Note: Descriptions are shown in the official language in which they were submitted.
Outokumpu Oy, Ou-tokumpu
80 0851
A method for the recovery of valuable metals from finely-
divided pyrite
The present invention relates to a process for the recovery of
valuable metals from impure, finely ground pyrite, either
directly or by smelting the pyrite by means of, for example,
non-oxidizing gases at an elevated temperature in order to
produce iron matte with a valuable-metal content ~nd dust-
bearing gases. This invention relates in particular to a
multiple-stage process for the recovery of valuable metals
from the roasting residue obtained from the oxidizing roasting
of pyrite or from the iron matte, dusts and gases obtained from
the smelting of pyrite.
From Finnish Patent 32 465 there is known the smelting of
finely ground pyrite in a flash smelting furnace at an elevated
temperature by means of non-oxidizing gases. The retention time
of the suspension in the reaction shaft is approx. 1-2 s,
during which time the solid materials heat up and melt in
accordance with known chemical reactions. Aftex leaving the
reaction shaft the gas contains sulfur dioxide in excess,
and therefore the gas has to be rèduced in the rising shaft
.. . ... . . ... . .. ..
of the flash smelting furnace in order to recover the
elemental sulfur, as described in, for example, Finnish Patents
44 797, 45 948 and 45 037. The reaction products obtained are
iron ma-tte, iron slag and gas, from which the sulfur is
recovered, and also dust.
The iron matte contains var:ious compounds of iron, copper, zinc,
cadmium, lead, cobalt, nickel, gold and silver. The dusts
separated from the gases, for their part, contain various
compounds of lead, zinc, arsenic, cadmium, copper, cobalt,
nickel, mercury and selenium.
The pyrite used as raw material contains varying amounts of
different valuable metals, and in a typical case the valuable-
metal contents can be, for example:
Cu 1-1.5 % Cd 0.001-0.1 %
Zn ~-4 Ag 0.001-0.1
Pb 1-2 Au 0.0001-0~01
As 0.5-2 Hg 0.001-0.1
Co 0.001-0.1 Se 0.001-0.1
Ni 0.001-0.1
In a flash smelting furnace, the distribution of these valuable
metals is as follows:
Zn 30-40 % in matte, the remainder in dusts
Pb 5-15 - " -
Cd 30-40 _ ~ _
Cu 85-95 - " -
Co 85-95 - '
Ni 85-95 - " -
Au and Ag approx. 100 % in the matte
Hg and Se approx. 95 % in the gases
As approx. 95 % in the gases, the remainder in the dusts
After leaving the e]ectric filter, the gases are directed to
the sulfur condensation towers, in which the elemental sulfur
is recovered by known methods.
s
Thereafter, the valuable metals present in the matte and the dust
can be recovered by various methods. Thus, there are numerous previously known
methods for the recovery of various valuable metals, and the problem has been
how to select those very methods which are best suited for use in conjunction
with each other so as to obtain as a final result a maximally economical and
disturbance-free total process for the recovery of valuable metals from impure
pyrite.
The object of the present invention is thus to produce a total process,
better than previous ones, for the recovery of valuable metals from finely
ground pyrite.
Thus, in one aspect the present invention provides a process for the
recovery of metal values from impure finely-divided pyrite either directly or
by first smelting the pyrite with the aid of non-oxidizing gases at an elevated
temperature in order to produce an iron matte with a valuable-metal content and
dusty gases, comprising in combination the following steps: a) part of the
pyrite or iron matte which contains valuable metals is first subjected to an
oxidizing roasting and thereafter, together with the remainder of the pyrite or
iron matte, to a sulfatizing roasting in order to convert the valuable metals
present in the pyrite and its roasting residue or in the iron matte to a
sulfate form and the iron to an oxide form, whereafter the sulfates are leac~ed
and the obtained solution is separated in order to recover the valuable metals
from thc solution, and the insoluble iron oxide is used for iron production;
b) the dusts which contain valuable metals are separated from the gases; c)
the separated dusts are fed directly to the sulfatizing roasting of step a)
after any arsenic present therein has first been removed; and d) the sulfur is
condensed from the gases.
In another aspect, the invention provides a process for the recovery
of metal values from impure finely-divided pyrite either directly or by first
- 3 -
smelting the pyrite with the aid of non-oxidizing gases at an elevated tempera-
ture in order to produce an iron matte with a valuable-metal content and dusty
gases, comprising in combination the following steps: a) part of the pyrite or
iron matte which contains valuable metals is first subjected to an oxidizing
roasting and thereafter, together with the remainder of the pyrite or iron matte,
to a sulfatizing roasting in order to convert the valuable metals present in the
pyrite and its roasting residue or in the iron matte to a sulfate form and the
iron to an oxide form, whereafter the sulfates are leached and the obtained
solution is separated in order to recover the valuable metals from the solution,
and the insoluble iron oxide is used for iron production; b) the dusts which
contain valuable metals are separated from the gases; c) the separated dusts
are directed to a separate sulfatizing roasting in order to convert the
valuable metals to sulfate form and the iron to oxide form, whereafter the
sulfates are leached in accordance with step a) and the insoluble, arsenic-
bearing iron oxide is rejected; and d) the sulfur is condensed from the gases.
The invention is described below in greater detail with reference
to the accompanying drawings, in which
Figures 1-3 depict flow diagrams of three different alternative total
processes,
Figure 4 depicts in greater detail the recovery of sulfur from the
gases of a sulf.ur smelting furnace, and
Figure 5 depicts one alternative method .for the sulfating roasting
of the roasting residue.
In the alternative process according to Figure 1, the sulide-oxide
matte from the flash smelting furnace is granulated during stage 1 and ground
during stage 2~ Part of the matte is dead roasted during stage 3, the
temperature being 950-1050C. The following reactions occur during this stage:
1) FeS t 1.75 2 ->0 5 Fe20 ~ S0
- 3a -
2) Fe304 ~ 0-25 o2~ 1.5 Pe2 3
3) FeO + 0.25 O2~ 0.5 Fe203
4) PbS t~ 1.5 2~ PbO ~ S02
3b -
5) ZnS -~ 1.5 2 + Fe2O3 ~ ZnO ~e2O3 + SO2
6) 2 2 2 3 > Cu2O Fe23 + S2
The sulfides of Co, Ni and Cd also oxidize, forming respective
oxides or ferrites. The calcine of this stage is directed to
the subsequent stage 4, i.e. the sulfating roasting (FI Patent
41 874), to which the final matte is also introduced. The
sulfating temperature is 65~-750 C. The iron compounds are
oxidized mostly to hematite and partly to magnetite and water-
soluble sulfates. The compounds of Cu, Zn and Pb are sulfated
to sulfates, the copper and zinc sulfates being water soluble,
and to some extent also to zinc and copper ferrites, which are
not soluble in water.
Co, Ni and-Cd are also sulfated to their respective sulfates.
The calcine is leached during the subsequent stage 5 in the
iron residue washing water, which is obtained from stage 6,
whereby the water-soluble sulfates dissolve. The iron residue
Fe2O3 is directed to the chloride leaching cycle. The acid
sulfate solution is directed to the recovery of metals.
The dusts emerging from the flash smelting furnace are directed
from the electric filters to the sulfating roasting 12, in
which the temperature is 620 -700 C. The necessary Na is added
in the form of, for example, Na2SO4. During this stage, most
of the iron sulfide oxidizes to hematite and to some extent
also to magnetite and water-soluble sulfates. The sulfides of
Cu and Zn are in part sulfated to sulfates and ir part oxidized to
water insoluble ferrites. Pb, Co, Ni and Cd are sulfated, lead
sulfate being water insoluble. The calcine is leached during
the subsequent stage 13 in the washing water of the precipitate
obtained from the iron removal stage 17, the washing water
coming from stage 18, and thereby the sulfates of the non-
ferrous metals present in the dusts are caused to dissolve. The
precipitate is directed to the strong acid leach stage (SAL)
1~, and to this stage the necessary amourlt of return acid is
- introduced from the zinc electrolysis of stage 22 in order
to dis~olve the ferrites of the dust calcine. The temperature
of the stage is 90-100 C, and its H2SO4 concentration is
30-150 g/l. The leach residue is washed during s-tage 15 and
directed to the chloride leach cycle. The acid sulfate solution
is directed to the metal recovery.
The sulfate solution obtained from the matte leaching stage 5
is directed to the Cu cementation stage 7, in which the copper
is cemented by means of iron and thereby cement copper is
recovered. The sulfate solution is directed to the iron removal
stage 8, during which that part of the electrolysis return acid
which is not used for the leach during stage 14 is neutralized
by means of, for example, limestone, and the iron in a ferrous
form is removed by oxidizing it by means of air, the temperature
being 85-95 C, and by precipitating it in the form of goethite
by using, for example, lime as a neutralizing agent. The
produced gypsum and iron precipi-tate is washed during the sub-
sequent stage 9 and removed, the washing water of stage 9 is
directed to stage 8 or 10, and the sulfate solution is directed
to the metal hydroxide precipitation stage 10, during which Zn,
Co, Ni and Cd are precipitated in the form of hydroxides by
means of, for example, lime. The precipitate is used as a
neutralizing agent in the iron removal stage 17 of the dust
cycle. Part of the metal-free sodium sulfate solution is used
for the precipitation of lead carbonate, when the lead is
recovered in the form of lead carbonate, otherwise it is
vaporized in a form suitable to be directed to the grinding
of matte during stage 2, through which the sodium sulfate
necessary for the sulfating roasting of matte is also ob-tained.
The dust of the sulfate matte is directed from the calcine
leaching stage 13 to the copper cementing stage 16, during !
which the precipitation takes place by means of, Eor example,
iron. The solution is directed to the iron removal stage 17,
to which the solution of the SAL stage is introduced, and par
of the washing water of the precipitate washing stage 15 is
neutralized by means of, for example, lime and the metal
hydroxide precipitate from stage 10. The iron and the gypsum
residue are washed with water during stage 18, and the washin~
water from this stage is used in the water leach stage 13. The Zn sulfate
solution is directed to the subsequent solution purification stage 19, during
which Cu, Co and Ni are cemented by known methods ~FI Patent 52 595 and Canadian
Patent No. 1,113,253) by means of zinc dust, using arsenic trioxide as the
reagent, and cadmium is cemented by means of Zn dust during stage 20 by methods
known per se (FI Patent 50 715).
The solution is concentrated by evaporation during stage 21. The
concentrated sulfate solution is directed to the zinc electrolysis 22, which
operates by the electrowinning method. The zinc is recovered rom stage 22,
and the return acid is used for -the purposes described above.
The iron residue from stage 6 is directed to the chloride leaching
stage 23. The iron residue is leached in a sodium-calcium chloride solution,
whereby the water-insoluble lead sulfate is brought to a water-soluble form.
Chlorine gas is added for the oxidation of the noble metals, and lime is used
as a neutralizer and for the removal of sulfates. In order to prevent the
concentration of certain impurities, e.g. iron, in the solution cycle, part of
the solution is removed, and in order to maintain the Na level, NaCl is added.
The obtained Fe203 is washed during stage 24 and fed directly to the iron
production. The chloride solution from stage 19, containing Pb, Ag and Au, is
directed to the second leach 15, during which the iron residue of the strong
acid leach of the sulfated dusts is leached. In this stage also, lime is used
for sulfate removal and pH control. Air blasting is used for removing the
excess chlorine gas. If iron has dissolved during the previous leach 23, it is
precipitated together with the arsenic of stage 14, carried by moisture. The
arsenic-bearing iron residue of the stage is taken to the disposal area after
the washing stage 26. The chloride solution, which contains some copper and
zinc in addition to the lead, silver and gold, is combined with the washing
water from stage 26 and directed to stage 27, during which the noble metals
- 6 -
s
and the copper are separated from the solution by, for example, cementation by
means of zinc dust. The lead and ~inc are precipitated during the subsequent
stage
- 6a -
28 by means of, for example, hydrogen sulfide. The precipitate
is washed during stage 29, and the washing water is combined
with the chloride solution and directed to the evaporation
stage 33, from which the solution, from which the excess
NaCl has been removed,is recycled to stage 23.
The sulfides of lead and zinc from stage 29 are dissolved in
water by oxidation with air iIl an autoclave during stage 31.
The lead sulfate precipitate is washed during stage 32, and the
washing water is directed to the autoclave. The zinc sulfate
solution is directed to the metal hydroxide precipitation stage
10. The lead sulfate is converted to lead carbonate during
stage 33 by leaching the residue in a sodium sulfa-te solution
(from stage 8) by means of sodium carbonate. The lead carbonate
precipitate is washed during stage 34 and the washing water is
recycled to the previous stage. The solution from stage 23
contains sodium sulfate, and it is used, alternatively with -the
NaS04 solution of stage 10, for the grinding of matte.
The roasting cycle for the dusts can be combined with the
matte roastin~ cycle, when the arsenic has first been removed
and thus cannot pollute the iron ore. The double roasting
replaces the strong acid leach of the dusts, and increases the
yield of sulfate.
In the alternative method illustrated in Figure 2, the leaching
cycles for the roasting residues of matte and dust have been
combined, whereby the strong acid leach of the dust line is
eliminated, and the arsenic is removed from the dusts during
a separate sta~e, ~and tshe calcine leach residue from the stage
9 is directed to the sulfating roasting ~ of matte.
In this manner, only one iron residue is obtained, and this
residue is directed to the chloride leaching cycle. The sulfate
solutions obtained from the leaching of the products of roasting
can now also be -treated in one cycle, in which case the solution
is directed via the copper cementation stage 10 to the iron
precipitation stage 11, to which is fed, as a neutralizing agent
in addition to the lime addition, the precipitate of the zinc
hydroxide precipitation stage 18 subsequent to the neutralization
16 of the return acid obtained from the zinc electrolysis. Th~
sulfate solution now passes via the Cu, Co, Ni and Cd removal
stages 13 and 14 to the evaporation s-tage 15 and further to t:he
zinc electrolysis. Since there is now only one Fe203 residue,
the separate chloride leaching stage for the dusts is eliminated
and the chloride solution cycle differs from -the process
alternative of Figure 1, subsequent to the washing stage for
the precipltate obtained from the lead su]fide and zinc sulfide
precipitation stage,in that the sulfates are removed from th~
solution during stage 25 by means of a lime addition, and the
obtained residual gypsun precipitate is washed and the solutions
are combined, and the excess NaCl is removed from the solution
prior to the evaporation stage. The lead carbonate is recovered
as in the alternative illustrated in Figure 1.
If it is desired, in accordance with Figure 3, to combine the
leaching cycles of the residues from the roasting of matte
and dusts and not -to carry out a separate arsenic removal from
the dusts, the roasting residue from the sulfating roasting
of the dusts is leached during the strong acid leach (SAL) 9
in order to dissolve the ferrites. However, in this case the
electrolysis return acid is not fed to the SAL as in the process
alternative of Figure 1, but a pure H2SO4 solution is used in
order to maintain a suitable sodium concentration in the
sulfate solution, and thus the sodium cannot interfere with
the Zn electrolysis. In other respects the sulfate solution is
pw:ified as in the process of Figure 1, i.e. the solutions
from the leaching stages 5 and 8 for the matte and the dusts
are combined and the sulfate solution passes via the copper
removal stage 11, the Fe removal stage 12, the Cu, Co, Ni
cementation stage 14, and the cadium removal stage 15 to
evaporation and further to ~he Zn electrolysis 17, the return
acid of which is neutralized during a separate stage 18 by
means of, for example, limestone, and during the subsequent
stage 19 the Zn(OH)2, which is used as a neutralizing agent
during the iron removal stage 12,is precipitated. The Na2SO4
solution is evaporated and used, alternatively with the Na2SO4
solution obtained from the lead carbonate precipitation stage,
for the grinding of matte.
The chloride leaching cycle is the same as in the alternative
of Figure 1, i.e. the Fe residue obtained from the dust cycle
is directed to the waste disposal area owing to the arsenic
present in it.
Figure 4 illustrates in greater detail how dusts are separated
from the dusty gases from the flash smelting furnace and how
the sul~ur is removed from the gases as elemental sulfur.
Figure 5 shows in greater detail one alternative method for
the sulfating roasting of the roasting residues, using pyrite
for fulfilling the heat requirement and for creating a suitable
gas atmosphere in the roasting rurnace.
Roasting residue is produced in the sulfuric acid industry,
which roasts pyrite in order to obtain sulfur dioxide, the raw
material of sulfuxic acid production. By means of the present
invention, such roasting residues and impure pyrite can be
exploited effectively.
The invention is described below with the aid of an example.
Example
The basis for the calculations is roasting and leaching stages
7000 h/a and solution purification and metal recovery stages
7700 h/a. Pyrite concentrate was fed into a flash smelting
furnace at a rate of 82 t/h. Oil was added at 3.1 t/h, and
carbon was fed into the gas reduction zone at ll.S t/h.
Matte was obtained from the furnace at 55 t/h and fly dust
at 12 t/h. Elemental sulfur was obtained from the smelting
gases at 14 t/h.
The analysis of the pyrite was as follows:
S 47 %
Fe 42 %
Cu 1~15 %
Zn 3.2
Pb 1.3
As 0.5
Co 0.02
Ni 0.002
Cd 0~006
Ag 0.004
Au 0.0008
Hg 0.01
Se 0.002
The matte was granulated in water and ground so that 80 %
was ~ 74 ~m and 100 % ~ 290 ~m. The amount of sodium in the
grinding was 0.94 t/h. To the oxidizing roasting matte was fed
at 34.5 t/h, and to the sulfating roasting matte was fed at
20.5 t/h and dead-roasted product at 33.4 t/h. The calcine from
the sulfating roasting of matte, the analysis of calcine being
as follows:
S 2.63 %
Fe 60.6 %
Cu 1.69 %
Zn 1.66 %
Pb 0.28 %
Na 0O54 %
was leached ùsing washing water at 35.1 m3/h, and residue
was obtained at 49.9 t/ho
The sulfate solution was directed to the cementation of copper,
to ~hich scrap iron was added at 1 t/h. This stage yielded
90-percent cement copper at 0.91 t/h. The sulfate solution at
23.9 m3/h was fed to the iron removal stage, to which limestone
was added at 1.82 t/h and lime at 2.22 t/h. The amount of
return acid was 9.9 m3/h, water arrived from the S~L stage at
6.0 m3/h and washing water at 19.8 m3/h~
The yield of goethite-gypsum precipitate was 9.9 t/h, and the
sulfate solution, 53.0 m3/h, was directed to the metal hydroxide
precipitation stage. Calclum at 1.89 t/h was added -to this
stage, and solution arrived -, at 4.4 m3/h from the lead
sulfide-zinc sulfide leaching stage. The amount of hydroxide
precipitate was 6.1 t/h and the amount of Na2SO4 solution was
53.3 m /h.
.
The dusts from the flash smelting furnace, 12 t/h, were fed to
the sulfati.ng roasting, to which the Na2SO4 addition was 0.2
t/h. The calcine, 12 01 t/h, which contained
S 8.9 %
Fe21.9 %
Cu0.61 %
Zn:'.5.8 %
P~8.4 %
Na0.53 %
was leached in the water arriving from the iron precipitate
washing stage at 15.3 m3/h, and residue was obtained at 7.5
t/h, and it was directed to the strong acid leach, to which
return acid arrived at 14.8 m3/h, and the amoun-t of water
arriving at the residue washing stage was 13.3 m3/h. The yield
of residue was 6.9 t/l~, and it was directed to the chloride
leaching cycle.
The sulfate solution at 12.8 m3/h was directed to copper
cementation, to which scrap iron was fed at 0.076 t/h. 90-
percent cement copper was obtained from this stage at 0.052
t/h~ and sulfate so].ution was fed to the iron removal stage
at 11.6 m /h, and to this stage metal hydroxide precipitate
was fed from the leaching cycle for the roasted matte at
6.1 t/h, the lime addition was 0.08 t/h, and solution arrived
from the strong acid leach at 14.8 m3/h, as well as washing
water fr~,m the strong acid leach at 6 m3/h and iron residue
washing.water at 13.9 m3/h. The yield of goethite-gypsum
precipitate was 4.9 t/h, and the sulfate solution at 33.3 m3/h
12
was directed to the Cu-Co-Ni cementation s-tage, to which zinc
dust was fed at 78 kg/h, As2O3 at 2 kg/h, and the yield of
Cu-Co-Ni precipitate was 0.~7 t/h. Solution was fed to the
cadmium removal stage at 3~.3 m3/h and zinc dust at 3 k~/h. rrhe
yield of cadmium cementate was 3 k~/h. Sulfate solution entered
the evaporation stage at 33.3 m3/h, and the output was 24.5 m3/h
and was fed to the zinc electrolysis, from which zinc cathodes
were obtained at a rate of 2.5 t/h and return acid at 24.5 m3/h.
To the first stage of the chloride leaching cycle, the leaching
of roasted matte, roasted matté leach residue entered at 45.4
t/h, chloride solution at 80 m3/h, washing water at 20 m3/h, C12
at 0.29 t/h, lime at 0.46 t/h, and NaCi at 0.47 t/h. Fe203, a
suitable raw material for iron production,was obtained at 45.5
t/h, and it contained:
Fe 6/.5
Zn 0.12 %
Cu o.og
Pb ~.04 ~
S ~.2 %
and chloride solution at 100 m3/h was directed to the leaching
stage for dust calcine. Residue from the roasting of dust was
obtained at 6.3 t/h, and washing water at 14.8 m3/h and lirne at
0.41 t/h. Reiectable iron residue was obtained at 5.6 t/h and
chloride solution at 11.5 m3/h. ~ext, Cu, Ag and Au were cemented
out from the chloride solution. Zinc dust was used at 0~04 t/h~
an~ silver-gold-copper cementate was obtained at 0.01~ t/ho
Chloride solution entered the lead and zinc precipitation stage
at 117.2 rn /h and hydrogen sulfide was added at 0.31 t/h, the
yield of ~h'-, ZnS precipitate was ln 3 t/h and the arnount of
washing w~ter was 2.9 m3/h. The amount of outlet solution was
4.5 m3/h, and chloride solution passed to the evaporat:ion stage
at 112.7 m3/h, and its output was ~ m3/h. The PbSr ZnS
precipitate was leached in an autoclave, to which w~shing water
was fed at 4s5 m3/h. L~ach residue was obtained at 1.6 t/h, and
rhe zinc sul~ate solution, 4.4 m3/h, was directed to the metal
hydroxide precipitation stage of the matte cycle. The leach
residue ~-as (~irected to the lead carbonate precipitation stage,
t:o which re~idual sodium s~llfate solution from the rnatte
cycle arrived at 5 m3/h, and the Na2C03 addition was 0.6 t/h
and the amount of washing water 4~5 m3/h. The sodium sulfate
solution, 9 m3/h, was direc-ted -to the grinding of matte, and
lead carbonate was obtained at 1.4 t/h, which was calcinated
to lead oxide, the amount of which was 1.16 t/h, and the lead
oxide was reduced to metallic lead, the yield being 1.0 t/h.
