U.S. patent number 3,728,430 [Application Number 05/097,851] was granted by the patent office on 1973-04-17 for method for processing copper values.
This patent grant is currently assigned to The Anlin Company of New Jersey. Invention is credited to Jay B. Clitheroe, Forrest H. Lacy, Jr..
United States Patent |
3,728,430 |
Clitheroe , et al. |
April 17, 1973 |
METHOD FOR PROCESSING COPPER VALUES
Abstract
The steps of leaching copper values from oxide ores, mixed
sulfide-oxide ores, and other oxide copper bearing materials in an
aqueous medium or suspension containing soluble sulfites or
bisulfites and precipitating the copper as copper sulfide in an
aqueous medium or suspension containing elemental sulfur and
soluble sulfites or bisulfites. The precipitated copper, along with
any other copper present as elemental copper or copper sulfide, is
then separated by conventional methods.
Inventors: |
Clitheroe; Jay B. (Salt Lake
City, UT), Lacy, Jr.; Forrest H. (Houston, TX) |
Assignee: |
The Anlin Company of New Jersey
(Houston, TX)
|
Family
ID: |
22265437 |
Appl.
No.: |
05/097,851 |
Filed: |
December 14, 1970 |
Current U.S.
Class: |
423/26; 423/27;
423/43; 299/5; 423/37 |
Current CPC
Class: |
C01G
3/12 (20130101); B03B 1/04 (20130101) |
Current International
Class: |
C01G
3/12 (20060101); B03B 1/00 (20060101); B03B
1/04 (20060101); C01g 003/12 (); C22b 015/08 () |
Field of
Search: |
;75/117,11R,108,2
;23/135 ;299/5 ;423/26,27,37,43 |
References Cited
[Referenced By]
U.S. Patent Documents
Primary Examiner: Ozaki; G. T.
Claims
What is claimed is:
1. In a method for processing oxide ores, mixed sulfide-oxide ores
and middlings, dumps and slag containing oxide copper values, the
combination of steps comprising:
leaching and substantially simultaneously precipitating said oxide
copper values in said material at a pH of about 1.0 to about 9.4
and at a temperature of from about 80.degree. F to about
212.degree. F in an aqueous pulp with elemental sulfur and a
material selected from the group consisting of soluble sulfites and
soluble bisulfites, whereby said copper is precipitated as copper
sulfide; and,
separating copper sulfide from said pulp by froth flotation.
2. The method of claim 1 wherein:
said leaching and precipitating step is carried out in the presence
of a stoichiometric excess of sulfur dioxide and elemental sulfur
in relation to the amount of oxide copper in said copper containing
material.
3. The method of claim 1 wherein:
said leaching and precipitating step includes the additional step
of adding an acid selected from the group consisting of sulfuric,
hydrochloric, nitric, hydroiodic, hydrofluoric, fluosilicic,
phosphoric, acetic and formic to said pulp prior to completion of
the precipitating portion of said leaching and precipitating step
in an amount sufficient to maintain the pH thereof in the range of
from about 1.0 to about 9.4.
4. The method of claim 1 wherein:
said leaching and precipitating step is carried out in said pulp in
the presence of a soluble chloride.
Description
BACKGROUND OF THE INVENTION
1. Field of the Invention
The field of the invention is methods for
hydrometallurgical-physical beneficiation of metalliferous ores,
mixed sulfide-oxide ores, and middlings, slags, tailings, and dumps
containing oxide copper values.
2. Description of the Prior Art
Ores of sufficiently high copper content may be smelted directly,
but lower grade copper ores, dumps, middlings, and slag usually
must be concentrated before shipment or smelting in order to
economically obtain the copper values therefrom.
Ores high in copper sulfide content quite often contain a portion
of "oxide" ores intermingled in the naturally occurring deposits.
The term "oxide" is meant to describe the nonsulfide ore
constituents, e.g., copper oxides, silicates, carbonates,
hydroxides, and sulfates. Rich sulfide ores are commonly
concentrated by grinding and flotation processes which, in many
cases, allow copper "oxides" to pass substantially unchanged
through the process to waste.
Certain oxide-containing ores are presently being concentrated by
leaching with sulfuric acid and precipitating copper or copper
sulfide by the use of ground or powdered iron or shredded scrap or
hydrogen sulfide. In many cases these steps are followed with
separation of the precipitate of cement copper or copper sulfide by
conventional froth flotation. Such a technique is familiarly known
as "LPF." Conventional LPF has many disadvantages, including the
following:
A. The reagents used in conventional LPF are relatively
expensive.
B. Some of the iron is consumed by the required free acid and by
oxidation and thus wasted.
C. Some of the soluble copper tends to plate out on coarser
particles of iron, thus wasting iron and sometimes rendering the
coarse particles of plated copper non-floatable.
D. Copper sulfide precipitated by hydrogen sulfide tends to be
colloidal in nature and therefore difficult to concentrate by
flotation, especially in the presence of slimes. In addition, the
use of an extremely toxic substance such as hydrogen sulfide poses
many safety engineering problems.
E. The sulfuric acid leach is often slow.
F. Acid consuming gangue often requires excessive quantities of
relatively expensive sulfuric acid.
Even though conventional LPF has been widely used, copper oxide
wastes and many copper oxide ores, some in relatively large
deposits, have been difficult or impossible to process
economically. This is especially true of those ores containing a
large proportion of slimes, copper silicates or acid-consuming
constituents.
BRIEF DESCRIPTION OF THE DRAWING
The drawing is a flow sheet showing one embodiment of the invention
preferred by the inventor for processing materials containing oxide
copper .
SUMMARY OF THE INVENTION
This invention improves oxide copper beneficiation processes and
includes the steps of leaching the oxide copper values from
metalliferous materials in the presence of water and soluble
sulfites or bisulfites and precipitating the copper values as
copper sulfide in the presence of elemental sulfur and soluble
sulfites or soluble bisulfites. The method may include the
additional step of separating the dissolved copper from the
undissolved ore prior to the precipitation of copper sulfide. After
precipitation, the copper sulfide is then separated from the
aqueous medium.
In the preferred embodiment of the invention, the leaching and
precipitating steps are carried out substantially simultaneously in
a pulp containing soluble sulfites or soluble bisulfites and
elemental sulfur. The precipitate is then separated from the pulp,
preferably by conventional froth flotation.
Another embodiment of the invention includes the steps of leaching
the oxide copper containing material in situ with a solution
containing soluble sulfites or bisulfites, transferring the leached
solution containing copper values to a precipitation vessel, and
precipitating the copper values as copper sulfide in the presence
of soluble sulfites and bisulfites and elemental sulfur.
Certain embodiments of the invention include the step of leaching
the oxide copper in the presence of water, soluble chloride, and a
material selected from the group consisting of soluble bisulfites
and sulfur dioxide.
DESCRIPTION OF THE PREFERRED EMBODIMENT
The chemistry of the embodiment preferred by the inventor may be
schematically represented as follows:
Cu X + H.sub.2 SO.sub.3 .fwdarw. Cu.sup.+.sup.+ + SO.sub.3
.sup.-.sup.- + H.sub.2 X (1) SO.sub.2 + H.sub.2 O .fwdarw. 2H.sup.+
+ SO.sub.3 .sup.-.sup.- (2)
H.sub.2 O + Cu.sup.+.sup.+ + SO.sub.3 .sup.-.sup.- + S.degree.
.fwdarw. Cu S .uparw. + H.sub.2 SO.sub.4 (3)
or overall:
Cu X + 2H.sub.2 O + SO.sub.2 + S.degree. .fwdarw. Cu S .uparw. +
H.sub.2 SO.sub.4 + H.sub.2 X
This embodiment of the invention is practiced by slurrying the
copper-containing material with water and reducing the material to
a suitable degree of fineness before entering the
leach-precipitation circuit. However, any method of exposing the
copper values of the raw material to the leaching-precipitation
medium is acceptable. The slurry thus formed is referred to as
"pulp." Elemental sulfur and water may be added prior to or after
the grinding operation, and if added prior to grinding, the sulfur
would, of course, be ground along with the raw material. The water
need not be fresh water. It may be salt water and, under certain
circumstances, more desirably may be salt water. Salt water is
defined as water which naturally contains sodium chloride or water
to which sodium chloride has been added. Such elemental sulfur may
be introduced into the leach-precipitation circuit at any point
prior to the precipitation step. From the grnding operation the
pulp flows to leach-precipitation agitators. Soluble sulfite
furnished in the form of hot gas containing sulfur dioxide is
contacted with the pulp, both heating the mixture and furnishing
sulfur dioxide for sulfurous acid. In the drawing, a series of
leach-precipitation agitators are shown for contacting the pulp and
hot SO.sub.2 - bearing gas. Soluble sulfates may be added to the
leach-precipitation agiator as accelerators. Oxide copper is
leached (as shown in equation [1]) by the sulfurous acid formed by
hydrolysis of the sulfur dioxide bubbled through the slurry (as
shown in equation [2]). A portion of the elemental sulfur in the
pulp serves as the precipitant, yielding a precipitate of copper
sulfide (as shown in equation [3]). It has been found beneficial to
include in the process an excess of elemental sulfur above that
amount which stoichiometrically reacts with the leached copper.
Some of the copper sulfide clings to excess free sulfur particles
present during the precipitation step. As may be seen from equation
(3), sulfuric acid is produced in the leach-precipitation reaction
and is available to react with all or a part of any acid-consuming
constituents in the ore, thereby lowering the chemical requirements
if the gangue is acid consuming.
It is believed that a preponderance of the copper, represented by
the Cu.sup.+.sup.+ symbol in equations (1)-(3), is present in the
precipitate as the cupric form. However, there may be some copper
present as cuprous sulfide after precipitation, depending on the
specific ore being processed. It is of no consequence to the
operation of the invention which valence state the copper
takes.
The elemental sulfur should be ground to, or supplied in, a fine
particle size, because its reaction rate is somewhat dependent upon
its available surface area. The preferred method for introducing
the sulfur is to grind it directly along with the ore being
processed. The source of sulfur may be in a number of forms, e.g.,
solid elemental sulfur, molten sulfur sprayed onto the ore, or even
naturally-occurring sulfur ores, many of which such sulfur ores are
not otherwise economically processable. The elemental sulfur for
the precipitation could be supplied by a reduction roast of
pyrite.
The precipitate is then physically separated, and in this preferred
embodiment, is floated by conventional techniques. Many of the
particles of copper sulfide surround and adhere to one or more
particles of elemental sulfur and thus are easily floated and
cleaned. The precipitate of copper sulfide was found to be coarse
and surprisingly easy to float from the pulp. No colloidal
precipitate, as found in H.sub.2 S precipitations, or adherence to
iron particles, as in iron precipitations, was noted. The ease of
recovery by flotation contributes to both the economy of the
process and its suitability for marginal ores.
Other materials containing copper are also floated with the copper
sulfide precipitate in this embodiment. Most of the excess sulfur
not consumed in the process is recovered in the flotation
concentrate, the flotation being carried out after the pulp-to-pulp
heat exchange step shown in the drawing. Naturally occurring copper
sulfide, as well as native copper, if present, are floated with the
precipitate and recovered in the froth flotation step. It has been
discovered that the process brightens (detarnishes) copper sulfide
and native copper particles originally present in the ore, thereby
improving their flotation characteristics. Some native copper, if
present, may be leached and precipitated in the leach-precipitation
agitator. This leached and precipitated native copper will appear
along with leached and precipitated oxide copper as copper
sulfide.
The concentrate, consisting primarily of element sulfur and copper
sulfide, is dewatered and roasted to yield heat and sulfur dioxide
for circulation to the leach-precipitation agitators. The roaster
may be operated so that only the elemental sulfur in the
concentrate is oxidized or it may be operated so that the copper
sulfide in the concentrate is also oxidized, yielding an excellent
dried concentrate which may be transferred to a smelter or
refinery.
The leach-precipitation agitators may be provided with a vent to
discharge any excess gaseous sulfur dioxide at an altitude which
will comply with any air pollution regulations, but preferably are
provided with a conduit to transfer any excess gaseous sulfur
dioxide to a basic metallic earth contact stage. Such a contactor
would contain a basic metallic earth material such as sodium or
potassium oxide, hydroxide, or carbonate, calcium or magnesium
oxide, hydroxide, or carbonate, ammonium hydroxide or carbonate, or
ferrous hydroxide, oxide, or carbonate. Such a contactor could also
contain sodium or potassium chloride, calcium or magnesium
chloride, ammonium chloride, and ferrous or ferric chloride.
Naturally occurring contactor materials could be dolomite, calcite,
soda ash, or limestone or ordinary salt. There should be present in
any such contact stage some water, preferably that which would
later be used in the grinding and the leach-precipitation steps,
all or a portion of which may come from the tailings pond. Such a
contact stage can provide a source of soluble fixed sulfites and
bisulfites or chloride to the process, all of which have been
demonstrated to be beneficial. In addition, utilization of the
waste gases in this manner is beneficial not only as a source of
soluble fixed sulfites and bisulfites and chlorides, but if the
materials in the contactor are properly proportioned, this
virtually eliminates the problem of venting any sulfur dioxide to
the atmosphere, a pollution problem which has plagued various metal
processing installations in the past. Soluble fixed sulfites and
bisulfites may be defined as any source of soluble sulfite or
bisulfite ions other than sulfurous acid. Sulfurous acid would be
included in the group defined as soluble sulfites and soluble
bisulfites, which group also includes the sulfites and bisulfites
of sodium, calcium, ammonium, magnesium, and potassium, as well as
the sulfite of iron. Soluble chlorides are defined as any water
soluble chloride-containing material which ionizes to furnish the
chloride in water solution.
Some or all of the heating and all or a part of the sulfur dioxide
for leaching could be supplied by utilizing hot smelter gases
containing sulfur dioxide or by other sulfur dioxide bearing stack
or vent gases. Heat and sulfur dioxide also could be supplied
simply by burning elemental sulfur with or without the roasting of
sulfur-bearing flotation concentrates. Soluble sulfite or
bisulfite, in the form of sulfur dioxide, for leaching could be
supplied by an oxidation roast of the FeS resulting from a
reduction roast of pyrite or could,of course, be supplied as a
solution of sulfurous acid.
It should be noted that in the preferred embodiment sulfur dioxide
introduced into the process from a roaster gas, smelter gas or
other sulfur dioxide bearing vent or stack gas does not have to be
in a clean gas stream. Such gas may contain substantial amounts of
normally objectionable particulate matter (including copper or its
compounds) and may even contain normally harmful or objectionable
elements, such as tellurium, selenium, etc. In the preferred
embodiment such materials end up either as a part of the recovered
concentrate, in the case of most of the copper, or as harmlessly
fixed materials in the tailings. Thus, both air and water pollution
is minimized or eliminated.
The process may be carried out at a temperature ranging from about
80.degree. to 212.degree. F. at atmospheric pressure and at a pH of
1.0 to 9.4. However, a temperature of about 130.degree. F to about
170.degree. F. and a pH of about 1.0 to about 6.0 has been
demonstrated to be practical, economical and desirable.
With the addition of pressure equipment at the leach-precipitation
stage, the process may be carried out above boiling. A convenient
source of heat to supply such elevated heat requirements would make
higher temperature extraction and precipitation with pressure
equipment attractive as an alternative embodiment of the
process.
A part of the heat requirements may be efficiently achieved by
indirect heat exchange of the final leach pulp with the pulp
entering the leach contactor. The sulfur dioxide and all or a part
of the heat requirements are obtained most easily by burning sulfur
directly or roasting the copper sulfide concentrate which will
normally contain some free sulfur. In some cases, roasting the
concentrate will not provide enough heat or enough sulfur dioxide
to reach optimum reaction conditions. In that case, additional
sulfur, natural gas or fuel oil may be burned to supply additional
heat, and, if sulfur is burned, additional sulfur dioxide. The
amount of sulfur dioxide and heat supplied by the hot gas stream
bubbled into the leach-precipitation agitators may vary with the
ore being processed.
In the case of an ore containing an oxide mineral that is somewhat
refractory with respect to sulfurous acid, the low-cost sulfurous
acid could be used in a primary leach-precipitation step to digest
and precipitate the easily dissolved metal minerals, while being
assisted to leach and precipitate such refractory minerals as a
secondary leaching step by the addition of an economically small
quantity of more valuable mineral acid, such as sulfuric,
hydrochloric, hydroflouric, nitric, hydroiodic, fluosilicic,
phosphoric, acetic and formic. Additionally, it has been
demonstrated that the addition of a soluble chloride in the
presence of sulfur dioxide or a soluble bisulfite aids in the
leaching portion of the process. The amount of such acid or soluble
chloride added or used will vary with the particular ore being
processed, and the acid, if added, should be present in an amount
sufficient to keep the leach-precipitation reaction within optimum
parameters. Both heat and low pH contribute to shorter
leach-precipitation periods and are, therefore, desirable
economically.
It has been demonstrated that the hot pulp treatment as described
above is very beneficial to flotation and that it causes, among
other things:
a. A rapid reaction of copper sulfide with collectors.
b. A partial dehydration of argillaceous slimes, reducing their
ion-exchange characteristics and mineral coating characteristics,
thereby improving flotation and causing a cleaner froth.
c. The brightening (detarnishing) of copper sulfide minerals and
native copper originally present, making such material more readily
reactive with collectors.
d. The copper sulfide precipitate by elemental sulfur according to
this embodiment is coarsely crystalline, making an excellent
flotation feed, rather than being colloidal as is usually the case
with hydrogen sulfide precipitation.
The following examples illustrate the process as hereinbefore
disclosed:
EXAMPLE 1
Red Bed copper ore from North Texas, containing about 85 percent
illite clays and assaying approximately 1.1 percent copper, of
which 67 percent was in "oxide" form, and containing the equivalent
of 1.5 percent calcium carbonate, was put into a ball mill with the
equivalent of 15 pounds per ton of elemental sulfur and wet ground
with water to approximately minus 200 mesh at about a 45 percent
solids level. All percentages and ratios are calculated on weight
basis. The finely ground elemental sulfur and copper ore were mixed
with the equivalent of 30 pounds sulfur dioxide per ton of ore in
the form of sulfurous acid and agitated for 20 minutes at a pH of
2.6-2.9 at 138.degree. F. Conventional froth flotation was then
carried out. Copper sulfide, some of which adhered to particles of
elemental sulfur, was then recovered by conventional froth
flotation. The cleaner concentrate contained 76.95 percent of the
copper originally in the ore with 7.51 percent of the copper
remaining in the cleaner tail and 15.54 percent of the copper
remaining in the rougher tail.
As a comparison, a different sample of the same ore was put into a
ball mill and ground to minus 100 mesh with water at a percent
solids level of approximately 50 percent. It was then processed by
conventional LPF techniques, e.g., leached for 40 minutes with the
equivalent of 80 pounds of sulfuric acid per ton of ore at a pH of
1.6, precipitated for 20 minutes with the equivalent of 80 pounds
of powdered iron per ton of ore (of which 27.3 pounds of iron was
consumed) and then treated for 20 minutes by conventional froth
flotation with two stages of recleaning. The result of this test
showed the distribution of the copper to be 65.4 percent in the LPF
concentrate, 15.8 percent in the cleaner tail, 5.2 percent in the
recleaner tailing, and 13.6 percent remaining in the rougher tail.
This test represented the best results obtained from an extensive
testing program of conventional LPF as applied to this ore.
EXAMPLE 2
The Red Bed copper ore of Example 1 was wet ground with water and
elemental sulfur in an amount equivalent to 30 pounds per ton of
ore and digested for 20 minutes at a temperature of 146.degree. F.
with the addition of the equivalent of ten pounds of ammonium
sulfate per ton of ore in the leach-precipitation agitator. During
this test the equivalent of 80 pounds per ton of sulfur dioxide was
added to the leach-precipitation agitator. The pH of the
leach-precipitation slurry remained at 2.7 during
leaching-precipitating. The cleaner concentrate showed a recovery
of 88.31 percent of the total copper originally in the ore with
2.89 percent in the cleaner tail and 8.80 percent in the rougher
tail.
EXAMPLE 3
The Red Bed copper ore of Examples 1 and 2 with the equivalent of
30 pounds elemental sulfur per ton of ore was ground to about a
minus 200 mesh with water at a percent solids level of
approximately 33 percent. The resulting slurry was mixed with the
equivalent of 30 pounds of sulfur dioxide per ton of ore and
digested for 20 minutes at 160.degree. F. The
leaching-precipitating slurry contained the equivalent of 50 pounds
of soluble sulfite salts per ton of ore. The ending pH of the
slurry was 5.6. The recovery by conventional flotation was 94.21
percent of the copper contained in the original ore sample with 0.5
percent of the copper remaining in the cleaner tail and 5.12
percent remaining in the rougher tail.
EXAMPLE 4
The Red Bed copper ore or examples 1 through 3, treated with the
equivalent of 30 pounds elemental sulfur per tom of ore, was ground
to about minus 200 mesh with water at a percent solids level of
approximately 33 percent. The resulting slurry was mixed with the
equivalent of 30 pounds of sulfur dioxide per ton of ore and
digested for 20 minutes at 80.degree. F., the approximate ambient
temperature. The leaching-precipitation slurry contained the
equivalent of 50 pounds of soluble sulfite salts per ton of ore.
The beginning pH of the leach-precipitation slurry was 3.1, and the
ending pH was 4.2. The recovery by conventional flotation was 67.32
percent of the copper contained in the original ore sample with
2.33 percent of the copper remaining in the cleaner tail and 30.35
percent remaining in the rougher tail.
EXAMPLE 5
The Red Bed copper ore of examples 1 through 4 was ground with the
equivalent of 30 pounds of elemental sulfur per ton of ore to about
minus 200 mesh with water at a percent solids level of
approximately 33 percent. The resulting slurry was mixed with the
equivalent of 110 pounds of soluble fixed sulfite salts per ton of
ore and digested for 20 minutes at 160.degree. F. The beginning pH
of the slurry was 8.4 and the ending pH was 9.4. The recovery by
conventional flotation was 45.03 percent of the copper contained in
the original ore sample with 1.94 percent of the copper remaining
in the cleaner tail and 53.03 percent remaining in the rougher
tail.
EXAMPLE 6
The Red Bed copper ore of examples 1 through 5 was ground with the
equivalent of 22 pounds of elemental sulfur per ton of ore to about
minus 200 mesh with water at a percent solids level of
approximately 33 percent. The resulting slurry was mixed with 40
pounds of sulfur dioxide per ton of ore and digested for 20 minutes
at 160.degree. F. with the equivalent of ten pounds per ton of
ammonium sulfate. The slurry also contained the equivalent of 73
pounds of sodium chloride per ton of ore. The recovery by
conventional flotation was 96.20 percent of the copper contained in
the original ore sample with 0.58 percent of the copper remaining
in the cleaner tail and 3.22 percent remaining in the rougher
tail.
Other examples demonstrating the efficacy of this embodiment of the
invention have been performed over a pH range of about 1.0 to about
9.4, at temperatures from 80.degree. F. to about 212.degree. F.,
and at elemental sulfur to copper molar ratios from 0.5 to 4.0.
Sulfur dioxide to copper molar ratios between 0.5 and 4.5, and
soluble fixed sulfite and bisulfite to copper molar ratios between
0.5 and 4.5 have been used in other examples and shown to be
operative.
EXAMPLE 7
Uinta Mountain sandstone ore from Utah containing azurite and
malachite with a copper content of 2.85 percent and a siliceous
matrix was wet ground with the equivalent of 80 pounds sulfur per
ton of ore and water to about minus 65 mesh at a pulp density of
39.2 percent solids. Eighty pounds of sulfur dioxide per ton of ore
and 50 pounds per ton of soluble fixed sulfite salts were digested
with the slurry at 170.degree. F. for 20 minutes at a pH of 1.5 to
1.7. The concentrate obtained by conventional froth flotation in
this example contained 93.05 percent of the total copper in the ore
sample with the cleaner tail containing 0.54 percent and a rougher
tail containing 6.41 percent of the total copper.
EXAMPLE 8
A conglomerate oxide ore from the Philippines, with a 2.59 percent
copper content and containing chrysocolla and tenorite in a
siliceous matrix, was ground with the equivalent of 104 pounds of
sulfur per ton of ore and water to a minus 65 mesh. The equivalent
of 10 pounds of ammonium sulfate per ton of ore and 77 pounds of
sulfur dioxide per ton of ore were digested for 60 minutes at a
temperature of 160.degree. F. at a pH of about 1.6. The
leach-precipitation slurry additionally contained the equivalent of
100 pounds of soluble fixed sulfite salts per ton of ore. In this
example, a seconary leach-precipitation step was conducted,
consisting of adding the equivalent of 50 pounds of sulfuric acid
per ton of ore to the leach-precipitator and
digesting-precipitating for an additional 20 minutes at a
temperature of 160.degree. F. and a pH of 1.1 to 1.6. The
concentrate obtained using conventional froth flotation contained
91.06 percent of the copper in the original ore, with a cleaner
tail containing 0.54 percent and a rougher tail containing 7.95
percent of the original copper.
A comparison test for leaching only as performed in the prior art,
showed that this ore ground to the same fineness yielded only 85
percent of its total copper in a 5 percent sulfuric acid solution
after four hours of conventional leaching.
EXAMPLE 9
An Arizona Strip oxide ore containing quartz, approximately 20
percent calcium carbonate, azurite, and malachite, for conventional
leaching, required 500 pounds of H.sub.2 SO.sub.4 per ton of ore
yielding 85 percent copper extraction in the leaching stage only,
without precipitation, in one hour on minus 65 mesh material. The
85 percent extraction level should not be confused with a recovery
level of the same percentage, because additional losses, as well as
the consumption of additional chemical reagents such as iron or
hydrogen sulfide, would be experienced with conventional recovery
techniques. The same Arizona ore was treated in accordance with the
preferred method by grinding with the equivalent of 117 pounds of
elemental sulfur per ton of ore and water to minus 65 mesh. The
resulting slurry was then treated with the equivalent of 348 pounds
of sulfur dioxide per ton of ore and the equivalent of 50 pounds of
soluble fixed sulfite salts per ton of ore for 40 minutes at
160.degree. F. at a pH of 3.8. The concentrate obtained using
conventional froth flotation contained 86.29 percent of the copper
originally in the ore, with a cleaner tail containing 0.56 percent
and a rougher tail containing 13.12 percent of the original
copper.
As an illustration of processing cost savings affected by the
preferred embodiment, reagent costs for processing the ore of this
example are less than half of the cost of leaching only by the
methods of the prior art.
The concentration of various reagents used in the process will
vary, of course, depending on the specific ore being processed.
Maximum efficiency of the process may be achieved by varying the
time, temperature and pH of the leaching, precipitating and
separating steps in accordance with generally known scientific
principles. For example, a high-silica or high-lime ore may require
the lowering of the pH of the leaching-precipitating medium or
increasing the reaction temperature for optimum results.
In addition to adaptation of the described embodiments for normal
ore processing techniques, one embodiment of the invention includes
working over subterranean deposits by leaching ores in situ. Heat
and pressure present at such underground mines might be utilized
containing the leaching process. The pregnant liquor containing the
metal values may then be pumped from the ore location to
precipitating and floating machinery at a plant.
From the foregoing description it can be seen that a practical and
economic processing technique for copper has been shown. Further
modifications and alternative embodiments of the invention will be
apparent to those skilled in the art in view of this description.
Accordingly, this description is to be construed as illustrative
only and is for the purpose of teaching those skilled in the art
the manner of carrying out the invention. It is to be understood
that the form of the invention herewith shown and described is to
be taken as the presently preferred embodiment. For example,
equivalent sequences of steps and process configurations may be
substituted for those illustrated and described herein, and certain
features of the invention may be utilized independently of the use
of other features, all as would be apparent to one skilled in the
art after having the benefit of this description of the
invention.
* * * * *