CA2016640A1 - Process for recovery of gold from refractory ores - Google Patents

Abstract

4LH ABR 5080.1 PATENT

PROCESS FOR RECOVERY OF GOLD FROM REFRACTORY ORES

Abstract of the Disclosure The present invention is directed to an improvement in a process for the recovery of gold from refractory sulfidic auriferous ores. The process comprises treatment of a slurry of the ore with sulfuric acid, oxidizing the treated slurry with oxygen gas under pressure in the presence of sulfuric acid, neutralizing the oxidized slurry, cyanidizing the neutralized slurry to leach gold therefrom, and recovering gold from the resultant leachate. In accordance with the improvement, the oxidized slurry is contacted with wash water in an oxidized pulp washing stage, wash water as introduced into the pulp washing stage being at a temperature lower than that of the oxidized slurry. A partial liquid/solids separation is effected within the washing stage, and a relatively high solids fraction comprising washed oxidized slurry and a liquid fraction comprising spent wash liquor are separately removed from the washing stage. The spent wash liquor is neutralized by mixing it with a base which forms a substantially insoluble sulfate salt upon reaction with sulfuric acid, after which the neutralized wash liquor is cooled. Precipitation solids are separated from the neutralized spent wash liquor and the cooled neutralized wash liquor is recycled to the pulp washing stage to provide wash water for contacting the oxidized slurry.

Description

~LH ABR 5~80.1 PATENT
Z01664(~

PROCESS FOR RECOVERY OF GOLD FROM REFRACTO~Y ORES
8ackground of the Inven~i~n This invention relates to the recovery of gold from ores and, more particularly to an improved pressure oxidation process for the recovery of gold from refractory ores.
In order to remove sulfide sulfur, refractory ores are conventionally treated by pressure oxidation before cyanide leaching. If the sulfide sulfur is not substantially oxidized, leaching is inhibited and leached gold remains locked in the sulfides. By treating the ore in an aqueous slurry at elevated temperature and oxygen pressure, the sulfur is oxidized and removed from the ore before it is contacted with cyanide leaching agent. Thereafter the gold is leached by the cyanide and acceptable yields are produced.
Pressure oxidation is an exothermic process but requires the use of a substantial amount of energy in pre-heating the ore slurry to a temperature at which the reaction is self-sustaining. The oxidized slurry may contain substantial amounts of iron, arsenic and other heavy metals which it is desirable to remove before cyanidation. These various metals are typically oxidized during the pressure oxidation step, but further measures are required if the salts and oxides of these undesired metals are to be removed from the process.
Weir et al U.S. patent 4,571,263 describes a process for pressure o~idation of refractory ores in which the effluent from the pressure osidation autoclave is subjected to a two step repulping operation with solids-liquid separations after each step. Liquid from the second separation step is recycled to the first repulping step. Liquid from the first separation step is in part recycled to pressure oxidation and in part subjected to a two step precipitation first with limestone and then with lime. Effluent slurry from the second precipitation 4LH ABR 5080.1 PATENT

~0~6640 step is subjected to solids-liquid separation and the liquid fraction is passed through a cooling pond and recycled to the second repulping step and the pressure oxidation step.
Weir 4,571,264 describes a pressure oxidation gold recovery process in which the effluent from the pressure oxidation step is repulped, thickened and then subjected to a two stage washing process. Water for washing derives from a liquid fraction produced in thickening the ore slurry after aci~ pretreatment prior to pressure oxidation. This liquid fraction is subjected to a two stage precipitation with limestone and lime, respectively. The effluent slurry from the second precipitation is-thickened, and the liquid overflow from the latter thickener is used as water in the second washing stage. A solids-liquid separation after the second washing stage produces a liquid fraction which is recycled and ser~es as the wash water in the first washing stage.
In both Weir et al '263 and Weir et al ~264, the neutralized oxidized slurry is subjected to cyanidation, followed by an eight stage carbon-in-leach adsorption process.

~oth patents disclose pressure oxidation at 160 to 200C and 700-5000 kPa (total pressure).
Weir 4,606,763 describes pressure oxidation at 165C
and 50-2000 kPa total pressure, using a compartment autoclave in which the fir~t compartment is approximately twice the size of each of the other compartments. Weir U.S. patent 4,605,439 di~clo~es a pressure oxidation process operated at 120 to 25~C and 350-6000 kPa. Mason et al U.S. patent 4,552,589 discloses alkaline pressure oxidation at 220-250C and 10-25 psia o~ygen partial pressure for 30 to 90 minutes. Matson et al U.S. patent 4,289,532 describes alkaline pressure oxidation at 140-190F using air.

4LH ABR 5080.1 PATENT
Z0166~!~

Summary of the Invention Among the several objects of the present invention may be noted the provision of an improved process for the recovery of gold from refractory ores; the provision of such a process which effectively removes sulfur, iron, arsenic, and other heavy metals; the provision of such a process which is effective for the removal of oxides, salts and any other o~idation products of iron, arsenic and other heavy metals;
the provision of such a process which can be implemented at relatively modest capital investment; the provision of such a process which is energy efficient; and the provision of such a process by which gold is recovered in high yield from relatively lean refractory ores.
~riefly, therefore, the present invention is directed to an improvement in a process for the recovery of gold from refractory sulfidic auriferous ores. The process comprises treatment of a slurry of the ore with sulfuric acid, oxidizing the treated slurry with oxygen gas under pressure in the pzesence of sulfuric acid, neutralizing the oxidized 81urry, cyanidizing the neutralized slurry to leach gold therefrom, and recovering gold from the resultant leachate.
In accordance with the improvement, the oxidized slurry is contacted with wash water in an oxidized pulp washing stage, wash water as introduced into the pulp washing stage being at a temperature lower than that of the oxidized slurry. A
partial liquid/solids separation is effected within the washing stage, and a relatively high solids fraction comprising washed osidized slurry and a liquid fraction comprising spent wash liquor are separately removed from the washing stage. The ~pent wash liquor is neutralized by mixing it with a base which forms a substantially insoluble sulfate salt upon reaction with sulfuric acid, after which the neutralized wash liquor is cooled. Precipitation solids are 4LH ABR 5080.1 P~TENT

X0~664S~
separated from the neutralized spent wash liquor and the cooled neutralized wash liquor is recycled to the pulp washing stage to provide wash water for contacting the oxidized slurry.
S The invention is further directed to an improvement in the aforesaid process in which a sulfuric acid treated auriferous ore slurry having a solids content of at least about 30% by weight is subjected to pressure oxidation in a horizontal autoclave at a temperature of between about 180 and about 225C, a total pressure of between about 275 and about 490 psia, and an oxygen partial pressure of at least about 25 psia for a period of at least 60 minutes. The sulfuric acid concentration of the slurry is between 5 and 40 gpl after the pressure oxidation is complete. The resultant oxidized slurry is washed with relatively cool water to reduce the iron and arsenic content of the slurry and cool the washed slurry to a temperature not greater than about 45F.
Other objects and features will be in part apparent and in part pointed out hereinafter.

~Lief Description of the Drawinqs Each of Figs. 1-4 is a flowsheet of a particular embodiment of the process of the invention;

Fig. 5 is a more detailed flowsheet illustrating the acidulation and pressure oxidation steps in a preferred processing scheme of the invention;

Fig. 6 is a more detailed flowsheet illustrating the o~idized slurry washing, cooling and neutralization steps in a preferred embodiment of the invention; and 4LH ABR 50~0.1 PATENT

Z0~6640 Fig. 7 is a more detailed flowsheet for the cyanidation, carbon-in-leach, and gold recovery steps in the process of the invention.

Corresponding reference characters indicate corresponding process and equipment features in the several views of the drawings.

~escription of the Preferred Embodiments The present invention provides an improved process for recovery of gold from refractory auriferous ores, including relatively lean ores containing as low as 0.10 oz Au per ton. The process is effective for recovery of gold from ores such as those found at the American Barrick Goldstrike property in Nevada, which are sulfidic, and contain iron, arsenic and other heavy metals. In accordance with the process, the various contaminants are oxidized under acidic conditions in a pressure oxidation operation, the sulfuric acid, oxides and salts produced in the pressure oxidation are removed in a combined washing and neutralization operation, and the washed neutralized slurry is subjected to carbon-in-leach cyanidation, preferably in a continuous countercurrent manner, for recovery of gold.
In order that the pressure oxidation step of the process of the invention operate autogenously, the ore used as feed to the process preferably contains at least about 3~ by weight sulfur in the form of sulfides. Exothermic oxidation of the sulfide sulfur generates the heat which brings the slurry to the temperature at which not only the sulfur but also the iron and other heavy metals are oxidized. However, by addition of steam to the pressure oxidation step, the process of the invention is also effective for the treatment of refractory sulfide ores containing as low as 1.5~ by weight s 4LH A~R 5080.1 PATENT

2016~
sulfide sulfur. As a further alternative, pyrite concentrates may be blended with the ore feed to provide an additional source of sulfide sulfur and assure that autogenous heat is sufficient to bring the autoclave to the desired reaction temperature and pressure. The latter alternative may provide a further advantage in allowing recovery of gold from the pyrite concentrate, where it is sometimes present in concentrations otherwise too low for economical recovery.
Illustrated in Fig. 1 is a preferred process of the invention. According to the process of this flowsheet, the ore is crushed and wet milled, and the ground ore slurry screened for trash or tramp material. Next the ground ore is thickened by removal of escess water in a solid-liquid separation operation. The carbonate content of the ore slurry is then substantially reduced by an acidulation treatment using fresh sulfuric acid. Removal of carbonate as CO2 in the acidulation step substantially reduces the volume of CO2 that must be vented from the autoclave, where it is generated from residual carbonate in the pressure oxidation step. Acidulation may be carried out in either a batch or continuous manner.
Next the acid treated ore slurry is subjected to pressure oxidation in the presence of sulfuric acid using oxygen gas at elevated pressure. The pressure oxidation step is typically conducted in a horizontal multi-compartmented autoclave, the compartments of which are preferably of substantially equal volume. Energy from the exothermic pressure oxidation is recovered by heat eschange between the oxidized slurry and acidulated feed to the autoclave. As indicated in Fig. 1, this heat e~change is preferably effected by letting down the pressure of the oxidized slurry, and using the steam which is flashed from the oxidized slurry to heat the acidulated autoclave feed, preferably by direct contact in splash condensers positioned ahead of the autoclave.

4LH ABR 5080.1 PATENT

201~
After it is partially cooled by flashing of steam, the oxidized slurry is directed to a washing operation where it is diluted with relatively cool water, and then subjected to a solids/liquid separation in which sulfuric acid and soluble metal salts produced in the pressure oxidation are separated in a spent wash solution liquid fraction. The spent wash solution is neutralized with lime or other base, thereby precipitating sulfate and heavy metals, clarified by solids/liquid separation, cooled, and recycled to the washing step where it serves as the source of fresh wash water. The solids stream from the clarifier is disposed of in a tailing facility.
Following the washing step and separation from the spent wash water, the cooled oxidized slurry is directed to a neutralization operation. Here lime or other base is added to increase the pH to allow for subsequent cyanide leaching.
Autoclave vent gas is added to oxidize the heavy metals remaining in the ore slurry. Gold is recovered from the washed and neutralized o~idized slurry by carbon-in-leach cyanidation, preferably in a continuous countercurrent system.
Referring to Fig. 5, ground ore slurry, a substantial fraction of which, for esample 65-85% by weight, passes 200 mesh, is directed to a trash screen 1 where rock, wood fiber and plastic larger than 20 mesh are separated and removed.
The ore slurry passing through the screen is directed to a mechanical thickening device 3, typically a vertical tank of large diameter which provides a net vertical flow low enough to permit sedimentation of the solid particles, In the thickener, the concentration of the ore slurry is increased from a range of about 10-25% by weight to a range of about 40-55%, preferably 45-50%, by weight. To promote separation of solids, a flocculant is preferably added to the thickener, for example, the polymeric thickener sold under the trade designation Percol 351, at a dosage of about 0.05 to about 0.2 4LH ABR 5080.1 PATENT
20166~0 pounds per ton of ore and a concentration in the thickener feed of between about 0.05% and about 2% by weight. Overflow from the thickener is recycled to the grinding circuit.
Thickened ore slurry underflow from the thickener is directed by a transfer pump 5 to a series of stirred acidulation tanks 7, 9 and 11, through which the slurry passes continuously. A
fresh sulfuric acid stream 13 is added to the acidulation tanks in order to release carbon dioxide from the carbonate contained in the slurry, and thereby reduce the equivalent carbon dio~ide levels in the ore to between about 0.05 and about 0.7% by weight, preferably not more than about 4 lbs/ton of ore. To promote removal of CO2, compressed air may be sparged into the acidulation tanks.
Residue slurry leaving the acidulation tanks, having lS an adjusted solids content of at least about 30%, preferably 40-SS~, optimum of 45-50~ by weight, i~ fed by a transfer pump lS to the first of a series of brick lined splash condensers 17, 19 and 21, in which the treated feed slurry for the pressure oxidation step is preheated by contact with steam flashed from the o~idized ~lurry leaving the pressure o~idation. The successive ~plash condensers are each, preferably, internally baffled to promote contact between steam and liquid, and are respectively operated at progressively higher pressure and temperature. Pumps are interposed to increa8e the pressure of the slurry between condensers, pump 23 transferring the slurry from condenser 17 to condenser 19, and pump 25 transferring the slurry from conden~er 19 to condenser 21. Preferably, condenser 17 is operated under a slight vacuum, condenser 19 is operated at substantially atmospheric pressure, and condenser 21 is operated under steam pressure.
Pressure o~idation is carried out in an autoclave 29, preferably multilined, the last lining being brick, to which the slurry i8 transferred, preferably by a diaphragm B

4LH A~R 5080.1 PATENT
X0166~

pump 27, from the last splash condenser 21. Addition of live steam to the slurry leaving the last splash condenser may be indicated for bringing the slurry to a temperature of at least about 175-180C, at which the exothermic pressure oxidation reactions become self-sustaining. In the autoclave, the slurry is passed through a plurality of compartments, where it is contacted in the presence of sulfuric acid with oxygen gas at a temperature of between about 180 and about 225C, an oxygen partial pressure of at least about 25 psia and a total pressure of between abaut 275 and about 490 psia. The final acidity of the slurry leaving the last compartment of the autoclave is between 5 and 40 grams sulfuric acid per liter of solution, and the final emf of the slurry is between about 480 and about 530 mv.
Noncondensables and steam generated during the pressure oxidation operation are vented preferably through a cyclone 31 which separates entrained solids for return to the autoclave. In the course of the autoclave operation, iron sulfides are oxidized to ferrous sulfate and sulfuric acid, further oxidation producing ferric sulfate; FeAsS is oxidized to arsenous acid and ferrous sulfate; and arsenous acid is oxidized to arsenic acid:

FeS + 72 + 2H20 - 2FeS04 ~ 2H2S04 4FeS04 + 2H2S04 + 2 ~ 2Fe2(S4)3 + 2H2 4FeAsS + 1102 + 2H20 ~ 4HAsO2 + 4FeS04 2HAs02 + 2 + 2H20 - 2H3As04 Oxidized slurry leaving the autoclave is passed through a choke 33 to reduce its pressure, then through a series of flash tanks 35, 37, and 39 where steam is flashed off to cool 4LH ABR 5080.1 PATENT

201664~
the slurry. Pressure of the slurry is progressively reduced by passage through chokes 41 and 43 between each flash tank.
Steam from each flash tank is recycled and contacted with autoclave feed slurry in a complementary splash condenser, operated at substantially the same pressure as the flash tank, for preheating the feed slurry. Thus, in the series as illustrated in the drawing, the first flash tank 35 is coupled to the last splash condenser 21, the second flash tank 37 is coupled with the second condenser 19, and the last flash tank 39 is coupled with the first splash condenser 17.
Steam leaving each of flash tanks 35, 37 and 39 is preferably passed through a cyclone 45, 47 or 49, respectively, for recovery of entrained solids. The recovered solids are blended back into the oxidized slurry.

Preferably, the temperature of the pressure oxidation is controlled at a level no higher than about 225C. Significantly higher temperatures can result in a runaway reaction and resultant overpressure of the autoclave.
Temperature can be controlled by a variety of means, including venting tailgas from the autoclave, venting steam from the first splash tank 35, and/or injecting cold water directly into the autoclave compartments.
Oxidized slurry having a solids content of at least about 30~ by weight and containing soluble sulfates, iron salts, arsenates, etc., leaves the last of the flash tanks at temperature in the neighborhood of 85 to 100C. Neither the soluble contaminants in this slurry nor the temperature thereof is conducive to efficient cyanidation for leaching of gold. In order to condition the slurry for gold recovery operations, the improved process of the invention subjects the hot contaminated oxidized slurry/to the novel washing, cooling and neutralization operation illustrated in Fig. 6. In accordance with this flow scheme, the slurry containing at least about 30% by weight solids is introduced continuously 4LH ABR 5080.1 PATENT

into an o~idized pulp washing stage comprising a washer/thickener tank 51 where it is diluted with relatively cool water in such relative proportions as to reduce the slurry temperature to between about 40-70C and dilute the solids content to 10-25%. Preferably, the slurry is cooled to a temperature no higher than about 45C and is diluted to a solids concentration in the neighborhood of 15% by weight.
Partial solids/liquid separation occurring in thickener 51 produces a continuous spent wash liquor overflow and a continuous washed oxidized slurry underflow having a strength of between about 35% and about 50%, preferably about 40% to about 45%, by weight solids. To assist in the separation, a flocculant is advantageously added to Lhickener 51.
Preferably, a nonionic polymeric flocculant, such as that sold under the trade designation Percol 351, is added at a concentration of about 0.05 to 0.2 lbs/ton of ore and about 0.05% to about 0.2% by weight solids.
The spent wash liquor is not recycled to the pressure oxidation or other early steps of the process, as in the prior art, but instead i8 reconditioned, cooled and recycled to thé step of washing the oxidized slurry. Thus, as shown in Fig. 6, the spent wash liquor is passed continuously through a series of stainless steel and carbon steel neutralization tanks 53, 55 and 57 where it is neutralized with lime to raise its pH to the neighborhood of 9.5 to 11, preferably about 10.0, and precipitate sulfates and compounds of iron, nickel, zinc and arsenic. Lime is highly preferred but the neutralization may be carried out with other bases which form substantially insoluble sulfate salts on reaction with sulfuric acid and are capable of raising the pH to a level at which iron and arsenate salts are precipitated.
Effluent from the spent wash neutralization is passed to a clarifier 59 where precipitated solids are removed in an underflow stream, typically containing 5-30% solids, that is 4LH ABR 5080.1 PATENT

Z0~664~
directed via a pump 61 to a tailings pond. Preferably, a flocculant is used in the neutralized spent wash clarifier also. In this instance, it is preferred to use a cationic polymeric flocculant such as that sold under the trade designation Percol E24 at a dosage of 1-4 ppm and a feed concentration of 0.02 to 0. 2 weight % solids.
Overflow from the neutralized spent wash clarifier is transferred by a pump 63 through open spray nozzles 65 to a cooling pond 67. Spraying of the neutralized and clarified spent wash liquor cools it by evaporation to a temperature near ambient, preferably not greater than about 32-40C, typically 20-38C. After it is cooled, the neutralized and clarified spent wash liquor flows to a wash maker making tank 69, where it is mixed with fresh water; and the combined stream is recycled via a pump 71 to washer/thickener 51 where it serves as the source of cool water for washing and cooling the oxidized slurry.
Underflow from washer/thickener 51 is transferred by a pump 73 through a series of rubber lined neutralization tanks 75, 77 and 79 where the pH of the slurry is raised to about 10.5 to 11.5, preferably about 1].0, by treatment with lime to condition it for cyanidation. Recovery of gold from the o~idized slurry by carbon-in-leach ~C-I-L) cyanidation is illustrated in Fig. 7. The washed and neutralized slurry is pumped, 81, to the first of a series of agitated carbon-in-leach tanks 87, 89, 91, 93, 95, and 97, and possibly more depending on retention time, countercurrently to a flow of granular activated carbon. Loaded carbon recovered from the carbon-in-leach operation is stripped with hot alkaline cyanide solution in a stripping vessel 99, and gold is recovered from the stripping solution by conventional means such as electrowinning and refining (not shown).
The process of the invention provides for high recovery of gold, for example, in a yield exceeding 80%, from 4LH ABR 5080.1 PATENT

~0~66~
relatively lean refractory auriferous ores containing 0.10 to 0.50 oz gold per ton. It is effective for removing contaminating elements such as iron, arsenic, nickel, and zinc from the oxidized slurry, and can be implemented with relatively modest capital investment. The autoclave conditions and means for recovery of exothermic reaction heat provide not only efficient gold recovery but efficient use of energy.
Illustrated in Figs. 2-4 are various alternative embodiments of the process of the invention. Fig. 2 varies from Fig. 1 in that an indirect heat exchanger 101 rather than coupled splash condensers and flash tanks are used for exchange of heat between the oxidized slurry and the feed to the pressure osidation autoclave. After leaving heat exchanger lOl, the oxidized slurry is further cooled in a slurry flash tank 103. Figs. 3 and 4 both use an indirect heat exchanger 105 rather than a cooling pond for cooling of the recycled neutralized and clarified spent wash liquor. The cooling fluid for the indirect heat exchanger comprises water circulated in a closed circuit through a cooling tower 107.
In the process of Fig. 3, heat exchange between oxidized slurry and autoclave feed is by means of coupled flash tanks and splash condensers as in Fig. 1, while in Fig. 4, heat exchange between oxidized slurry and autoclave feed is accompli~hed via an indirect heat exchanger 101 as in Fig. 2.
In those embodiments in which transfer of heat from the oxidized slurry to the treated slurry autoclave feed is accomplished by indirect heat exchange rather than by coupled flash tanks and splash condensers, the indirect heat exchanger is preferably a double pipe exchanger in which the inner pipe is constructed of titanium and the outer pipe of steel. The oxidized slurry is passed through the interior of the titanium pipe and the relatively cold pressure oxidation feed slurry is passed through the ar.nular space between the pipes.

4LH ABR 5080. 1 PATENT

Further illustrations of the process of this invention are given below.
Three samples of each Betze and Deep Post ore were pulverized to minus 200 mesh and identified as Head Samples A, B, and C. Each of the three head samples was fire assayed for gold and silver, and the A samples for each ore type were assayed for other elements of interest.
The analyses are listed in Table 1 and show average gold analyses of 0.205 and 0.336 oz Au/ton for the Betze and Deep Post composites, respectively. The average of 75 assay determinations for the Betze ore during the course of testing was 0.201 oz Au/ton. Likewise, the average of 11 determinations for the Deep Post ore was 0.328 oz Au/ton.

4dfb `_ Patent ~01664 a~ ~. r-~

~c l oo C4 r-l C n~ a j o o ~ ~ o r~ o ~ r~ co o r~l o o o o o~ ~r~ t~ o O o o o o ~ ~r o ~ o~ u~ o a l ~ l ~
c l lu ~r ~ ¦ o O

N o ~P
m a o o ~ ~ ~0 0 ~ O
o o o o O~ ~ O 1` r~
o o o o o ~ o ~I

O ~ O C4 C4 ~
~o 0~ 0~ d o~ o~
N " O~o ~ .r ` C C C ~ d 3 o ,~ o O ~ I n~ n~ n~
, ~ u~ o ~ ~ ~ ~,q ~.q t~ ~ o ~ ~;

4LH AB~ 5080.1 PATENT

E~AMPLE 1 20~66~

A pilot plant arranged in a manner generally corresponding to the flow sheet of Fig. 1, initially was operated continuously for about 2 1/2 days (62 hours) using the Betze ore. The total retention time in the circuit was estimated to be approximately 36 hours, so that a large portion of the time during this first period of operation was required to fill the autoclave thickener and the various neutralization and CIL tanks.
The autoclave conditions during this period are listed in Table 2. Sulfuric acid in a proportion of 100 lbs.
per ton of ore feed was added to the autoclave feed. This was the stoichiometric amount required to neutralize the natural carbonate in the ore.

~`
4LH AaR 5080. 1 PATENT

TAI~LE 2 Z01664~

Start-Up Conditions for ~ilot Plant ~utoclave Retention time 90 minutes Feed 40% solids by weight ~85% passing 200 mesh Temperature 435F (225C~

Pressure 420 psig total O~ygen 2 standard liters f 2 per stage (total of 8 SLPM) Acid addition 100 lb H2SO4/ton ore Mixers 500 rpm The circuit finished filling during the morning of the secona 24 hour period and was shut down a day later. The CIL and neutralization circuits were both profiled and the gold fire assays of the solids determine~. The data indicated rapid gold leaching; i.e., the CIL tailings after 2 hours of leaching in the first stage assayed 0.024 oz Au/ton compared to 0.019 oz Au/ton for the last stage after 16 hours of leaching. The calculated gold dissolution was 90% based on final tailings of 0.019 oz Au/ton and the average neutralization product ~CIL feed) which assayed 0.190 oz Au/ton.

4LH A~R 5080.1 PATENT

The solution in the autoclave discharge ~ rlng this period contained approximately 22 to 25 grams of free acid per liter and the off-gas from the autoclave measured approximately 96% oxygen and 3% carbon dioxide. Typical solution analyses show arsenic concentrations of approximately 70 to 100 mg As/liter in the autoclave discharge liquor and 0.2 to 0.3 mg As/liter in both the CIL tailings and clarifier overflow liquors. The mercury in all the liquors was below the detection limit of 0.005 mg per liter. The iron in the autoclave discharge liquor was as high as 996 mg per liter but assayed 1.1 to 1.7 mg Fe/liter in the CIL tailings and clarifier overflow liquors.
Cyanide addition of 1 lb NaCN/ton of ore was initially added to the first leach stage. The cyanide concentration in the final CIL tailings solution measured 0.06 g NaCN/liter (<0.12 lb NaCN/ton solution). The cyanide addition was increased to as high as 2 lb NaCN/ton of ore in sùbsequent operations, but good leach results were obtained at levels averaging 1.5 lb NaCN/ton of ore feed.
Analysis of the laboratory CIL tests indicated that the sulfide sulfur in the autoclave solids averaged 0.05~ S~, the gold assay of the CIL tailings averaged 0.013 oz Au/ton, and the calculated gold dissolution averaged 93.4%.

In this sxample the pilot plant of Example 1 was operated continuously for 60 hours. The acid addition to the autoclave feed was reduced from 100 to 80 pounds of sulfuric acid per ton of sre feed, but otherwise the autoclave conditions were the same as for Example 1. Plant recycle water was used to dilute the autoclave discharge slurry prior to thickening.

9LH A~R 5080.1 PATENT

A summary of analyses for this example is given in Table 3. The average ore feed assayed 0.207 oz Au/ton and, with the exception of one 12-hour composite, the CIL tailings averaged 0.014 oz Au/ton. This corresponds to a calculated gold dissolution of 93.2% (assuming no weight changes).

4dfb _ A~R 5080 PATENT

~016640 ,~ ~
C~ o ~o ~ ~ o ~ o v o, CO o ~o " o o ~ o ~ e c O O ~ O ,~ O O ~ O ~
P- t~ V ~ V V ~ V O

' ~ I~ U r~ ~ ~ Ul, u~ ~¢ & E ~1 0 r I cO ~1 u7 o _I O ~1 0 _~
P~ . o~
I ~o ~
.- ~ .- ~r ~ ~ u, ~ u~ ~ I
t~l t O ~ 1` ~ o ~r ~D O ~ a~
~ _ ~ ~ v ~ vl ~ v o~o~

r l ~ U ~ o ' o E ~ O O O CO. _~ 'JJ
C~
tl ~ rO` o o ~o o, ~r o ,~ ~" , o ,~
~ ~ ~1 0~1 0 r l O 0 ~1 ~O O O o ~o N vv v r-l ~
O ~0 l ~ O ~ 1 o .
, .,~ ~ '~ ~
o ~ O t~ O ~ ~
~a o ,,~ ~a o u, ,~ :~ O
3 ~ U~
O ,~ ~ V~ ,0~ 30 3 0 ~ tn tn rc v1 ~1 la ,a ~ ,1 _I u~ . .
~ ~ ~ U ~ Q
O u~ :~
~ b P ~ ~ ~ ~' ,~
14 14 ~ C~ t~ O O ~ Is~ E~ ~ n1 ~ ~ ~ 3 ul n~ C '~u nl ~ C I O X ~
P~ V .~ _) n1 4LH ABR 5080.l PATENT

Z0~6640 There was one 12-hour period when the CIL tailings composite assayed 0.027 oz Au/ton. The exclusion of this value during averaging produced the numbers in parentheses.
It is of particular interest to note the presence of gold in the CIL feed liquor. There was no cyanide in the process at this point and it is not known why there should be soluble gold. However, the presence of soluble gold was substantiated by solution assays for other periods of pilot plant operation.
Also, the low gold assays for many of the CIL feed solids are indicative of soluble gold. These solids assayed only 0.152 oz Au/ton compared to the autoclave discharge solids which assayed 0.204 oz Au/ton. This is a significant difference and could occur only if (1) there was a substantial weight gain ~appro~imately 30%) during neutralization, or (2) if there was soluble gold in the CIL feed liquors. It is considered e~tremely improbable that there was such a large weight gain and, in fact, later laboratory tests demonstrated a minimal weight gain of les~ than 2%. Calculations indicate a gold balance is achieved when the solubilized gold is considered in the balance.
The mercury analyses in Table 3 show less than 0.005 mg/l in ~olution throughout the c1rcuit except for the CIL
tailings ~olution which as~ayed 0.212 mg Hg/liter. The arSeniC values show a concentration of less than 100 ppm in the autoclave discharge solution, 0.08 mg As/liter in the clarifier overflow, and appro~imately 0.3 mg As/liter in both the CIL feed and tailings liquors.
A Jummary of the laboratory CIL tests showed that the average CIL tailings assayed 0.015 oz Au/ton with a corresponding average gold dissolution of 93.0%. These values agree extremely well with the results obtained in the plant and summarized in Table 3. The average sulfide sulfur analyses in the autoclave discharge products used for bath CIL leaching was 0.05%S'.

4LH A~R 5080.1 PATENT

During the last two days (48 hours) of the 60 hour operation of Example 2 the CIL circuit was profiled. The solids and li~uor gold analyses are listed in Table 4. Both sets of data show very rapid rates of gold dissolution. The assays from the first day show that the gold in the solids from the first stage are not significantly different from those in the last stage. On the second day, the solids assays in the first and last stages differed by only 0.004 oz Au/ton.
Both sets of data show high tailings at mid-points in the CIL
circuit. On day one, for example, the solids in Stages 3 and 4 assayed 0.027 and 0.024 oz Au/ton, respectively. It is not known if these are anomalous data or if they are indicative of aberrations in the operation of the autoclave of CIL circuit.

TABLe 4 Profiles of the Pilot Plant CIL Circuit Day 1 Day 2 CIL SolidsSolution SolidsSolution ~ta~e oz Au/ton ma Au/l oz Au/ton ma Au/l 1 0.017 1.33 0.020 0.65 2 0.019 0.24 0.021 0.07 3 0.027 0.04 0.017 0.013 4 0.024 0.01 0.016 <0.004 0.017 0.01 0.027 <0.004 6 0.015 0.01 0.027 >0.004 7 0.017 0.01 0.016 <0.004 8 0.016 <0.01 0.014 <0.004 4LH ABR 5080.1 PATENT

20166~
E~AMPLE 3 For this example the pilot plant of Example l was operated so as to study autoclave retention times of 90, 80, 70 and 60 minutes using the Betze ore and a sulfuric acid addition of ~0 pounds per ton of ore. Other autoclave conditions were the same as in Example l, i.e., a teMperature of 435F and a total pressure of 420 psig ~50 psig overpressure).
The retention times were varied by changing the feed rate to the autoclave and adjusting the acid and 02ygen additions accordingly. After each change the autoclave was allowed to equilibrate for a time period equal to three times the tested retention time. For example, 240 minutes (4 hours) were allowed for equilibration after switching to 80 minutes retention time.
The data for each test include sulfide sulfur analyses for the solids in each autoclave compartment, the corresponding redox potentials (emf), free acids, and pH
values for the liquors, and gold dissolutions from the solids. The gold dissolutions for the compartment solids were determined by shaking 10 grams of sollds with an excess of cyanide and carbon in a test tube for 16 hours. The gold dissolutions of the final autoclave products were determined by dupllcate standard CIL tests of splits from a 2-liter sample of the autoclave discharge.
The data in Table 5 show that each of the tested retention times achieved e~cellent sulfide sulfur oxidations.
In every case the autoclave discharges contained 0.07~ or less sulfide sulfur corresponding to oxidations in excess of 95%.
The other data indicate that the final autoclave products leached very well and the tailings from the batch CIL tests were in a range of O.OlS oz Au/ton with calculated gold dissolutions of 92% or better.

- 4LH ABR 5080.1 PATENT

Z0~664~3 The data in Table 5 for the four tests using 80 pounds of acid addition and retention times of 60 to 90 minutes were used to determine the rate of sulfide sulfur oxidation. The percentages of sulfide sulfur oxidation were 5 calculated based on analyses for the solids in each of the autoclave compartments and a head value of 2.4%S~.
The data showed rapid rates of oxidation:
approximately 95% of the sulfide sulfur is oxidized within the first 30 minutes. Longer times increased the oxidation by only 2 to 3 percentage points.
The data in Table 5 include two periods of pilot plant operation using retention times of 75 and 90 minutes and no acid addition to the autoclave feed. The sulfide sulfur analyses of 0.06 to 0.07%S~ for the autoclave discharge solids were comparable to similar tests using acid. The batch CIL
tailings appear to be slightly higher than tests with acid, but the gold dissolutions of the final autoclave discharges remained in excess of 90%.

4dfb -- -- ABR 5080 PATENT

20~66~0 ~ 1~ o o o I o ~o ~ ~ ~ u~
o o ~ o ~

O O ~ , I O

c a . aa) g' .~ 'u~
_~ C _ ~ O~ ,~ u~ o o 1~ N 1~ ~ ~ 0 t` o ~J ~- ul a~ r~N ~ ~ ~ i ~
e v a r ~ a , o .~ ~ ~ ~ I~ , o ,~ oo oooo oo oo oo v~
i~ i~ ~- ¦~NO O O O O O O O O O O O ~N
'C O ~uOl o ~ ~ ~ u~ r~ a o C~ C , ~ D~ ~ ~ U WO
rc +
:~ ~ ~
r~ 1~ ~ ~ o o o o o o o o o o O O O c ~i c r~ o o o o o o o o o o o o ~

~ t Ct N ~
~ JJ a ~ a o c o o o o o o o o v~ u~
.~ ~3 ~$ ;~ o~ o~ a~ a~ ~ o c u o W V

~-~1 v o o o o o o o o o o o o ~ 3 ~ ~q oo ~ oo a~ o a~
a 3 u~ o ~ I ,~ ~~ o E~ u~
D

4LH ABR 5080.1 - PATENT

ZO:16fi4~

Autoclave oxidation of the Betze ore using reduced oxygen additions with 80 pounds of acid addition was investigated during the runs of Example 3. The amount of oxygen used was measured using a high-pressure mass flowmeter calibrated to indicate standard liters of oxygen flow per minute. Oxygen flow rate was controlled by precision needle values. The efficiency of the autoclave treatment was evaluated by sampling the four autoclave compartments and assaying the solids for sulfide sulfur and the liquors for free acid, pH, and emf. Samples of the final autoclave products were batch leached in the laboratory; however, none of the solids from the individual autoclave compartments were leached.
Results of these tests and others, summarized in Table 6, show good results even with the oxygen reduced to as low as 92 pounds added per ton of ore feed. The use of the lowest oxygen producéd sufficient oxidation to make the ore amenable to cyanidation. Tests 67 and 68 are replicate tests of the autoclave product using only 92 pounds of oxygen; CIL
tailings were 0.017 and 0.018 oz Au/ton with calculated gold dissolutions in excess of 90%.
There were, however, differences in the rates and completion of the sulfide sulfur oxidation. For example, using the 92 pounds of osygen achieved a sulfide sulfur analysis of 0.11% for the solids in the second autoclave compartment. By comparison, the use of 154 pounds of oxygen achieved solids assaying 0.07~ sulfide sulfur from the same compartment.

4dfb -- -- ABR 5080 PATENT

Z01664~

~ ~o D J ~ N

J,~3J. ~'q ~
a o VVo .~

~~ O N O ~ r 0 P~ ~ ~ ~ ~ ~ ~ I 1~ ~ ~ N O~
, t. 1 O~
~g ~8, ~
~ u.,a ~
~9~ % a ~ a ,, ~ .,7 ,, ~ ~ '',, _, '` ~
~ ~ ~ o o o O O O O O O O O
o.~ ~ ."~3 o o o o o o o o o o o ~ ~ ~ . ~ O
C r. ~'D r~
o c ~1 o oo o o oo o o o o ,~
.,~ ~q .. .. .. .... ..
.,1., ~ ~ oo oo oo oo oo oo u~c, . , J ~ V~
3 P le ~ u~ o; o .~ ~ 1 o~ a~
c ` ~
~J~ OO OO OO OO OO OO
~ ~ o 0 oD ~ c~

~ ~ h 2 7 :, 4LH ABR 5080.1 PATENT

PLE 5 201664~

The pilot plant of Example 1 was also operated continuously for approximately 52 hours using Deep Post ore feed. The autoclave conditions included the following: a feed of 40% solids by weight, a retention of 90 minutes, 435F
(225OC), and 420 psig total pressure (50 psig overpressure).
An acid addition of only 12 pounds per ton of ore feed was required to achieve 10 to 15 g H2SO4/I in the final autoclave discharge solution. During one six hour period, the temp-10 rature was reduced to 410F and the pressure to 325 psig total (50 psig overpressure).
The autoclave oxidation was evaluated by leaching samples of the autoclave slurry in the laboratory using the standard CIL procedure described earlier.
The entire pilot plant was operated during this period but only the autoclave portion of the circuit was in equilibrium. This is shown by the low gold assays for the solids in the thickener underflow and CIL feed samples for the last day of preparation. Both samples assayed only 0.2 oz Au/ton compared to the average aùtoclave discharge solids assay of 0.35 oz Au/ton and the Deep Post ore assay of 0.336 oz Au/ton.
Although there was approximately 1.4 mg Au/liter in the CIL feed solution, the inclusion of this gold would increase the eguivalent CIL feed solids assay to only 0.27 oz Au/ton. This is significantly lower than the Deep Post but higher than the Betze material and it would seem, therefore, that the slurries in the neutralization and CIL circuits were mixtures of the Betze and Deep Post materials.
Results of the laboratory batch CIL tests are summarized in Table 7. The CIL tailings of the material produced using the standard autoclave conditions (420 psig total pressure, 50 psig overpressure, and 435F) averaged 9LH ABR 5080.1 PATENT

20166~0 0.019 oz Au/ton with a corresponding calculated gold dissolution of 94.6%. The reduced temperature and pressure conditions resulted in a CIL response which was only slightly less; the tailings assayed 0.022 oz Au/ton and the calculated gold dissolution was 93.8%.

Table 7 Summary of Laboratory Batch CIL
Tests of DeeD Post Autoclave Product Laboratory Batch Autoclave Conditionsl CLL_Tests Temp, Pressure, Acid Product,Tailings %Au F psia Total lk/ton %Soz Au/ton Dissolution 435 420 lZ 0.~8 0.019 94.6 410 325 12 0.07 0.022 93.8 1/ Other conditions include 40~ solids, 90 minutes retention, and 154 pound 02/ton of ore feed.

The results indicate that the Deep Post ore was very amenable to cyanidation after autoclave o~idation using conditions which were at or near the standard conditions used for the Betze ore.

,~

Three batch laboratory autoclave tests were made using the ~etze ore auger composite. The products were cyanide leached at temperatures of 25, 37 and 49C. The 4LH ABR 5080.1 PATENT

20166a~3 autoclave products from two different pilot plant tests using the Betze ore feed were also leached in the laboratory at temperatures of 25, 30, 40, 50, and 65C.
Results for the laboratory batch autoclave tests in S Table 8 show a trend of increasing gold dissolutions corresponding to increased leaching temperatures. The leach tailings, for example, decreased from 0.020 to 0.016 oz Au/ton when the temperature was increased from 25 to 49C.

4dfb . _ ABR 5080 PATENT
20~6 o U~
o~ll o o o a~
C n~ ~ ,~ o~ r o ,~ n o r~ ,~
1 ::~ ~ r~ N ~ ~ ~ ~ ,~ o ,~
~ U~¢ ~ , n N O o o O O o o O O O O o O ~

~1~ r1 ~) ~

.C~ ~ ~ ~ ~ O O~
n t_) Z ~ ~ ~ ~ ~ ~ ~ n . ,-1 N ~.
~ql , .~
.~: c ~ 0 ~ D O CO ~O Y
,~
O O O O O O O O O O O O O
~ ~ ~ oOOOO oOOOO OOo a ~3 ~ .
c ~a a) E~ ~ ~ t~ u~ ~ N 0~ r al ~ ,~1 p ~,) ~ ~ N ~ ~ N N N O ,~1 N P
~ ~ ~ cn ~ 'U

~, n~
~ .d n ~ ~ O O O ~ o O O u u~
l~ l~ N ~ ~ N ~ 0 N
. J~
. ~0 P ~ ~I N . _I

~ ~ ~ ~ 3 J~ ~ ~
U E~ ~ D t~ 0 ~N N N N N ~ ~ a~ U O

9LH ABR 5080.1 `~ PATENT

Z0166~3 Results obtained by leaching the pilot products indicate no trend associated with the leaching temperature.
There were only minimal differences in the tailings assays which were within normal expected experimental and analytical precision. However, the data, and particularly those from Example 2, indicate increased cyanide consumptions with increasing temperature. This is not unexpected and is indicative of either increased reactivity with the cyanicides or higher rates of cyanide volatilization.

Autoclave discharge slurry from Example 2 (using the Betze ore) was tested to determine what effects, if any, different neutralization techniques might have on the subsequent cyanide gold dissolutions.
In test CIL-17, -18, and -19 the autoclave slurry was neutralized at temperatures of 25, 40, and 70C, respectively. The neutralization in each case was made by adding milk of lime directly to the autoclave slurry to achieve pH lO.S, holding the slurry at the temperature and pH
for 2 hours, and batch CIL testing for 16 hours.
For test CIL-21, the autoclave slurry was heated to 90C, diluted to 10~ solids ~by weight) using fresh water, thickened to 40~ solids, and then decanted. The thickened slurry, which wa~ now at ambient temperature, was neutralized to pH 10.5 using milk of lime. This procedure simulated the pilot plant operation.
The results are summarized in Table 9 and with the exception of Test 18 at 40C, the leach tailings all assayed 0.014 oz Au/ton. The higher tailings assay for Test 18 make interpretation of the results difficult. However, these results appear to be anomalous and, if so, the gold dissolutions were independent of the tested neutralization techniques.

4dfb `--- PATENT

Z0166~1 .~
C C U~
o-,l 3 ~
.0 ~ rl ~ ~ E
nl 5~ ~ ~ ~ ~ ~ a~ E
rl U ~ 1 Orl O rl n~ O
~ ¢ ~ S ~ vl ~, ~ N O O O O ~ ~ O
n O ,~~ r1 a) ~ e ',~
~ ~ ~ Z ~a ~ e ~ ~ ~ 3 ~1 ", z :~ ~ ~ a) ~ u~ u~ ~ JJ ~ ~ 3 o~ o~
C C E4 . . . . a~
~ Z ~ ~ O ~1 ~1 E~ .C n~ C ~ '~
O C: 3 p E~ ~ n ,~ o 3 N ~ ~ C ~ ~ ~ ~r O a .,~ ~ n .-1 ~'C O O O o ~ e;~ N
~ NO O O O I~
:~ ~ ~ 01 ~
Z '~ ~1 ~ O O E-~ :1 nl ,C O~O a~
O ~! ~ ~ ~ .a ,~r~ o ~D V ~ ~ O
~ . O O n~ ~u~ ~ o 1~ 0 3 .~0 ~ 1,) O~ ooo~ cn ~ ~ ~ o nl ~1 ~I P ~ P~
:~ ~ ~ 'V ~ C
i N o ~ m o O a~
C n~ ~ > '1 _ JJ ~u~ o O u~ U d ~ ~ ~ V
~ :1~ ~ a c~
:z; ~ s ~ ~ ~ a ~ ~ o ~
~ s~
U 11 E~
~n In ..
oO~o o c ~ a Q) I~ a~ a~ ,1 O
, 4LH ABR 5080.1 PATENT

Z0~66~0 The cyanide consumptions for Tests 19 and 21 were significantly higher than the other two tests. In the case of Test 19, this was probably due to starting the cyanide leaching immediately after neutralizing while the slurry was still at 70C. It is not known why the consumption for Test 21 was higher since the dilution and thickening process allowed the slurry to cool to ambient conditions prio~ to cyanidation. However, since this process simulated plant conditions, it correlates with experience in the plant.

E~N~PLE 8 A sample of autoclave slurry (Betze ore) was split into six equal portions which were each aged (held in storage at ambient temperature) for times of 0 to 96 hours followed by CIL testing. Prior to cyanidation, the aged slurries were filtered, the solids washed using three water displacements and repulped with fresh water. The objective was to determine whether aging the autoclave slurry had any deleterious effects on the gold dissolutions.
The results summarized in Table 10 show no significant differences in the gold dissolutions, but there were measurable differences in the cyanide consumptions. The latter two tests using 72 and 96 hours of aging demonstrated lower consumptions than the other tests.

4df b ~ PATENT

~0166~

V~q O O~ ~ ,, ~ ~r o ,~ ~ ,~
U ~ ~ ~ O O O O O O , ~ o ~J
o `
a CD O r~
~3 3 ~ 11 u~ o o o r- OD
~; d Z C ~ ~ ~ ~r~
~.)r~
Q~
PIn d r ~ ~ ~ O
O 1~ ~
J~ 'C ~ _~ ~ O O O O O O
~ I P N O o o o o o q J~_ .~~ ~ ~
~1~ O ~ ~ . . ~ . .
u o ~ ~ ~ rJ~r .C~ p O
O P
~,~
~C~
~." ~ ~ O U~ ~CO ~ ~D
~;v ~

~ r ~ ~r ~
u~ oooooo ~ o o o o o o ~I
J~
co ~ o r~t 4LH ABR 5080.1 PATENT

A sample of the pilot plant thickener un~erflow (~etze ore) taken during the same time period during which the autoclave slurry samples were taken was neutralized in the laboratory from pH 1.3 to 10.5 using milk of lime to determine weight changes during neutralization. A worst case weight change would occur if all the lime was converted to CaSO42H2O
(gypsum), and this would result in a weight gain of 1.6% with the lime addition of 11.5 pounds of CaO (90% basis) per short ton of dry autoclave solids.

During the run of Example 2 samples of the Betze ore autoclave feed and discharge were screened at sizes from 48 to 400 mesh and each fraction was assayed for sulfide sulfur and gold. In each case there was only a trace of plus 48-mesh material and it was not assayed. There was insufficient material to assay the minus 65- and plus 100-mesh fractions of gold.
The results, given in Table 11, show sulfide sulfur values in the feed which were more or less evenly distributed throughout the various size fractions. There was a slight concentration of gold in the minu8 400-mesh fraction and it contained 67.6% of the gold compared to 60.7~ of the total solids weight.
By comparison, results for the autoclave discharge show major differences in the weight and gold or sulfide sulfur distributions. Approximately 42% of the sulfide sulfur values were contained in the plus 270 size fractions compared to a of only 22% of the total solids, and the minus 400-mesh fraction contained 72~ of the total solids and only 50% of the sulfide sulfur.
The results also show a concentration of almost 89%
of the gold values in the minus 400-mesh fraction of the 4LH ABR 5080.1 PATENT

;~0166~
autoclave discharge slurry compared to 72% of the total solids. This high concentration of gold in the minus 400-mesh size fraction is probably a major reason for the rapid gold dissolutions which were demonstrated in the pilot plant CIL
circuit.
In view of the above, it will be seen that the several objects of the invention are achieved and other advantageous results attained.
As various changes could be made in the above process without departing from the scope of the invention, it is intended that all matter contained in the above description or shown in the accompanying drawings shall be interpreted as illustrative and not in a limiting sense.

Claims (31)

1. In a process for the recovery of gold from a refractory sulfidic auriferous ore, comprising treatment of a slurry of the ore with sulfuric acid, oxidizing the treated slurry with oxygen gas under pressure in the presence of sulfuric acid, neutralizing the oxidized slurry, cyanidizing the neutralized slurry to leach gold therefrom, and recovering gold from the resultant leachate, the improvement which comprises:

contacting the oxidized slurry with wash water in an oxidized pulp washing stage, the wash water as introduced into the pulp washing stage being at a temperature lower than that of the oxidized slurry;

effecting a partial liquid/solid separation within said washing stage, and separately removing from said washing stage a relatively high solids fraction comprising a washed oxidized slurry and a liquid fraction comprising a spent wash liquor;

neutralizing the spent wash liquor by mixing it with a base which forms a substantially insoluble sulfate salt on reaction with sulfuric acid;

separating precipitated solids from the neutralized wash liquor;
cooling the neutralized wash liquor; and recycling cooled neutralized wash liquor to said pulp washing stage to provide wash water for contacting said oxidized slurry.

4LH ABR 5080.1 PATENT

2. An improved process as set forth in claim 1 wherein the concentration of solids in said high solids fraction is at least about 35% by weight.

3. An improved process as set forth in claim 2 wherein the concentration of solids in the oxidized slurry entering said pulp washing stage is at least about 30% by weight, and the amount of wash water entering said stage is sufficient that the average concentration of solids in the combination of the oxidized slurry and wash water feed streams is not higher than about 15% by weight.

4. An improved process as set forth in claim 3 wherein said oxidized ore slurry and said wash water are fed continuously to, and said high solids fraction and said spent wash liquor are removed continuously from, said pulp washing stage.

5. An improved process as set forth in claim 4 wherein said pulp washing stage comprises a single gravity thickener.

6. An improved process as set forth in claim 3 wherein the temperature of said high solids fraction leaving said pulp washing stage is not higher than about 45°C.

7. An improved process as set forth in claim 6 wherein the temperature of said cooled neutralized wash liquor is not greater than about 35°C when it is introduced into said pulp washing stage.

4LH ABR 5080.1 PATENT

8. An improved process as set forth in claim 7 wherein said base is effective for the precipitation of sulfate, iron, and arsenate from said spent wash liquor.

9. An improved process as set forth in claim 8 wherein said base comprises lime.

10. An improved process as set forth in claim 1 wherein said neutralized wash liquor is used entirely for washing of said oxidized slurry, and is not used for dilution or neutralization of the slurry fed to the pressure oxidation step.

11. An improved process as set forth in claim 1 wherein there is no separation of liquid from said oxidized slurry prior to washing of the slurry in said pulp washing stage.

12. An improved process as set forth in claim 1 wherein said spent wash liquor is cooled by use of a cooling pond.

13. An improved process as set forth in claim 12 wherein the oxidized slurry leaving the pressure oxidation step is cooled prior to introduction into said pulp washing stage by exchange of heat with the treated slurry fed to said pressure oxidation step.

4LH ABR 5080.1 PATENT

14. An improved process as set forth in claim 13 wherein said oxidized slurry is cooled by passage sequentially through a series of flash tanks, the pressure of the slurry being reduced between each of these flash tanks and any immediately subsequent flash tank in the series, steam from each flash tank being contacted with treated feed slurry in a corresponding contact condenser of a series of contact condensers complementary to said flash tanks, the treated slurry pressure oxidation feed being passed through said contact condensers for purposes of preheating, each of said corresponding contact condensers being operated at substantially the same pressure as the flash tank from which it receives steam, the pressure of the feed slurry being increased by pumping means between each contact condenser and any immediately subsequent contact condenser in the series.

15. An improved process as set forth in claim 13 wherein said oxidized slurry is cooled by indirect heat exchange with the treated slurry fed to said pressure oxidation step, and is further cooled after said indirect heat exchange by reducing the pressure of the oxidized slurry and flashing steam therefrom.

16. An improved process as set forth in claim 15 w herein indirect heat exchange is effected by passing said oxidized slurry and said treated pressure oxidation feed slurry through opposite sides of a double pipe heat exchanger.

17. An improved process as set forth in claim 16 wherein the inner pipe of said heat exchanger is constructed of titanium and the outer pipe is constructed of steel, the 4LH ABR 5080.1 PATENT

oxidized slurry being passed through the interior of the titanium pipe and the relatively cold pressure oxidation feed slurry being passed through the annular space between the pipes.

18. An improved process as set forth in claim 1 wherein said spent wash liquor is cooled by indirect transfer of heat to cooling tower water.

19. An improved process as set forth in claim 18 wherein the oxidized slurry leaving the pressure oxidation step is cooled prior to introduction into said pulp washing stage by exchange of heat with the treated slurry fed to said pressure oxidation step.

20. An improved process as set forth in claim 19 wherein said oxidized slurry is cooled by passage sequentially through a series of flash tanks operating at progressively decreasing pressure, and the feed slurry to the pressure oxidation step is heated by passage through a series of contact condensers operating at progressively increasing pressure, each flash tank being coupled by a steam line to a corresponding one of said contact condensers that is operated at substantially the same pressure as the flash tank, the feed slurry in each contact condenser being contacted with steam received from the flash tank to which the condenser is coupled.

21. An improved process as set forth in claim 19 wherein said oxidized slurry is cooled by indirect heat exchange with the treated slurry fed to said pressure 4LH ABR 5080.1 PATENT

oxidation step, and is further cooled after said indirect heat exchange by reducing the pressure of the oxidized slurry and flashing steam therefrom.

22. An improved process as set forth in claim 21 wherein indirect heat exchange is effected by passing said oxidized slurry and said treated pressure oxidation feed slurry through opposite sides of a double pipe heat exchanger.

23. An improved process as set forth in claim 22 wherein the inner pipe of said heat exchanger is constructed of titanium and the outer pipe is constructed of steel, the oxidized slurry being passed through the interior of the titanium pipe and the relatively cold pressure oxidation feed slurry being passed through the annular space between the pipes.

24. In a process for the recovery of gold from a refractory sulfidic auriferous ore, comprising treatment of a slurry of the ore with sulfuric acid, oxidizing the treated slurry with oxygen gas under pressure in the presence of sulfuric acid, neutralizing the oxidized slurry, cyanidizing of the neutralized slurry to leach gold therefrom, and recovering gold from the resultant leachate, the improvement which comprises:

subjecting a sulfuric acid treated auriferous ore slurry having a solids content of at least about 30% by weight to pressure oxidation in a horizontal autoclave at a temperature of between about 180° and about 225°C, a total pressure of between about 275 and about 490 psia, and an oxygen partial 4LH ABR 5080.1 PATENT

pressure of at least 25 psia for a period of at least about 60 minutes, the sulfuric acid concentration of said slurry being at being between about 5 and about 40 gpl after said pressure oxidation is completed; and washing the resultant oxidized slurry with relatively cool water to reduce the iron and arsenic content of the slurry and cool the washed slurry to a temperature not greater than about 45°C.

25. An improved process as set forth in claim 24 wherein the oxidized slurry leaving the pressure oxidation step is cooled prior to introduction into said pulp washing stage by exchange of heat with the treated slurry fed to said pressure oxidation step.

26. An improved process as set forth in claim 25 wherein said oxidized slurry is cooled by passage sequentially through a series of flash tanks, the pressure of the slurry being reduced between each of these flash tanks and any immediately subsequent flash tank in the series, steam from each flash tank being contacted with treated feed slurry in a corresponding contact condenser of a series of contact condensers complementary to said flash tanks, the treated slurry pressure oxidation feed being passed through said contact condensers for purposes of preheating, each of said corresponding contact condensers being operated at substantially the same pressure as the flash tank from which it receives steam, the pressure of the feed slurry being increased by pumping means between each contact condenser and any immediately subsequent contact condenser in the series.

4LH ABR 5080.1 PATENT

27. An improved process as set forth wherein said oxidized slurry is cooled by indirect heat exchange with the treated slurry fed to said pressure oxidation step, and is further cooled after said indirect heat exchange by reducing the pressure of the oxidized slurry and flashing steam therefrom.

28. An improved process as set forth in claim 27 wherein indirect heat exchange is effected by passing said oxidized slurry and said treated pressure oxidation feed slurry through opposite sides of a double pipe heat exchanger.

29. An improved process as set forth in claim 28 wherein the inner pipe of said heat exchanger is constructed of titanium and the outer pipe is constructed of steel, the oxidized slurry being passed through the interior of the titanium pipe and the relatively cold pressure oxidation feed slurry being passed through the annular space between the pipes.

30. An improved process as set forth in claim 24 wherein said auriferous ore initially contains at least about 1.5% by weight sulfur in the form of sulfides.

31. An improved process as set forth in claim 24 wherein prior to pressure oxidation the ore slurry contains pyrite concentrate and the sum of the sulfide moiety contained in the ore and the sulfide content of the pyrite concentrate constitutes at least about 1.5% by weight of the solids in said slurry.