United States Patent [19]
Mason et ale
[11] Patent Number:
[45] Date of Patent:
4,552,589
Nov. 12, 1985
[54] PROCESS FOR THE RECOVERY OF GOLD
FROM REFRACTORY ORES BY PRESSURE
OXIDATION
[57] ABSTRACT
2,315,187 3/1943 Chapman et al. 75/106 X
2,869,529 1/1959 Forward et al. 75/118 R
3,935,006 1/1976 Fischer 75/118 R
4,038,362 7/1977 Guay 423/40
4,053,305 11/1977 Smyres et al. 75/118 R X
4,188,208 2/1980 Guay 423/29 X
4,259,107 3/1981 Guay 423/29 X
4,267,069 5/1981 Davidson et al 75/118 R X
4,289,532 9/1981 Matson et al. 423/29 X
Primary Examiner-Andrew H. Metz
Assistant Examiner-Nam X. Nguyen
Attorney, Agent, or Firm-Lyon & Lyon
A process for recovering precious metals from a refractory
ore by forming a slurry which is heated to a temperature
in excess of about 1500 C. at an oxygen partial
pressure in excess of about 10 psia for an effective
amount of time to form a partially oxidized slurry which
is then subjected to carbon-in-Ieach treatment to separate
the precious metals.
27 Claims, 2 Drawing Figures
AppI. No.: 574,667
Filed: Jan. 27, 1984
Int. Cl.4 C22B 11/08
U.S. Cl 75/105; 75/118 R;
423/29; 423/27
Field of Search 75/105, 118 R; 423/23,
423/27, 29, 30, 31
References Cited
U.S. PATENT DOCUMENTS
2,147,009 2/1939 Chapman 75/106 X
[75] Inventors: Peter G. Mason, Scarborough,
Canada; Frank D. Wicks, Stansbury
Park, Calif.; John C. Gathje, Arvada,
Colo.
Assignee: Getty Oil Company, Salt Lake City,
Utah
[21]
[22]
[51]
[52]
[58]
[56]
[73]
9·08 r----r---r--,...--..,--..,---,---,---,--....,
z o.err
~
~ 0.06
...J o
(!) 0.05
l:i
en 0.04
...J
<:(
I-' 0.03
IJJ
°Z00·2
~
U 0.01
-<>- 50psig 02 COMPOSITE I
--0- 100 psig 02 COMPOSITE I
-&- 25 psig 02 COMPOSITE DC-2
-<>- 25 psig 02 COMPOSITE 2
0.00 L....._.L..-_.L..-_..L-_...L...._...L...._....L..._...L_....L._--l
100 120 140 J60 180 200 220 240 260 280
TEMPERATURE. °C
u.s. Patent Nov. 12, 1985 Sheet 1 of2 4,552,589
0.08 --~--.----.-----.-~~----,.----.------
-0- 50 psig 02 COMPOSITE I
-0- 100 psig 02 COMPOSITE I
-&- 25 psig 02 COMPOSITE DC-2
-0- 25 psig 02 COMPOSITE 2
Z 0.07 g
d 0.06
...J o
<.9 0.05
~
en 0.04
...J
«t-= 0.03
wo
0.02
Z
~
U 0.01
0.00 L..--_'""'--_......L..-_.....L-_--J-_---L._--'-_---I__'------I
100 120 140 160 180 200 220 240 260 280
TEMPERATURE, °C
.IIG.l.
u.s. Patent Nov. 12, 1985 Sheet 2 of2 4,552,589
0.12
0.11
0.10
Z009 o'
~
"00
.08
o-'
<.9 0.07
~
.. 0.06
~
~ O.co
wo 0.04
Z
~
U 0.03
0.02
0.01
-0- COMPOSITE I
-D- COMPOSITE DC-2
--t:::r- COMPOSITE DC-I
-0-- COMPOSITE DC-I
0.00 l....-_.L...-_.L--_.L--_""--_""--_""--_""--_-L-------J
o 10 20 30 40 50 60 70 80 90
TIME, MINUTES
~.2.
2
In an attempt to overcome some of these problems, the
United States Bureau of Mines has conducted experiments
in which they used a wide variety of oxidation
pretreatment systems including ozone, sodium hypo-
5 chlorite, calcium hypochlorite, permanganates, perchlorates,
chlorates and oxygen. As a result of these and
other experiments, various processes dealing with pretreatment
systems have been patented, including U.S.
Pat. No. 1,461,807, which discloses the use of certain
mineral oils for "blinding" the action of the carbonaceous
impurities on the cyanide complex formation;
U.S. Pat. No. 2,234,140, which discloses that certain
wetting agents can make the ore more amenable to
cyanidation; U.S. Pat. No. 3,639,925, which discloses
the use of sodium hypochlorite and calcium hypochlorite
as agents for oxidizing the carbonaceous materials
so as to prevent them from adsorbing the gold cyanide;
U.S. Pat. No. 3,846,124, which discloses a chlorine
pretreatment of the ore in the absence of any alkaline
material in order to decompose the organic carbonaceous
components and remove them prior to cyanidation;
U.S. Pat. No. 3,574,600, which discloses that certain
acids can be used in conjunction with an ozone
treatment prior to cyanidation to oxidize the carbonaceous
impurities; and U.S. Pat. No. 4,038,362, which
discloses a pre-oxidation technique carried out in the
absence of extraneous alkaline material for reducing the
amount of chlorine needed to pretreat the ore.
Another process utilizing air oxidation followed by
chlorination prior to cyanidation is disclosed by W. J.
Guay in the article "How Carlin Treats Gold Ores by
Double Oxidation" in World Mining, March, 1980,
pages 47-49. In this process, an ore bearing gold of
which a substantial portion of the ore was not amenable
to standard cyanidation techniques because of the presence
of activated carbon in pyrite is disbursed in an
aqueous slurry of ground ore at 40-50% solids and
temperatures of 800 _860 C. until a considerable portion
of the pyrite is oxidized to iron oxides. Although some
of the carbonaceous materials are decomposed by the
oxidation, it was found necessary to follow the air oxidation
with chlorination in order to complete the oxidation
of the carbonaceous materials and pyrite. This
process also utilized the addition of a solution of sodium
carbonate over a period of several hours during the air
oxygenation to react the air and the soda ash with pyrite
to form soluble sulfates and iron oxides. The soda ash is
added to the slurry in amounts varying from 25-100
pounds per ton of ore.
In addition to various processes for the pretreatment
of carbonaceous ore, other processes have been disclosed
which involve modifications to the actual cyanidation
process. Thus, U.S. Pat. Nos. 2,147,009 and
2,315,187 disclose the use of finely divided charcoal
during cyanidation to simultaneously leach the gold
from the ore and absorb the gold on the charcoal so as
to maintain the solution continuously depleted of gold
and thereby improve cyanidation efficiency. More re-
60 cently, the prior processes utilizing finely divided charcoal
have been improved by replacing the finely divided
charcoal with granular activated charcoal. Such a
process, which is commonly known in the art as "carbon-
in-Ieach," is disclosed in U.S. Pat. No. 4,289,532,
the disclosure of which is hereby specifically incorporated
by reference, in which a slurry is initially treated
with an oxygen-containing gas for at least an hour and
contacted with the source of hypochlorite ions for at
4,552,589
1
PROCESS FOR THE RECOVERY OF GOLD FROM
REFRACfORY ORES BY PRESSURE OXIDATION
BACKGROUND OF THE INVENTION
The field of the present invention relates generally to
a process for recovering gold or other precious metals
from refractory ores by pressure oxidation followed by
carbon-in-leach treatment.
Gold and other precious metals naturally occur in 10
ores in several different forms and complexes. Unfortunately,
however, the gold-bearing ores found in Nevada,
Utah, California and other states in the United
States, as well as other countries throughout the world,
often contain refractory material which interferes with 15
the extraction of gold and other precious metals. Further,
the actual gold content of such ores is variable and
an ore with a relatively small actual gold content can
have less than one-tenth of an ounce of gold per ton of
mined ore. When an ore with a relatively small actual 20
gold content is processed, the adverse effects of refractory
material may make the recovery of gold prohibitive
unless an effective process is utilized for coping
with the refractory material.
In the prior art, many processes are disclosed for 25
treating "sedimentary carbonaceous gold-bearing ores"
or refractory ores. However, these processes rarely
define the actual constituents posing the problem to the
particular process for a particular ore. Thus, as noted by
W. J. Guay and M. A. Gross in Preprint 81-34 of the 30
Society of Mining Engineers of AIME, the term "carbonaceous"
has been rather loosely applied to ore constituents
of widely varying characteristics, including:
(1) an activated carbon component capable ofadsorbing
gold chloride or gold chloride complexes from solu- 35
tions; (2) a mixture of high molecular weight hydrocarbons
usually associated with the activated carbon components;
and (3) an organic acid, similar to "humic
acid", containing functional groups capable of interacting
with gold complexes to form organic gold com- 40
pounds. While the structure of such compounds is unknown,
it is possible that they are formed by chelation
wherein ligands, such as N, S or 0, in organic acids
form stable gold chelates.
In addition to the arbitrary nature of any definition of 45
the term "carbonaceous", the term "refractory" has
also been the subject of a rather unsettling, all-inclusive,
vague definition. Thus, loosely speaking, the term "refractory"
has been used to define an ore containing any
substance which interferes with the recovery of gold 50
from said ore by standard cyanidation techniques. One
such substance which is commonly known to render an
ore refractory is pyrite, which may occlude finely disseminated
particles of gold in spheroidal or cubic clusters.
Such occluded gold particles may have a size of 55
less than .02 microns. Other materials which may be
deemed to render an ore refractory, besides the general
class of carbonaceous materials, include clay minerals
which can adsorb the gold cyanide complex and certain
sulfur bearing compounds other than pyrite.
In the past, extensive research has been conducted
into methods for dealing with the problem of carbonaceous
impurities in gold-containing ores. Some of these
studies have indicated that the carbonaceous material
comprises active carbon which appears to adsorb the 65
gold cyanide complex Au(CNh- from cyanide leaching
solutions, as well as long chain organic compounds
which appear to form stable complexes with the gold.
4
DETAILED DESCRIPTION OF THE
PREFERRED EMBODIMENT
The present invention is directed to a process for
recovering gold from refractory ore. It is contemplated
that such an ore may contain one or more of the following
substances: carbonaceous compounds, sulfide minerals
of iron, arsenic, antimony and other metals. It is
anticipated that the gold in ore processed according to
the present invention will have an average particle size
on the order of 10 microns or less. Further, it is anticipated
that the carbonaceous compound content of such
an ore will not exceed 5% carbon by weight, and will
generally be between about 0.5 to about 1.0% carbon by
weight.
In order to provide specific examples of the process
of the present invention, the following description of
the preferred embodiment will focus, by way of example
only, on ore samples taken from the Mercur mine in
the Mercur Canyon in the Oquirrh Mountains, Tooele
County, about 56 miles by road from Salt Lake City,
Utah. The Oquirrh mountain range is predominantly
25 composed of Paleozoic sedimentary rocks which have
been folded, faulted and in some places intruded by
igneous rocks. The Mercur ore bodies are found in the
Mississippian age rocks consisting of thin-bedded carbonaceous
limestones. Intrusive rocks in the area consist
of irregular masses of porphyritic rhyolite which are
barren of gold mineralization. The gold mineralization,
which occurs as disseminations of micron sized particles,
forms discontinuous stratiform bodies. Deposition
of the gold took place in the carbonate host rocks from
hydrothermal solutions in a hot springs environment.
Minerals associated with the gold are pyrite, marcasite,
orpiment, realgar, barite and remobilized carbon. Gold
is found in four forms: as native gold, as gold included
in pyrite and marcasite, as gold adsorbed onto organic
carbon and as gold-hydrocarbon complexes. Decalcification
of the lime rocks, formation of the kaolinite and
introduction of silica to form jasperoid are the main
alteration features of the deposit. A significant portion
of the Mercur ore is oxidized and the oxidized zones
contain alteration products of the sulfide minerals while
being absent of organic matter. Accordingly, the composition
for a given sample ofgold-bearing ore from this
mine, as with any mine, can vary tremendously. Subject
to the foregoing limitations, various ore samples were
taken of refractory and oxidized ore both above and
below 6,850 feet elevation. The samples were combined
into composite samples and analyzed. A quantitative
55 analysis of composites 1-6 is given in Table 1 while an
X-ray analysis of composites 1-6 is given in Table 2.
Another series of composite samples, DC-l through
DC-3, were compiled and a quantitative analysis of
these samples is given in Table 3 while an X-ray analysis
60 of these samples is given in Table 4.
In general, for purpose of classification of the Mercur
ore, the term "refractory ore" shall be defined as an ore
which is less than 60% amenable to gold extraction by
65 standard cyanidation leaching techniques. Thus, said
refractory ore will have a percent extraction of gold
which is less than sixty by standard cyanide for leaching
technique.
4,552,589
SUMMARY OF THE INVENTION
3
least one hour, followed by the simultaneous contact of
the oxidized aqueous slurry, in a plurality ofstages, with
a cyanide complexing agent and granular activated
carbon, the carbon flowing in countercurrent fashion
5
with the slurry such that the gold is transferred to the
granular activated charcoal. However, this process still
consumes large amounts of cyanide, which may be the
single most expensive operating cost in an ongoing gold
extraction process, and the process is heavily time-con- 10
suming because of the retention times involved in the
oxidation and chlorination steps.
DESCRIPTION OF THE DRAWINGS
FIG. 1 is a graph of gold present in cyanidation tails
versus autoclave temperature for tests conducted according
to the present invention.
FIG. 2 is a graph of gold present in cyanidation tails
versus time in the pressure oxidation step for tests conducted
according to the present invention.
In the present invention, refractory gold bearing ores 15
are subjected to a treatment process in which at least
part of the substances which cause the ore to be refractory
are oxidized at an elevated temperature and oxygen
partial pressure to form an oxidized slurry which is 20
then subjected to a carbon-in-leach process. In another
aspect of the present invention, the amount of cyanide
consumed in the present invention is minimized while
water and a portion of the heat required in the present
invention are recovered and recycled.
When a refractory gold bearing ore is treated in accordance
with the present invention, an ore slurry is
formed which is introduced into at least one autoclave
maintained at an autoclave temperature in excess of 30
about 150· C. and an autoclave oxygen partial pressure
in excess of about 10 psia. The slurry is agitated sufficiently
to allow oxygen to oxidize at least part of the
substances which cause the ore to be refractory to
thereby form an oxidized slurry which is passed to a 35
carbon-in-leach process which utilizes granular activated
carbon in a countercurrent flow process. The
partial oxidation which occurs in the autoclave increases
the recovery of gold from particularly refrac- 40
tory ores in a relatively short time while reducing the
consumption of expensive cyanide. Thus, the present
invention provides a process by which refractory ores,
which may contain both carbonaceous and sulfur-bearing
compounds, may be economically treated for the 45
recovery of gold.
When the tailings water stream from the tailings pond
of the present process is recycled, the discharge of toxic
cyanide is eliminated and problems associated with
50 scarcity of water in some mining locations are at least
partially alleviated.
Accordingly, it is a primary object of the present
invention to provide an improved process for treating
refractory gold bearing ores.
This and further objects and advantages will be apparent
to those skilled in the art in connection with the
drawings and the detailed description of the preferred
embodiment set forth below.
5
4,552,589
6
TABLE 1
Quantitative Analysis of Composites I to 6
Composite
I 2 3 4 5 6
Refractory Oxidized Composite Refractory Oxidized Composite
Above Above Above Below Below Below
Element 6,850' 6,850' 6,850' 6,850' 6.850' 6.850'
Gold, oz/ton, FA 0.130 0.145 0.079 0.140 0.125 0.052
Gold, oz/ton, AA 0.125 1 0.128 0.068 0.128 0.1I5 0.048
Silver, oz/ton, FA 0.064 0.081 0.081 0.031 0.028 0.1I3
Silver, ozlton, AA 0.32 0.16 0.21 0.21 0.16 0.24
Copper, % 0.013 0.006 0.004 0.005 0.017 0.003
Lead, % 0.009 0.006 0.006 0.006 0.007 0.006
Zinc, % O.oI8 0.024 0.019 0.014 0.025 0.017
Iron, % 2.32 2.29 2.33 2.44 2.04 1.55
Total carbon, % 4.78 4.59 4.99 3.34 4.01 4.86
C03=, % 23.8 22.0 23.2 14.5 19.5 23.3
Organic carbon, % 0.42 0.19 0.35 0.44 0.1I 0.20
Total sulfur, % 2.60 1.65 1.75 3.09 0.90 0.95
S04=, % 3.07 2.66 2.00 4.15 1.23 1.15
Sulfide sulfur, % 1.16 0.61 0.88 1.02 0.31 0.37
Mercury, ppm 28 28 15 49 18 14
Thallium, ppm 249 147 131 175 125 73
Arsenic, % 0.25 0.20 0.13 0.05 0.08 0.04
Antimony, % 0.031 0.035 0.041 0.030 0.033 0.039
Magnesium, % 0.30 0.44 0.54 0.26 0.33 0.39
Calcium, % 14.9 16.8 15.8 10.1 12.4 15.5
Barium, % 0.38 0.26 0.1I 1.05 0.81 0.78
Chloride, ppm 470 190 280 260 180 50
Aluminum, % 3.41 3.39 3.77 4.16 3.36 2.64
Silica, % Si02 40.7 40.6 38.1 47.3 47.3 42.7
Nickel, % 0.006 0.006 0.006 0.006 0.005 0.005
Cadmium, % 0.001 0.001 0.001 0.001 0.001 0.001
Potassium, % 0.67 0.68 0.69 0.95 0.83 0.68
Sodium, % 0.50 0.035 0.1I 0.11 0.042 0.030
Phosphate, % P04 0.27 0.24 0.11 0.12 0.21 0.16
Selenium, ppm 1.0 0.8 1.1 1.5 0.7 0.7
Fluorine, % 0.17 0.16 0.14 0.13 0.12 0.12
Specific Gravity 2.67 2.67 2.73 2.71 2.68 2.71
FA = fire assay
AA - atomic absorption
INeutroD activation gold analysis of Composite 1 gave the following results: Gold = 4.203 ± 0.038 ppm; equivalent
to 0.123 oz gold/ton.
TABLE 2
X-Ray Analysis of Composites I to 6 40
Composite
Element, % 2 4 5 6 TABLE 3-continued
Copper O.oI8 0.010 0.010 O.oI8 0.013 0.020 Quantitative Analysis of
Silver 0.004 Composites DC-I to DC-3
Zinc 0.01I 0.023 0.022 0.007 0.023 0.015 45 Composite
Lead 0.012 0.026 0.019 0.016 0.026 0.03 Element DC-I DC-2 DC-3
Arsenic 0.34 0.28 0.19 0.045 0.071 0.036
Antimony 0.009 O.oI5 Silver, oz/ton, AA 0.1I 0.14 0.08
Iodine 0.007 0.016 Copper, % 0.004 0.004 0.002
Iron 1.4 1.4 1.3 1.5 1.3 I.2 Lead, % 0.007 0.007 0.007
Nickel 0.006 0.008 0.004 0.004 0.008 0.002 50
Zinc, % 0.008 0.020 0.006
Rubidium 0.01I 0.015 0.014 0.016 0.010 0.003 Iron, % 2.65 2.69 1.95
Barium 0.62 0.48 0.24 2.2 1.5 1.7 Total carbon, % 4.15 3.26 4.21
Strontium 0.077 0.084 0.076 0.078 0.074 0.090 C03=, % 18.8 25.9 14.6
Titanium 0.051 0.034 0.034 Organic carbon, % 0.47 0.42 0.12
Molybdenum 0.002 0.003 Total sulfur, % 2.28 1.90 0.39
Manganese 0.022 0.026 0.024 0.019 0.025 0.014 S04=, % 1.66 1.83 1.04
Yttrium 0.002
55 Sulfide sulfur, % 1.86 1.35 0.18
Zirconium 0.008 Mercury, ppm 29.6 25.2 33.0
Thallium, ppm 155 285 695
The above numbers represent a semi-qualitative and semi-quantitative analysis. No Arsenic, % 0.09 0.1I 0.13
check was made for elements with atomic number less than 22. Elements not Iised Antimony, % 0.07 <0.01 0.04
or for which no value is given were not detected. Magnesium. % 0.31 0.37 0.30
60 Calcium, % 12.5 16.7 12.8
TABLE 3 Barium, % 0.34 0.71 0.60
Chloride, ppm 270 300 170
Quantitative Analysis of Aluminum, % 4.58 3.20 3.70
Composites DC-I to DC-3 Silica, % Si02 44.7 38.5 47.4
Composite Nickel, % 0.008 0.009 0.007
Element DC-I DC-2 DC-3 65 Cadmium, % 0.001 <0.001 <0.001
Potassium, % 1.25 0.88 1.16
Gold, oz/ton, FA 0.088 0.1I0 0.1I6 Sodium, % 0.34 0.30 0.009
Gold, oz/ton, AA 0.083 0.113 0.108 Phosphate, %, P04 0.27 0.23 0.22
Silver, oz/ton, FA <0.01 <0.01 <0.01 Selenium, ppm 1.7 1.0 2.6
2FeS2+4CaC03 + 711202-Fe203 +4CaSo4X4C02
5
8
tory ore amenable to cyanidation. Gold is absorbed
from the leach pulp using the carbon-in-leach technique.
The gold is then stripped under pressure, and
recovered by electrolysis.
In the first step of the process of the present invention,
the ore is converted into a slurry. To form a slurry
from the ore initially mined, the ore is crushed and
blended. Thereafter, one or more stages of grinding
reduce the size of the ore and the ground ore is thick-
10 ened with the aid of a flocculant. The clear supernatant
is then recycled to the grinding circuit. Tailings liquor
is treated for removal of cyanides and recycled for use
within the grinding circuit, as will be discussed more
fUlly hereinafter, with additional fresh water make-up
15 being used as needed.
In the second step of the process of the present invention,
the ore slurry is rendered more amenable to cyanidation
by treatment through a mechanically agitated
autoclave or a plurality of mechanically agitated auto-
20 claves. For the purpose of example only, it will be assumed
hereafter that a plurality of autoclaves in series is
being used. The autoclave vessels are held under oxygen
pressure at an elevated temperature. Final heating
25 to operating temperature is accomplished by injection
of live steam and oxygen which may be introduced into
each autoclave. Gases should flow countercurrent to
the pulp, with a bleed being drawn from the first stage
to remove inert reaction products.
In the pressure oxidation step of the present invention,
oxygen consumption is dependent upon the refractory
nature of the ore. In the present example, it is assumed
that a range from about 0.30 to about 1.85%
sulfide sulfur, with an average level of 1.07%, is present
in the gold-bearing ore. Carbon dioxide is generated in
the process following the oxidation of the sulfides as
follows:
The liberation of carbon dioxide will begin during pulp
heat-up, between about 1500 C. and 1750 C., and may be
expected to continue during pulp cooling. However,
when the ore also contains carbonates, the sulfate ions
45 produced during the autoclave oxidation of the sulfide
react with the carbonates to produce calcium sulfate,
commonly known as gypsum, which may deposit on the
inside of the autoclave to form a relatively insoluble
50 scale to cause operational problems. One way to control
such formation is to add a basic compound which will
provide ions, such as sodium, which will combine with
the sulfate ions and which will not form an insoluble
species. When sodium carbonate was added to the auto-
55 clave, the gold analysis of the final cyanidation tails
went down for most composites in most tests. Less
extensive testing for sodium bicarbonate and sodium
hydroxide tended to indicate a similar beneficial effect.
The autoclave temperature is a critical parameter in
60 the pressure oxidation step where high temperatures
can lead to prohibitively expensive operational costs
while low temperatures will be inadequate to effectuate
improved gold recovery. By way of example only, a
summary of tests evaluating the effect of autoclave
65 temperature is given in Table 6 and shown graphically
in FIG. 1. The data show a decrease in the final cyanidation
tails gold assay with increased autoclave temperature.
A minimum in the tails assay was achieved at
4,552,589
88
95
90
88
89
92
0.12
DC-3
DC-3
0.025
0.014
0.11
0.010
0.16
2.0
0.019
0.014
1.2
0.093
0.087
0.055
0.029
0.009
0.11
DC-2
DC-2
0.021
0.016
0.042
0.009
0.24
2.1
0.010
0.016
1.3
0.10
0.089
0.064
0.058
0.011
Percent Extraction of Gold
7
0.12
DC-I
Standard Direct Pressure
Cyani- Carbon-in- Oxidation/
dation Leach (CIL) CIL
TABLE 4
TABLE 5
DC-l
0.004
0.008
0.017
0.008
0.13
2.2
0.013
0.56
0.062
0.10
0.037
0.041
0.008
Quantitative Analysis of
Composites DC-Ito DC-3
Composite
TABLE 3-continued
X-ray Analysis of DC Composites
Composite
Head Grade
(oz Au/t)
Element
Fluorine, %
Element, %
Copper
Zinc
Thallium
Lead
Arsenic
Iron
Nickel
Rubidium
Barium
Strontium
Titanium
Zirconium
Manganese
Yttrium
Selenium
Samples
I 0.125 53 65
2 0.128 78 82
3 0.068 27 69
4 0.128 22 78
5 0.115 88 88
6 0.048 57 81
DC
composites
1 0.083 38 70
2 0.113 46 65
3 0.108 91 87
FA = fire assay.
AA = atomic absorption
The above numbers represent a semi~qualitative and semi-quantitative analysis. No
check was made for elements with atomic numbers less than 22. Elements not listed
or for which no values are given were not detected.
In general, the process of the present invention begins
with the initial removal of the ore from the ground.
Thereafter, the mined ore is crushed and blended by
stockpiling in layers, the objective being to reduce fluctuations
in the constituents which cause the ore to be
refractory. These fluctuations may otherwise affect
process efficiency. The crushed ore is ground in a semiautogenous
(SAG)/ball mill circuit with cyclone classification.
The thickened product is treated under oxygen
pressure at elevated temperature to render the refrac-
Composites 1-6 and DC-I through DC-3 were sub- 30
jected to standard cyanidation and direct carbon-inleach
treatment, and most of the samples were also
subjected to the process of the present invention to
compare the improvement in extraction of gold from
the samples of varying compositions. The wide range in 35
the refractory nature of the samples is demonstrated by
the variable response to leaching by standard cyanida-
, tion techniques. A significant improvement in extraction
from the more refractory material was achieved by
the process of the present invention as is shown by the 40
results set forth in Table 5.
0.013
0.D15
0.014
0.016
0.D15
0.017
0.015
30
40
40
60
30
90
15
10
TABLE 7
50
50
50
50
50
50
25
Effect of Autoclave Time
02
Pressure Time Cyanide Tails Assay
psig min oz gold/ton
Composite 1:
10
4,552,589
TABLE 6
Effect of Autoclave Temperature
9
approximately 220·-250· C. This minimum was essentially
independent of the initial oxygen pressure. For
composite 1, the data show that the difference in the
tails for 50 or 100 psig oxygen pressure is not significant.
Accordingly, the dependence of the pressure oxidation 5
stage on the partial pressure of oxygen appears to be
minimal, provided a partial pressure of oxygen of about
10-25 psia is exceeded and adequate mixing is obtained.
30 All teSls at 225' C. and 40% solids.
0.022
0.015
0.020
0.D15
0.D18
0.017
0.023 .
0.018
0.012
0.013
0.010
0.112
0.026
0.019
0.013
0.012
0.010
0.009
5
10
15
30
60
90
5
10
15
30
60
o5
10
15
30
60
90
25
~
~
~
~
I~
In the third stage of the process of the present invention,
pulp from the pressure oxidation step is passed
through a surge tank to which a basic chemical may be
35 added for pH adjustment. It has been found that the
pulp from the pressure oxidation step should preferably
be cooled to a temperature below about 50· C. before it
is introduced to carbon-in-Ieach treatment. This cooling
may take place in a heat exchanger which recovers part
of the heat to be used upstream from the heat exchanger.
In addition to cooling the pulp, it may be
necessary to dilute the pulp before the carbon-in-Ieach
treatment wherein cyanide may be added to the first of
a plurality of mechanically agitated vessels in series in
which gold extraction from ore by cyanidation and
carbon absorption will proceed simultaneously. Pulp is
transferred continuously downstream through interstage
screens from a first vessel to the following vessel
in the series of vessels while activated granular charcoal
carbon is advanced from the last vessel toward the first
vessel. Fresh reactivated granular charcoal carbon is
added to the last stage and the loaded charcoal is withdrawn
from the first stage. Pulp leaving the last stage is
passed through an additional screen to scavenge some
attrited carbon before being discarded to a tailings pond
from which the tailings water may be recycled to the
slurry formation step.
When liquid from the tailings pond is recycled upstream
of the autoclaves, it has been found that free
cyanide in the recycled liquid unexpectedly decreases
the efficiency of the gold recovery of the present invention.
This is in contrast to conventional cyanidation
techniques in which the excess free cyanide would in
fact be utilized to help oxidize the slurry to be treated.
To overcome this unexpected problem, the free cyanide
must be removed by any suitable means.
The invention will be further illustrated in the following
example in which the pressure oxidation step is
20 Composite DC-I:
Composite DC-2:
25
Temp 0c.
180 50 0.028
200 ~ 0.020
215 ~ 0.017
225 ~ 0.016
249 ~ 0.021
250 ~ 0.016
150 100 0.070
180 ~ 0.031
180 ~ 0.027
215 ~ 0.021
225 ~ 0.011
Composite 2:
150 25 0.020
180 ~ 0.017
225 ~ 0.007
Composite DC-2:
150 0.060
180 0.023
200 0.014
215 0.014
225 0.010
All tests at 60 minutes and 40% solids.
In order to achieve adequate mixing without unnecessarily
diluting the slurry, it has been found preferable
to maintain the solids content of the slurry between
about 40% to about 50% and it is postulated that the
rate of reaction in the pressure oxidation stage is con- 40
trolled by the mass transport of oxygen to the solids'
surface. However, the selection of the proper type of
mixing equipment and the mixing speed utilized within
the autoclaves is deemed to be well within the scope of
one of ordinary skill in the art. 45
Another critical variable in the pressure oxidation
step is the residence time which the ore slurry spends in
the autoclaves. By way of example only, a summary of
tests evaluating the effect of autoclave time is given in
Table 7 and shown graphically in FIG. 2. The data 50
show a curve with minimum slope from approximately
30-90 minutes. Progressively higher tails assays result
from shorter times. Complete oxidation of the sulfides is
unnecessary; within an average 30-minute treatment
period found to be adequate, less than 50% of the sul- 55
fides may be oxidized. To maintain the oxygen pressure,
a bleed is required from the autoclaves as carbon dioxide
is evolved from the reaction of carbonates noted
above. Additionally, since the amount of oxygen consumed
will inevitably vary depending upon the refrac- 60
tory nature of the ore being treated, the oxygen partial
pressure must be monitored to assure it does not drop
below a minimum of about 10-25 psia. The partial pressure
of oxygen may be maintained by the introduction
of either pure oxygen, air or a mixture of both into the 65
autoclaves to ensure that an effective amount of the
substances which cause the ore to be refractory will be
oxidized in the pressure oxidation step.
4,552,589
60
11
described in greater detail. Thus, the operating conditions
in the autoclaves are as follows: a temperature of
about 2250 C., a retention time of about 30 minutes, an
oxygen overpressure· of about 25 psia, an agitation of
about 100 hp/5,OOO gallon unit, a pulp density of about 5
45% solids, a grind of about 80% passing through 325
mesh, a reagent addition of about 8 pounds NaOH/ton
ore, an oxygen supply (average) of about 45 lb/ton ore
and a steam supply (maximum) of about 250 1b/ton of
ore. At least six stages are utilized to avoid by-passing 10
inefficiency. This might be achieved through six vertical
units or one or two multistage horizontal autoclaves.
The gas bleed and oxygen input are controlled by manifolding
the bleed gas from each stage and exhausting to
control pressure, while injecting oxygen to maintain a 15
desired oxygen concentration in the atmosphere of each
vessel. The agitation is designed to reentrain gas at the
pulp surface. A flash heat exchange system is used to
recover heat, followed by steam injection to attain the
desired pulp temperature in the first stage. Bleed gas is 20
contacted with feed pulp to recover some heat as well
as to provide initial scrubbing. Although the pulp
would be alkaline, pitting corrosion could occur above
the pulp interface. Accordingly, a mild steel autoclave 25
would require at least a partial protective lining.
Having fully described the present invention, it will
be apparent from the above description and drawings
that various modifications in the process of the present
invention may be made within the scope of this inven- 30
tion. Therefore, this invention is not intended to be
limited except as may be required by the lawful scope of
the following claims.
What is claimed is:
1. A process for recovering gold from a refractory 35
ore slurry, said process comprising the steps of:
introducing the slurry to at least one agitated autoclave
maintained at an autoclave temperature in
excess of about 1500 C. and an autoclave oxygen
partial pressure in excess of about 10 psia; 40
agitating the slurry to allow oxygen to oxidize at least
part of the substances which cause the ore to be
refractory to form an oxidized slurry;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cya- 45
nide and activated charcoal, the charcoal flowing
in countercurrent fashion with said oxidized slurry
whereby precious metals are transferred to the
activated charcoal; and
separating the activated charcoal from the oxidized 50
slurry.
2. The process as recited in claim 1 wherein the autoclave
temperature is maintained between about 1800 C.
to about 2250 C.
3. The process is recited in claims 1 or 2 wherein the 55
autoclave oxygen partial pressure is maintained at a
minimum between about 10 psia and about 25 psia.
4. A process as recited in claim 1 wherein a basic
compound, is added to the ore slurry to form an insoluble
species with free sulfate ions.
5. A process as recited in claim 4 wherein the basic
compound comprises sodium carbonate.
6. A process as recited in claim 1, comprising the
further step of:
cooling the oxidized slurry before the oxidized slurry 65
is contacted with the cyanide.
7. A process as recited in claim 1, comprising the
further step of:
12
raising the pH of the oxidized slurry before the oxidized
slurry is contacted with the cyanide.
8. A process as recited in claim 1, comprising the
further step of:
diluting the oxidized slurry before the oxidized slurry
is contacted with the cyanide.
9. A process as recited in claim 1 wherein the ore
slurry is maintained within the autoclave step from
approximately 30-90 minutes.
10. A process for recovering gold, said process comprising
the steps of:
forming an ore slurry from a refractory gold bearing
ore;
introducing the slurry to to at least one autoclave
which is maintained at an autoclave temperature in
excess of about 1500 C. and an autoclave oxygen
partial pressure maintained in excess of about 10
psia;
agitating the slurry sufficiently to oxidize at least part
of the substances which cause the ore to be refractory
to form an oxidized slurry;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cyanide
and activated charcoal at a cyanidation temperature
less than about 500 C., said charcoal flowing
in countercurrent fashion with said oxidized
slurry whereby gold is transferred to the activated
charcoal; and
separating the activated charcoal from the oxidized
slurry.
11. The process as recited in claim 10 wherein the
autoclave temperature is maintained between about
1800 C. to about 2250 C.
12. The process as recited in claims 10 or 11 wherein
the autoclave oxygen partial pressure is maintained at
about 15 psia.
13. A process as recited in claim 10 wherein a basic
compound is added to the ore slurry to form an insoluble
species with free sulfate ions.
14. A process as recited in claim 13 wherein the basic
compound is selected from the group consisting of sodium
carbonate, sodium bicarbonate and sodium hydroxide.
15. A process as recited in claim 10 wherein the gold
in the ore exists in a finely disseminated form with an
average size of less than 10 microns.
16. A process as recited in claim 10 wherein the refractory
ore contains a sulfur bearing compound which
is at least partially oxidized in at least one autoclave.
17. A process as recited in claim 10 wherein the ore is
less than 60% amenable to standard cyanidation techniques.
18. A process as recited in claim 17 wherein more
than 85% of the gold existing in the ore slurry is transferred
to the activated charcoal.
19. A process as recited in claim 10 wherein the ore
slurry is maintained within at least autoclave for from
about ten minutes to about thirty minutes.
20. A process as recited in claim 10 wherein the activated
charcoal is separated from the oxidized slurry
from about 4 hours to about 8 hours after said activated
charcoal is contacted with the oxidized slurry.
21. A process for recovering gold, said process comprising
the steps of:
forming an ore slurry by grinding a refractory gold
bearing ore to form a slurry which is mixed with a
diluent stream;
4,552,589
13
introducing the slurry to at least one autoclave which
is maintained at an autoclave temperature in excess
of about 150· C. and an autoclave oxygen partial
pressure maintained in excess of about 10 psia;
agitating the slurry to oxidize substances which cause 5
the ore to be refractory to form an oxidized slurry;
cooling the oxidized slurry to a temperature less than
about 150· C.;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cya- 10
nide and activated charcoal, the activated charcoal
flowing in countercurrent fashion with the oxidized
slurry whereby gold is transferred to the
activated charcoal; 15
separating the activated charcoal from the oxidized
slurry;
collecting a tailings stream from the oxidized slurry;
treating a water portion of the tailings stream to destroy
any free cyanide and form a treated reclaimed 20
water stream; and
recycling the treated reclaimed water stream to a step
ofthe process upstream from the step of passing the
oxidized slurry to a plurality of stages.
14
22. A process as recited in claim 21 wherein the
treated reclaimed water stream is combined with a feed
stream to form the diluent stream.
23. A process as recited in claim 21 wherein the oxidized
slurry is cooled by a heat exchanger which is
utilized to preheat the slurry.
24. A process as recited in claim 21, comprising the
further step of:
diluting the oxidized slurry to dethicken said oxidized
slurry before the oxidized slurry is contacted with
the cyanide.
25. A process as recited in claim 24 wherein the reclaimed
water stream is used to dilute the oxidized
slurry.
26. A process as recited in claim 21 wherein the ore
slurry is maintained within at least one autoclave from
about 10 minutes to less than about 60 minutes and the
carbon-in-leach residence time is from about 4 hours to
about 8 hours.
27. A process as recited in claim 26 wherein the ore
slurry is less than 60% amenable to standard cyanidation
techniques and more than 85% of the gold in said
ore slurry is transferred to the activated charcoal.
* * * * *
25
30
35
40
45
50
55
60
65
ginati} ��oe@s� ut-grid-align:none;text-autospace:none'>portion of said high pressure leachate of step (c) and at
least a portion of said second fraction of step (a), are
contacted in an atmospheric leach at atmospheric pressure
and at a temperature below about 100· C. to form 20
an atmospheric leach residue and an atmospheric leachate
prior to contact in the low pressure leach of step (d).
4. A method according to claim 3 further comprising
separating from said first fraction a coarser third frac- 25
tion, said third fraction containing more magnesium
than said first fraction and less than said second fraction,
and contacting at least a portion of said third fraction
and at least a portion of said atmospheric leachate in the
low pressure leach of step (d). 30
5. A method according to claim 1 further comprising
separating from said first fraction a coarser third fraction,
said third fraction containing more magnesium
than said first fraction and less than said second fraction,
and contacting at least a portion of said third fraction in 35
the low pressure leach of step (d).
6. A method according to claim 1 or claim 3, wherein
at least a portion ofsaid low pressure leachate is neutralized
by the addition of a neutralization agent selected
from the group consisting of alkali and alkaline earth 40
oxides and alkali and alkaline earth hydroxides to form
a neutralized low pressure leachate.
7. A method according to claim 1, or claim 2, or claim
3, or claim 4, wherein at least a portion of said low
pressure leach residue is recycled to said high pressure 45
leach.
8. A method according to claim 3 or claim 4, wherein
at least a portion of said low pressure leach residue is
recycled to said atmospheric leach.
9. A method according to claim 8 wherein a coarser 50
fraction of said atmospheric leach residue is separated
from said atmospheric leachate and the remainder of
said atmospheric leach residue, and further comprising
filtering said coarser fraction into a filtrate and a filter
cake from which chromite is recovered.
10. A method according to claim 9 wherein at least a
portion of said filtrate is recycled to said atmospheric
leach.
60
65