one mole of acid being formed for
each mole of zinc or manganese which is electro
deposited. The process for the simultanecus electro-
winning of zinc and manganese has been fully expounded
in our said South African Patent No. ~0/1590.
11799~i6
Needless to say, if only relatively small amounts
of manganese are dissolved throughout the process,
alternative means for controlling the dissolved
manganese at desired levels exist, of which those
skilled in the art would be aware.
The purified solution from step 13, or optionally
from step 14 proceeds to step 15 where sufficient zinc
is electrowon to maintain a zinc balance throughout
the circuit by m~ans of a conventional zinc electro-
winning process.
The solution, or so-called return electrolyte
from step 15 containing typically 140 - 180g/~
sulphuric acid, 40 - 60g/~ zinc iron and 5 -20g/~
manganese ions proceeds to the second leaching stage
at step 5 where it is contacted at typically between
. 70C and 119C but preferably at 90C - 100C with the
- filtered solids from step 4 which contain unleached
sulphide mineral, elemental sulphur and precipitated
iron.
In step 5 much of the iron precipitate is
~ rapidly dissolved by the acid in the return electro-
.~ /
-
1:179~6
-16-
lyte to form dissolved ferric ions in solution, and
this results in a rapid decrease in the sulphuric
acid concentration of the solution. The ferric
ions produced by the dissolvèd iron precipitate react
with the sulphides present to dissolve the sulphides
and form ferrous ions. However, since oxygen at an
oxygen overpressure of between 100 - 1000kPa, and
preferably, at about 300 - 500kPa is present, the
ferrous ions are continuously oxidised to their ferric
state, thus consuming acid and permitting dissolution
of the sulphides by the ferric ions to proceed.
Dissolution of the sulphides virtually to completion
could conceivably be attained under pressure conditions
in step 5, but this is not necessary and the retention
time in step 5 can, as provided by the invention, be
shortened considerably.
This second leaching stage can be terminated
when the leach solution contains at least stoichiometri-
cally sufficient ferric ions to permit dissolution of
the sulphides to proceed substantially to completion
without any elemental oxygen being present. At this
stage the liquid/solid mixture proceeds to the third
leaching stage at step 6.
'99~6
-17-
It is desirable that conditions be adjusted
so that at the end of leaching step 6 the sulphuric
acid concentration be as low as possible without
precipitation of the ferric ions actually taking
S place and should be less than about 15g/~.
The temperature in step 6 should be between
70C and the boiling point of the solution, but
preferably above 90C. Leaching in step 6 should be
permitted to proceed until economically negligible
further reduction in the ferric ion concentration
with time, due to leaching is detected. It is
desirable that the ferric ion concentration at the
end of step 6 be minimised, yet be such that realistic
rates of leaching of the residual soluble sulphides
be attained.
The leach slurry from step 6 is filtered in
step 7 to produce a filter cake containing unleached
sulphides, elemental sulphur, gangue and some acid
insoluble iron precipitates (e.g. such as plumbo-
jarosite, if soluble lead was present in the originalconcentrate).
.
J 17~ 6
-18-
The elemental sulphur may be recovered from
this filter cake in step 8 by methods known to those
skilled in the art. Optionally, methods exist for
the recovery of the elemental sulphur directly from
the leach solution after step 6. The filtrate from
step 7 should now be at a composition similar to that
previously described for the solution entering the
first leaching stage step 2. If the sulphuric acid
concentration in the solution emanating from step 7 is
undesirably high (e.g greater than about 15g/~) it
is possible to reduce this acid level to a point at
which hydrolysis of the ferric species is just avoided
by the oxidation of ferrous ions to their ferric state
using air, oxygen or ground oxidic manganese ore.
If air or oxygen is used, it is probable that an
oxygen over-pressure of say 100kPa to 500kPa will be
required to reduce the retention time in step 9 to
acceptable limits.
The reason for reducing the acid concentration
in step 9 and the ferric ion concentration in step 2
is to improve the flocculation and filtration charac-
teristics of the iron precipitate produced in step 3,
as it is well known to those skilled in the art that
the production of goethite-type iron precipitates in
. . .
1179966
-19-
the temperature range of 70C - 119C is favoured
by higher pH's and low~ ferric ion concentrations.
EXAMPLE 1
In this example 830,0g of a dry flotation
concentrate assaying Zn - 50,9%; Fe - 9,1%;
Mn - 2,5%; Pb - 0,47~; S(total - 30,8%);
Cu - 0,14%; Cd - 0,08%; Co -0,01~; Si 2 -
2,19% was rod milled until 100,0% of the con-
centrate passed through a 45,0~m screen. The
size distribution of this milled product, when
measured in a WARMAN CYCLOSIZER was as follows :-
TABLE 1
. I
Size in Microns % Greater than
j 44 i i
41 0,8
3~ 6,1
23 24,8
: . 16 42,7
12 54,2
I _ i
* Trade mark
117g~
-20-
This 850,0g of dry milled concentrate was added to
5,0~ leach solution which initially contained 97,0g/~
ZN; 18,6g/~ Mn; 38,9g/~ Fe ; 0,5g/~ Fe ; and
9,6g/~ H2 S04 . The reaction in the first leaching stage
was allowed to proceed for 45,0 mins at 100,0C with
a 500kPa oxygen overpressure. After 45,0 mins. the
leach solution contained 11,3g/~ Fe and about
121,0g/~ Zn and 19,96g/~ Mn~ The leach solids were
flocculated and filtered but not weighed and assayed.
All the leach solids referred to above were
then added to 5,0~ simulated return electrolyte con-
taining 50,6g/~ Zn; 21,9g/~ Mn; 146,0g/~ H2 SO4 .
The second leaching stage reaction was permitted to
proceed for 30,0 minutes at 100,0C with a 500kPa
oxygen overpressure. At this point the solution
contained 22,2g/~ Fe ; 11,8g/~ Fe ; and 19,2g/~
H2 S04 and the pressure was diminished to atmospheric
and the supply of oxygen terminated. The ferric
species present in the leaching solution was then
permitted to effect leaching for a further 9~hours
without oxygen present to allow dissolution of the
sphalerite to continue. This constituted the third
'
:
1~799~
-21-
leaching stage. The final leach solution assayed
25,7g/~ Fe ; 7,2g/~ Fe ; 14,8 g/~ H2SO4; 108,1g/~
Zn and about 24,8g/~ Mn. Table 2 below summarises
the change with time of the ferric ion, ferrous ion and
S sulphuric acid concentration, the percentage zinc
remaining in the leach residue, and the calculated total
zinc extraction in respect Gf the third leaching stage.
The final leach solids which were filtered, washed and
dried, weighed 257,0g and assayed 0,62% Zn; <0,2% Mn
and 8,8% Fe. The overall percentage of zinc,
manganese and iron leached in this example were thus
about 99,5%; 97,0% and 74,0% respectively.
TABLE 2
Time Fe3+ Fe2+ H2 S04 Zn in re- Total dis-
(mins) g/~ g/~ g/~ i solution of
0 22,2 11,8 19,2 15,50 88,11
19,0 15,4 15,4 12,20 90,96
15,1 19,0 14,8 9,05 93,71
12,5 21,3 15,0 6,46 95,55
150 10,6 23,2 14,8 5,50 96,24
210 9,1 24,4 15,0 3,34 97,73
282 8,7 25,4 15,0 2,76 98,15
380 7,6 25,3 14,9 1,67 98,89
S&0 7,2 25,7 14,& 0,62 99,50
li~79~ 6
-22-
EXAMPLE 2
The flotation concentrate described in Example
1 was used in this example "as received" from the
flotation cells without further milling. This
concentrate was coarse (relative to that used in the
first example), as only 36,7~ of the concentrate passed
through a 45~m screen. The size distribution of
this material, when measured on a Warman Cyclosizer
was as follows :-
TABLE 3
Size in % Greater
Microns than
29,3 61,52
22,8 71,21
16,4 79,80
11,1 84,41
8,5 100
935,0g of this concentrate containing about 10,0
moisture was added to 5,0~ leach solution whichinitially contained 98,0g/~ Zn; 18,5g/~ Mn;
li'~9
( -23-
38,9g/~ Fe ; 0,5g/~ Fe ; and 9,6g/~ H2SO4. The
first leach stage reaction was allowed to proceed for
50 minutes at 100,0C with a 500,0kPa oxygen overpressure.
After 50 minutes the leach solution contained 8,9g/~
Fe ; 1,0g/~ Fe ; 3,0g/~ H2SO4 and about 137,0g/~
Zn and 20,5g/~ Mn. The leach solids were flocculated
and filtered, but not weighed and assayed.
All the leach solids referred to above were
then added to 5,0~ simulated return electrolyte con-
taining 50,6g/~ Zn; 21,9g/~ Mn and 146,0g/~ H2SO4 and
zero dissolved iron. The second leach stage reaction
was permitted to proceed for 110,0 minutes at 100,0C
with 500,OkPa oxygen o~erpressure. At this point
the solution contained 28,2g/~ Fe +; 6,2g/~ Fe + and
13,2g/~ H2 S04, and because of the ferric species
present, the leaching was permitted to proceed by
way of the third leach stage for a further 5,0 hours
at 100,0C without oxygen present (i.e no oxygen
overpressure) to allow dissolution of the sphalerite
to continue. The final solution assayed 17,2g/~
Fe ; 16,6g/~ Fe ; 11,0 g/R H2SO4; 95,0g/~ Zn
and about 24,1g/~ ~n. The final leach solids which
were filtered, washed and dried, weighed 247,7g
. . .
?~
-24-
(dry) and assayed 3,6% Zn; 0,3% Mn; 8,4% Fe and
60,4% elemental sulphur.
The overall percentages of zinc, manganese and
iron leached in this example were thus about 98,0%
96,5~ and 73,0% respectively. It will thus be noted
that the coarse material leached extremely well only
not quite as well as the finely milled concentrate.
EXAMPLE 3
This example serves to illustrate the optional
use of an oxidic manganese ore as a neutralising and
oxidising agent to remove dissolved iron from a leach
solution.
180,0g of a pyrolusite-type oxidic manganese
ore (assaying Mn - 24,9%; Fe - 15,6% and having 85,2%
passing through a 53,0~m screen) was added to 5,0
solution assaying Zn - 150,0g/~; Fe - Og/~;
Fe - 9,6g/~; H2SO 4 - 9,8g/R; Mn - 20,3g/R which
was at 90,0C and was vigorously stirred. Table 4
summarises the change with time of the Mn , Fe
and H2 S04 concentrations, and the pH. It is observed
. . .
li799~
-25-
that the total acid and iron concentrations drop
to about 2,4g/~ and 1,0g/~ respectively in only
two hours.
TABLE 4
_
Time Mn2+ Fe2+ Fe3+ H2 S04 pH
(mins) g/~ gi~ g/~ g/~ __
0 20,3 9,6 0 9,8 0,5
22,1 3,4 2,0 0 1,57
25,4 2,3 1,4 1,0 1,48
25,3 1,6 1,3 1,0 1,40
26,3 1,0 1,1 2,0 1,36
27,3 0,2 1,0 2,2 1,29
. 120 28,3 0 1,0 2,4 1,24
180 27,3 0 1,0 2,6 1,22
Use of 180,0g oxidic manganese ore to remove iron
from 5,0~ solution at 90,0C which initially contained
150,0g/~ zinc ions.
Production of iron precipitates demonstrating
improved flocculation and filtration characteristics
was favoured by somewhat slower rates of iron removal.
11799f~
-26-
EXAMPLE 4
This is another example of the use of
oxidic manganese to remove dissolved iron from
typical zinc leach solution. Sixteen grams of a high
grade manganese dioxide (assaying 74~ as MnO2) was
added per litre of solution which initially contained
12,1g/~ Fe and 1,4g/~ Fe . After 60 minutes
stirring at 90C the solution contained less than
0,1g/~ Fe and 2r5g/~ Fe . After a further 50
minutes 14,7 grams of zinc dioxide calcine was added
to the solution and after 40 minutes the solution
contained no Fe and about 0,6g/~ Fe
EXAMPLE 5
850g of dry flotation concentrate was
leached without additional milling in steps 2 and
5. This example gives the results of leaching the
residue contained in the solution obtained at the
end of step 5 in step 6 at a constant temperature
of 90C. The conditions and results are summarised
in Table 5.
. . .
1179~
-27-
TABLE 5
Time Fe Fe3+ H2 S04 % Zn in ~ Dissolu-
(mins' g/~ g/~ g/~ residue tion of Zn
0 6,0 32,8 11,2 __ 92,2
30 9,5 29,2 11,4 __ 94,6
10014,1 24,4 11,8 9,2 97,7
16016,7 21,9 12,8 7,0 98,2
24018,4 20,4 13,6 5,0 98,7
26019,4 20,6 13,2 3,9 99,0
It is observed that the percentage dissolution
of zinc increases with time to a very high value, in-
dicating that relatively high dissolutions of zinc
can be obtained when leaching flotation concentrate
which has not been ground finer. Only additional
time is required and this must be weighed up relative
to the cost of milling.
EXAMPLE 6
This example is similar to example 5, except
in that the temperature of the leach solution in step
6 was permitted to decrease with time from 97C to
70C.
. . .
1179966
85Gg of dry flotation concentrate was leached
without additional milling in steps 2 and 5. Table
6 summarises the conditions and results of leaching
the residue in the solution obtained at the end of
step 5.
It is observed by comparing the results of
Tables 3 and 4 that the effect of the decrease in
temperature during step 5 of Table 6 is to increase
the leaching time required to achieve a given percentage
dissolution of zinc. It is intereating to note in
Table 4 that the H2 S04 concentration decreases with
time, and the total dissolved iron (Fe + Fe
increases with time.
TABLE 6
.
Time Temp Fe2+ Fe3+ H2SO~ % Zinc % dis
(mins) C g/~ g/~ g/~ in res- solu-
idue t on
. -- _
0 97,0 7,5 28,5 9,5 11,2 91,8
86,0 11,1 25,5 9,0 9,2 93,3
71,0 13,0 23,6 8,5 8,2 95,6
150 70,0 14,0 22,8 8,4 5,8 95,6
360 70,0 15,9 21,2 6,8 3,8 97,3
480 70,0 15,7 21,8 3,6 2,2 98,5
780 70,0 13,7 23,6 1,0 1,5 99,0
~17~9~6
-29-
It is possible that as the temperature de-
creases the solubility of ferric ions in solution
increases thus permitting iron precipitate which had
previously hydrolysed, to redissolve, thus con-
suming acid and increasing the tot~l dissolved ironconcentration in solution. This phenomenon was
not observed in Table 5, possibly because the
temperature was maintained at gOC.
The invention therefor provides an effective
and economic process for the leaching of zinc sulphide
containing materials.
