4,552,589 Process for the recovery of gold from refractory ores by pressure oxidation

Patent Number/Link:

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

0.02

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

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.

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

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.

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

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

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.

4,552,589

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

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

0.12

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

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

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

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

TABLE 7

Effect of Autoclave Time

Pressure Time Cyanide Tails Assay

psig min oz gold/ton

Composite 1:

4,552,589

TABLE 6

Effect of Autoclave Temperature

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

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:

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

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:

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

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.

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.

* * * * *

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.