CA1174055A

CA1174055A - Cyanidation of gold ores

Info

Publication number: CA1174055A
Application number: CA000391260A
Authority: CA
Inventors: David Weston
Original assignee: Individual
Current assignee: Individual
Priority date: 1981-12-01
Filing date: 1981-12-01
Publication date: 1984-09-11

Abstract

ABSTRACT OF THE DISCLOSURE
This invention relates to a cyanide process for the recovery of gold from complex sulphide bearing gold ores wherein in a preferred embodiment of the invention, a pulp of the ore is subjected to aeration conditioning at at least one optimum pH
point resulting in gold recoveries exceeding 90%, reduction in cyanidation of gold dissolution time to a maximum of 24 hours, and making it possible to recycle substantially all of the barren solution.

Description

~L~74~5 1 BACKGROUND OF T~E INVEN~ION

This invention is primarily applicable to the recovery of gold from gold bearing ores using at least cyanide as the gold solvent.
The ores to which the invention are applicable, carry in addition to gold values, one or more sulphide minerals of which the most common are pyrite, marcasite, pyrrhotite, arseno-pyrite and stibnite.
In treating the more complex of such ores, current practice uses flotation to produce a bulk sulphide concentrate containing varied percentages of the contained gold values in the ore. The bulk sulphide concentrate is roasted and the roasted concentrate is treated in the cyanidation process to recover the contained gold values. Not only is this an expensive process but in addition, causes serious losses in cyanide solution as the roasted sulphides form deleterious salts that inhibit the gold dissolution and if not bled from the circuit, gold recovery is materially reduced. As these solutions contain free cyanide, they pose an environmental problem.
In addition during the roasting process sulphur dioxide fumes are produced which again pose an environmental problem.
Further, in ores containing arsenopyrite and stibnite, arsenous oxide and antimony oxide are formed and normally, at least the arsenous oxide recovered. ln the recovery and handling of arsenous oxide, unless extreme care is taken, there is a serious question of a health hazard to the workmen.
I have invented a cyanide process whereby I not only can eliminate at least roasting but also can recirculate the plants' cyanide solution, thus not only eliminating the costs ~.~74e~55 1 involved in roasting but also a major part of the environmental and health hazards associated with current practices. In addition, the gold recovery is at least equal to current practice and in some of my reagent balances indicated to be appreciably higher.
In applying my invention to this type of ore, I use at least one stage of aeration conditioning in preferably a violent-ly mechanically agitated circuit wherein the pH is closely con-trolled to one or more optimally determined pH points and in the presence of one ore more agents selected from the group consisting of dispersants, chloride salts, lead salts, and cyanide salts.
In the past, aeration of pulps has been used particular-ly for gold ores containing pyrrhotite. This process was normal-ly carried out with no cyanide present in the solution. The object was simply to oxidize the pyrrhotite and the pH of the solutions considered of little importance. In some cases the solutions thus produced were discarded to waste and the solids then treated in a cyanide circuit. In other cases where c~anidation directly followed, it was necessary, following the gold removal from the solutions, to discard at least part of the solutions to waste which not only entailed a loss in cyanide but also created an environmental problem in discarding free cyanide.
For clarity, particular terminology used herein and for purposes of describing the invention are defined as follows:
When I use the term "pounds per ton" lime, lead salts, dispersants, and chloride salts I mean "pounds per short ton of ore". "Cyanide strengths" means "pounds of NaCN per short ton of solution".
Where I refer to an "optimum pH point" I mean "the practi-cal pH point at which the pulp can be maintained". For instance, if I refer to an optimum pH point of 10.5, in plant practice it ~7~55 1 may vary plus or minus approximately 0.2, with erratics due to changes in plant operating conditions or poor operating plant control.
"Reducing Power" or RP is a measurement of the effect-iveness of the cyanide solution to dissolve the gold contained in the ore. The lower the RP number, the higher the eff~ciency of the cyanide solution. For instance an RP of l.0 is highly efficient, while an RP of 5.0 indicates a low efficiency solution that would normally be discarded to waste.

When I refer to "cyanide", I mean the commercial cyanides normally used in the gold recovery process su~h as calciurn cyanide, sodium cyanide, and potassium cyanide.
"Dispersants" - Although the major part of my research work was carried out with the family of sodium silicates, there are probably other families which will have similar results, and for this reason the invention is not to be considered limited to this family only.
"Chloride Salts" - The salts used in the programme con-sisted of calcium chloride, ammonium chloride, and potassium20 chloride.
"Lead Salts" - The lead salts used in the programme were lead nitrate and lead acetate. There are other applicable lead salts such as litharge.
"Alkaline Agent" - To raise or maintain the pH of a pulp of the ore, my preferred alkaline agent is lime in the dry form or in slurry form as calcium hydroxide.
The primary objects of this invention are firstly to eliminate the roasting process used in complex sulphide mineral containing gold bearing ores. Furtherr in the cyanide process for the dissolution and precipitation of the contained gold s~
1 values, to place the cyanide solution in such condition that after the removal of the gold values from the solution, substantially all of the barren solution can be returned to the circuit for continuous dissolution and recovery of the gold values.
By obtaining the above noted objectives, both workmen and the major environmental hazards are automatically removed in addition to appreciably lowering operating costs.
A further object of the invention is to improve gold recovery in less complex sulphide mineral bearing ores.
SUMMARY OF THE INVENTION
_ .
I have invented a process wherein in using cyanide as a gold solvent, I am able to treat sulphide containing gold bear-ing ores without the use of roasting while achieving recoveries of the gold values appreciably in excess of 90%.
In a preferred embodiment of my invention, which in-cludes treating complex sulphide minexal bearing ores, I use the following steps:
(a) Grind in solution with or without cyanide present.
(b) Where substantial free gold is present in the ore r a gravity concentration stage wherein the gravity concentrate is amalgamated and the amalgam treated to recover the gold contained therein.
Alternately I may omit this stage.
(c) Thickening stage. Normally the pulp following the grinding stage or gravity concentration stage is too dilute to feed directly into a cyanide gold dissolution circuit and re~uires thickening pre-ferably in the range of about 55 to 70% solids.
The overflow from the thickening stage may go ~74~55 I directly to the gold recovery circuit for recovery by precipitation or alternately partially or wholly close-circuited with the grinding and gravity concentration stages~
(d) The thickener stage pulp underflow will either go to regrind followed by at least one aeration conditioning circuit or directly to at least one aeration conditioning circuit.

In my aeration conditioning circuit I prefer to use mechanical agitation wherein the~pulp has a minimum residence time period of 60 minutes with a preferably continuous flow of air disseminated throughout the pulp.
In this aeration conditioning circuit I
maintain at least one optimum pH point in the pH
range of about 8.5 to 12.0 using at least lime as the alkaline agent and in the presence of agents selected from the group consisting of dispersants, chloride salts, lead salts, and cyanide salts.
(e) Following the aeration conditioning circuit the pulp passes to the conventional cyanide gold dissolution and gold recovery circuits.
(f) In using a second aeration conditioning circuit I
prefer to insert this circuit following 12 to 24 hours residence time in (e), and repeating the aeration conditioning circuit described in (d).
(g) Following the aeration conditioning circuit, the pulp passes to the conventional cyanide gold dissolution and gold recovery circuits.

:~174~:D~
1 In treating generally less complex types of sulphide mineral bearing ores where the major object is simply to increase recovery, I prefer to use the following steps:
(a) Grind in solution in the presence of a dispersant and at least lime as an alkaline agent and at an optimum pH point in the range of about 8.5 to 12.0 with or without cyanide present. My preferred dispersant is chosen from the family of sodium silicates.

lb) Where substantial free gold is present in the ore, a gravity concentration stage wherein the gravity-concentrate is amalgamated and the amalgam treated to recover the gold contained therein.
Alternately I may omit this stage.
(c) Thickening stage. Normally the pulp following the grinding stage or gravity concentration stage is too dilute to feed directly into a cyanide gold dissolution circuit and re~uires thickening preferably in the range of about 55 to 70% solids.
The overflow from the thickening stage may go directly to the gold recovery circuit for recovery by precipitation or alternately partially or wholly close-circuited with the grinding and gravity con-centration stages.
I have found that for the maximum rate of thickening, the density of the pulp of the feed to the thickening stage should be less than 35~
solids, part cularly in us-ing a flocculating agent (d) The thickener stage pulp underflow may be treated in the following alternate circuits:

'7~D55 1 (i) Directly or to regrind and directly to the cyanide gold dissolution circuit.
(ii) To regrind followed by at least one aeration conditioning circuit, followed by one or more gold dissolution circuits, and finally to the gold recovery circuit.
(iii) Directly to at least one aeration con-ditioning circuit followed by one or more gold dissolution circuits and finally the gold recovery circuit.
DESCRIPTION OF THE DRAWING~
Figure 1 This figure shows a preferred flowsheet of the invention.
To release gold that is normally associated with the sulphides, I prefer to use at least one stage of wet grinding wherein the grinding unit is in closed circuit with a classifier to produce a product that is at least 70% minus 200 mesh. In the grinding circuit shown at 10 I normally prefer to have cyanide present as shown at 11.
The discharged product of the at least one grinding stage may either pass directly to the gravity reCQVery circuit 14 or alternately to a classifier (not shown) and the classifier overflow (not shown) to the gravity recovery stage. Alternately as shown at 13 the classifier overflow (not shown~ may pass directly to the thickener stage shown at 22.
The tailings 21 from the gravity recovery circuit may either be returned to a classifer (not shown) in the grinding circuit or alternately to the thickening stage shown at 22.

The gravity concentrate shown at 15 would normally 74~?55 1 be treated in the amalgamation circuit shown at 16 to produce a gold amalgam 17 which is retorted (not shown) to recover the mercury and gold bullion. The tailings from the amalgamation circuit shown at 19 will normally consist of a high percentage of the coarsest ground sulphides and returned to the grinding circuit. The underflow 23 from the thickening stage 22 may either pass to a regrind stage 25 or alternately to an aeration condition-ing circuit 27 as shown at 24.
If a regrind stage as shown at 25~is used, it is preferably in open circuit and the product 26 passes directly to the aeration conditioning circuit shown at 27. Not only does this circuit bring about the high gold recoveries previously re-ferred to but most surprisingly has the following beneficial met-allurgical effects:
With optimized chemical reagent balance, the cyanidation time required following this circuit is a maximum of 24 hours.
The normal cyanidation residence time in a conventional circuit is 72 hours. This factor alone results in large capital cost savings. In addition it prevents the fouling of the cyanide solutions allowing the complete circulation of barren solution instead of normal practice wherein such solutions are discarded to waste with the ensuant high cost of cyanide and lime losses together with creating an environmental problem.
Further, with the use of comparatively high concen-trations of dispersant in this circuit, following 24 hours cyanidation, surprisingly, thickening characteristics of the pulp were excellent together with outstanding clarity of the thickener overflow solutions. The product 2~ from the aeration conditioning circuit is fed to the conventional cyanide dissolution and gold recovery circuits shown at 29.

~.~1 '74q:DSS
1 Figure 2 The gold bearing ore shown at 30 is prepared for the grinding circuit 31 by conventional means such as crushing to the appropriate feed size.
To the grinding circuit 31, I prefer to add at least a dispersant 35 preferably selected from the family of sodium silicates and at least either a solution ~ontaining calcium hydro-xide or lime in the dry or slaked form or any combination there~
of as shown at 34.
In addition, I prefer to have an agent as shown at 36 added to the grinding circuit and selected from the group con-sisting of cyanide salts, lead salts, and chloride salts. The cyanide salts may be added in the form of circulated barren or mill solution. (not shown) The product 32, normally from the last grinding unit in the grinding circuit 31 is preferably closed-circuited with the classifier shown at 33 with the underflow shown at 52 returned to preferably the last grinding unit Inot shown). Where the gravity concentration circuit shown at 37 is used, I prefer to pass the product 32 through the gravity concentration stage to produce a gravity concentrate 38 which will contain the bulk of the coarse free gold contained in the ore and subsequently recovered in the subsequent amalgamation circuit shown at 39.
The gravity concentration circuit tailings shown at 40 proceed to the classifier 33.
The classifier overflow shown at 42 is adjusted to the following conditions as shown at 43: the pulp density to a maximum of 35% solids and preferably less than 30~ solids; the pH to a minimum of 9.5 and preferably to within the pH range of 10.0 to 12.0 using CaO or Ca(OH)2 as the alkaline agent; the ~ ~ 74~55 1 addition of a flocculant when required and preferably selected from the group consisting of guar gums and polyacrylamides.
The adjusted and modified classifier overfloW 42 is fed to the thickener stage shown at 44. The solution overflow 45 may alternately be sent to the conventional gold recovery circuit shown at 46 or partially or wholly closed-circuited with the grinding and gravity concentration circuits shown at 47. The thickener stage 44 pulp underflow shown at 48 has the following alternate treatment stages. As shown at 49 it is reground in a preferably open circuit mill and passes to the aeration conditioning circuit shown at 50 and previously described in the specification. Following this circuit, it passes to conventional cyanidation and gold recovery circuits as shown at 51. The second alternative is to by-pass the regrind stage as shown at 53 and feed directly to the aeration conditioning circuit shown at 50. The third alternative is to feed the underflow pulp 48 and as shown at 54 directly or through a regrind mill (not shown?
to the conventional cyanidation and gold recovery circuits as shown at 51.
EXAMPLES OF THE INVENTION
The product used for the following examples was the thickener underflow from an operating plant in which the pre-ceeding steps involved grinding in cyanide solution followed by a jig, and amalgamation of the jig product. During the period of testing, the gold recovery by cyanidation in the grinding and amalgamation circuits was approximately 55%. The calculated head values for the tests were arrived at by dividing the actual thickener underflow solids assay by 45%. The ore itself was a refractory gold oré containing pyrite, pyrrhotite, arsenopyrite and stibnite, and for many years used flotation and roasting with discarding to waste, the bulk of the barren solution.

3 ~1.74~
1 The cyanide plant results prior to flotation and follow-ing approximately 60 hours cyanidation residence time was a tail-ings analysis averaging 0.08 ozs. gold per ton at a head value of approximately 0.50 ozs. gold per ton. The bulk of the barren sol-ution was bled off to waste carryinq approximately l.Q lbs. cyan-ide per ton and approximately 2.0 lbs. lime equivalent per ton.

The RP factor used for discarding the barren solution was 4.5.
In using a grindin~ step on the thickener underflow samples, the followin~ were the approximate screen analyses at the various times used.

1.5 Mins....................... 75% minus 200 Mesh 3.0 Mins....................... 85% minus 200 Mesh 4.5 Mins....................... 95% minus 200 Mesh The thickener underflow averaged about 65% minus 200 mesh.
As the grinding was carried out in batches, it would tend to give inferior metallurgical results than in plant prac-tice where a more desired differential grind would be obtained with a ball mill in closed-circuit with a classifier~ different-ially grinding the sulphides and thus obtaining better gold liberation. Further, the fine iron produced by laboratory batch grinding is far higher in proportion than in plant practice.
As this fine iron is soluble in cyanide, fouling of the solutions is far more seriously affected in this type of laboratory testing.
For the above noted reasons the results shown in the examples will probably be appreciably improved upon in plant practice.
In all of the following examples, the thickener under-flow samples were batch re-ground in a laboratory rod mill fol-lowed by mechanical agitation and aeration conditioning in a :~7~55 1 Denver laboratory flotation cell with the air control valve normally at the fully open position.
The optimum pH points of the pulp during the aeration eonditioning period were maintained by intermittent additions of slaked lime. A pH probe constantly in the pulp showed the pH
over the complete aeration conditioning cycle.
Example 1 Regrind time was 2.0 mins. with the addition of 0.2 gms. CaO to mill. Aeration eyele was 90 mins. at pH of 11.0 followed by 30 mins. at pH of 10.5.

2.0 lbs. of Na2SiO3 per ton ore was added to the beginning of the aeration conditioning cycle and 2.0 lbs. of NaCN per ton ore for the last 30 mins. of the aeration cycle.
Following the aeration cycle, the pulp was transferred to a eyanidation jar and placed-on a rolls for 24 hours.
At the end of the 24 hours cyanidation, the RP of the solution was 2.1 which is outstanding considering the intense aeration eonditioning period of 2 hours.

The caleulated head analysis of the ore was 0.25 ozs.
Au/Ton and the tailings analyzed 0.019 ozs. Au/Ton for a Au reeovery of 92.8~.
Following the 24 hour cyanidation period, the pulp was filtered, reground for 2 mins., the filtrate returned to the pulp, and the pulp subjected to an additional 60 min. aeration cycle at a pH of l0.5 followed by an additional 24 hours cyanidation.
0.65 lbs. Pb(N03)2 was added to the regrind stage.
The CaO consumption in the first and second aeration conditioning stages was 8.4 and 4.8 lbs./Ton respectively.
At the end of the additional 24 hours cyanidation, there was no change in the tailings analysis showing that ~.~7'~P55 1 maximum gold recovery had been achieved in the first 24 hours of cyanidation.
It is obvious that on this ore, a second aeration conditioning circuit would not be used. In plant practice using the single stage aeration conditioning circuit and recycling the barren solution, the CaO consumption would be approximately 6.5 lbs./Ton.
Example 2 The thickener underflow sample was batch reground for 4.5 mins. with the addition of 2.0 lbs. Na2SiO3 to the mill.
The sample was transferred to the Denver Cell and conditioned for 120 mins. at a pH of 10.5. Following the aer-ation conditioning stage, 3.0 lbs. NaCN were added to the pulp and the pulp transferred to the cyanidation rolls for a 24 hours cyanidation treatment.
The calculated head analysis of the ore was 0.58 ozs.
Au/Ton, and the tailings analyzed 0.030 ozs. Au~Ton for a Au recovery of 94.8%.
The RP of the solution at the end of 24 hours cyan-idation was 2.8. Total CaO consumption was 7.9 lbs./Ton.
The tailing analysis of 0.03Q ozs. Au/Ton is remarkablein comparison to the plant results which showed a cyanide tailing of 0.08 to 0.09 for a comparable head analysis and 60 hours cyanidation time versus 24 hours in this example.
Example 3 This test was a duplicate of Example 2 with the exception that 0.45 lbs. (ph(No3)2 was added to the regrind mill and the cyanide was reduced to 1.5 lbs. of NaCN.

The calculated head analysis of the ore was 0.45 ozs.
Au/Ton and the tailings analyzed 0.0315 ozs. Au/Ton for a Au ~7~ 5 1 recovery of 93.0%. Total CaO consumption 7.3 lbs./Ton.
The RP of the solution at the end of 24 hours cyanida-tion was 1.3 showing the effect of the aeration circuit in the combined presence of a lead salt and Na2SiO3, with outstanding gold recovery and highly effective gold dissolution properties of the cyanide solution.
Example 4 -The following series of tests were run to study the effect of the chloride ion.
The thickener underflow samples were batch reground for

3 mins. in the presence of 2.0 lbs. Na2SiO3, and 10 lbs. CaC12, 8.0 lbs. NH4Cl, and ~.0 lbs. KCl respectively.
Following grindingj they were transferred to the Denver Cell and conditioned for 150 mins. at a pH of 10.5.
Following conditioning, they were transferred to the cyan;dation rolls and cyanided for a 24 hours period with the addition of 3.0 lbs. NaCN.
CaO consumption with CaC12 was 11.5 lbs./Ton; with MH4Cl was 17.6 lbs./Ton and with ~Cl was 9.1 lbs./Ton.
The calculated head analysis was 0.40 ozs. Au/Ton.

The results in all tests were nearly identical showing an average t~ilm~s analysis of 0.0285 ozs. Au/Ton for a gold recovery of 92.9~.
~xample 5 ., This example is of a two stage aeration condit~oning system wherein the thickener underflow sample was batch reground for 3 mins. in the presence of 2 lbs. Na2SiO3, conditioned in the Denver Cell for 150 mins. at a pH of 10.5 followed by 24 hours cyanidation on the rolls with the addition of 3.0 lbs.NaCN, filtered and the filtrate recombined with the filter cake, ~ 7(~

1 conditioned in the Denver Cell for 60 mins. at a p~I of 10.5 with the addition of 0.68 lbs. Pb(NO3)2 and 0.75 lbs. Na2SiO3, placed back on the rolls for an additional 24 hours cyanidation period.
The calculated heads were 0.21 ozs. Au/Ton and the tailings analysed 0.0168 ozs. Au/Ton for a Au recovery of 92.0%.
CaO consumption in the first and second aeration conditioning stage was 7.9 and 6.0 lbs./Ton respectively.
The RP of the solution at the end of 48 hours cyanidation was 1.7 showing that all of the barren solution could be recirculated.
Example 6 . .
This example was a repeat of example 5 except the first aeration conditioning cycle was 140 mins. at a pH of 9.5, and the second aeration conditioning at a pH of 9.5, and 0.36 lbs. Pb(NO3)2 added to the first conditioning stage, and 0.15 lbs. Pb(NO3)2 to the second stage, and 2.5 lbs. Na2SiO3 to the first stage. The NaCN addition was 1.5 lbs. NaCN following the first stage, and 0.75 lbs. NaCN to the second stage.
The calculated head was 0.47 ozs. Au/Ton and following 48 hours cyanidation, the tailings analysis was 0.030 ozs. Au/
Ton for a Au recovery of 93.6~.
CaO consumption in the first and second aeration con ditioning stages was 4.8 and 1.0 lbs/Ton respectively.
The RP of the cyanide solution following 48 hours cyanidation was 1.0 showing outstanding Au dissolution character-istics.
Example 7 The thickener underflow sample was batch reground 7.0 mins. with the addition of 0.5 lbs. Pb(NO3)2, and transferred to 74¢~55 1 the Denver Cell. The conditioning time was 60 mins. at a p~
of 9.5, and 80 mins. at a pEI of 10.5. 1.5 lbs. NcCN was added to the pulp and it was transferred to the cyanida-tion rolls for a 24 hour cyanidation period.
Following this period, the sample was filtered, the filtrate recombined with the filter cake, conditioned for an additional 60 mins. at a pH of 10.5 and the addition of 0.3 lbs.
Pb(NO3)2. It was then transferred to the cyanidation rolls for a 24 hour cyanidation period with the addition of 1.5 lbs. NaCN.
The calculated head was 0.27 ozs. Au/Ton and the tailings ana-lyzed 0.025 ozs. Au/Ton for a Au recovery of 90.4%. The RP
of the cyanide solution following 48 hours cyanidation was 1.10.
CaO consumption in the first and second aeration conditioning stages was 8.0 and 7.2 lbs./Ton respectively.
It will be noted that even with the extremely fine grind, the gold recovery was poorer than with those tests that combined the use of dispersant in the aeration conditioning circuit or circuits.

Example 8 The thickener underflow sample was reground for 3 mins.

with the addition of 0.5 gms. CaO and 2.0 lbs. NazSiO3. Follow-ing the transfer of the pulp to the Denver Cell, it was condition-ed for 120 mins. at a pH of 10.5, 3.0 lbs. NaCN added to the pulp and transferred to the cyanidation rolls for 24 hours cyanidation treatment. At the end of this period the pulp was filtered and transferred to the regrind mill with the filtrate as make-up solution and ground for 2 mins. with the addition of 0.6 lbs. Ph(NO3)2. It was then transferred to the Denver Cell and conditioned for 60 mins. at a pH of 10.0 and transferred to the cyanidation rolls for a 24 hours cyanidation treatment.

~7'~ 5S
1 Total CaO consumption was 7.9 lbs./Ton. The calculated head was 0.36 ozs. Au/Ton and the tailings analyzed 0.023 ozs. Au/Ton for a Au recovery of 93.6~.
The surprising effectiveness of my aeration conditioning circuit is in many ways contrary to current accepted technology in the art. For instance, in the use of comparatively lar~e ~uantities of dispersant, that is, as high as 3.0 lbs. Na2SiO3 per ton, theoretically one would expect gold recovery to suffer due to the production of silicate surfaces on various particles of the ore thus inhibiting cyanide contact with the ground particles of ore. However, I have found the opposite to be the case.
In addition, with the dispersing effect of the silicate, the expected result would be thickening problems and dirty thickener overflows. However, I have found that with 24 hours cyanidation following this stage, the thickening characteristics of the pulp are excellent and the thickener overflow solution as clear as without the use of the dispersant.
In the most simplified version of my invention, where I use a dispersant in a primary and possibly secondary or tertiary grinding circuit, as compared with the use of dispersants in the regrind circuit as illustrated by some of the examples and where gold recovery is the major objective, the time factor in the circuit prior to thickening may be too short to bring about flocculation of the ground particles of ore even at my preferred minimum pH of 9.5 wherein lime acts as a flocculant.
For this reason, in addition to maintaining the thickener feed pulp density below 35~ solids, in such cases I prefer to add an additional flocculant to the thickener feed.

My preferred flocculants are selected from the group consisting of guar gums and polyacrylamides. However there are ~.~L74qgS5 1 probably other flocculants that will be applicable and the invention should not be considered as limited to my preferred group.
Although the disclosure describes and illustrates a number of embodiments oE the invention, it is to be understood that the invention is not to be considered restricted thereto, as anyone skilled in the art may find different applications of the invention to the cyanidation of gold bearing ore than as described herein.

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Claims

The embodiments of the invention in which an exclusive property or privilege is claimed are defined as follows:

1. A process for the recovery of gold from gold bearing ore containing sulphide mineral in the group consisting of pyrite, marcasite, pyrrhotite, arsenopyrite and stibnite comprising: subjecting said ore to at least one stage of wet grinding to produce a finely ground product contained in a pulp of said ore and in the presence of at least a dispersing agent and at least calcium hydroxide as alkaline agent and agent selected from the group consisting of chloride salts, lead salts and cyanide salts; subsequently subjecting said pulp to a thickener stage wherein the pH of the said pulp has been adjusted by said at least lime as the alkaline agent to a min-imum 9.5 and the density of the said pulp is maintained at less than about 35% solids to produce a thickener overflow solution and a thickener underflow pulp containing said ground product of said ore; subsequently subjecting at least said thickener underflow to cyanide treatment to dissolve the gold values contained in the said ground product of said ore to produce a cyanide solution enriched in gold values and said ground product impoverished in said gold values; and subsequently recovering the gold values by conventionally practised methods from the said enriched cyanide solution and said ground product is impoverished in said gold values.

2. The process of claim 1 wherein the pH of the said pulp produced in the said at least one stage of wet grinding is maintained at a minimum pH of 9.5 by the addition of at least the said lime as the alkaline agent.

3. The process of claim 1 wherein the said cyanide agent is present during the said at least one stage of wet grinding.

4. The process of claim 1 wherein the said dispersing agent is selected from the family of sodium silicates.

5. The process of claim 1 wherein preceeding the said thickener stage, part of the gold contained in the said gold bearing ore is recovered by gravity concentration and amalgamation.

6. The process of claim 1 wherein the said lead salt is selected from the group consisting of lead nitrate, lead acetate, and litharge.

7. The process of claim 1 wherein a flocculating agent is present in said pulp to said thickener stage.

8. The process of claim 7 wherein the said flocculating agent is selected from the group consisting of guar gums and polyacrylamides.

9. A process for the recovery of gold from gold bearing ore containing sulphide mineral in the group consisting of pyrite, marcasite, pyrrhotite, arsenopyrite, and stibnite com-prising: subjecting said ore to at least one stage of wet grinding to produce a finely ground product contained in a pulp of said ore and in the presence of at least a dispersing agent selected from the family consisting of sodium silicates and in the presence of cyanide salt agent selected from the group consisting of calcium cyanide, sodium cyanide and potassium cyanide and in the presence of at least calcium hydroxide and wherein said pulp is maintained at a minimum pH of 9.0; sub-sequently subjecting said pulp to a thickener stage wherein the

Claim 9 continued density of the said pulp to said thickener stage is a maximum of 35% solids and a flocculant selected from the group consist-ing of guar gums and polyacrylamides has been added to said pulp to said thickener stage and the pH of the said pulp has been maintained by at least calcium hydroxide to a minimum of 9.5 and wherein the products of said thickener stage consist of an overflow solution enriched in said gold and underflow pulp impoverished in said gold; subsequently subjecting said overflow solution to conventional gold recovery and subsequently subject-ing said underflow pulp to conventional cyanidation leaching practice to recover the residual gold in the said finely ground product contained in the said pulp.

10. The process of claim 9 wherein preceeding the said thickener stage the said pulp is subjected to a gravity con-centration and gold recovery stages.

11. A process for the recovery of gold from gold bearing ore containing sulphide mineral in the group consisting of pyrite, marcasite, pyrrhotite, arsenopyrite, and stibnite comprising: subjecting said ore to at least one stage of wet grinding to produce a finely ground product contained in a pulp of said ore; subsequently subjecting said pulp to an agitation conditioning circuit wherein the pH of the said pulp is maintained for at least 60 mins. at at least one optimum pH
point in the pH range of about 8.5 to 12.0 and wherein at least calcium hydroxide is added to the said pulp during said agitation conditioning and said agitation conditioning is carried out with air disseminated throughout the said pulp and in the presence of

Claim 11 continued agent selected from the group consisting of dispersants, chloride salts, lead salts, and cyanide salts to produce a conditioned pulp; subsequently subjecting said conditioned pulp to a con-ventional cyanide gold dissolution and gold recovery circuit to recover the bulk of the said gold contained in the said con-ditioned pulp and the said finely ground product impoverished in the said gold.

12. The process of claim 11 wherein within the said at least one stage of wet grinding a gravity concentration stage is used to recover at least part of said gold in a gravity concentrate; subsequently subjecting said concentrate to amal-gamation and retorting to recover said gold in at least semi-bullion form.

13. The process of claim 12 wherein the said gravity con-centration stage follows the production of the said finely ground product.

14. A process for the recovery of gold from gold bearing ore containing sulphide mineral in the group consisting of pyrite, marcasite, pyrrhotite, arsenopyrite, and stibnite comprising; subjecting said ore to at least one stage of wet grinding to produce a finely ground product contained in a pulp of said ore; subsequently subjecting said pulp to an agitation conditioning circuit wherein the pH of the said pulp is main-tained for at least 60 mins. at at least one optimum pH point in the pH range of about 8.5 to 12.0 and wherein at least calcium hydroxide is added to the pulp during said agitation conditioning and said agitation conditioning is carried out

Claim 14 continued with air disseminated throughout the said pulp and in the pre-sence of dispersing agent selected from the family of sodium silicates to produce a conditioned pulp; subsequently subjecting said conditioned pulp to a conventional cyanide gold dis-solution and gold recovery circuit to recover the bulk of the said gold contained in the said conditioned pulp and the said finely ground product impoverished in the said gold.

15. A process for the recovery of gold from gold bearing ore containing sulphide mineral in the group consisting of pyrite, marcasite, pyrrhotite, arsenopyrite, and stibnite com-prising; subjecting said ore to at least one stage of wet grind-ing to produce a finely ground product contained in a pulp of said ore; subsequently subjecting said pulp to an agitation con-ditioning circuit wherein the pH of the said pulp is maintained for at least 60 mins. at at least one optimum pH point in the pH range of about 8.5 to 12.0 and wherein at least calcium hydroxide is added to the pulp during said agitation con-ditioning and said agitation conditioning is carried out with air disseminated throughout the said pulp and in the presence of dispersing agent selected from the family of sodium silicates and lead salts selected from the group consisting of lead nitrate, lead acetate, and litharge to produce a conditioned pulp;
subsequently subjecting said conditioned pulp to a conventional cyanide gold dissolution and gold recovery circuit to recover the bulk of the said gold contained in the said conditioned pulp and the said finely ground product impoverished in the said gold.

16. The process of claim 14 wherein during said agitation conditioning,cyanide salt agent selected from the group consist-ing of calcium cyanide, sodium cyanide, and potassium cyanide is present.

17. The process of claim 15 wherein during said agitation conditioning,cyanide salt agent selected from the group consist-ing of calcium cyanide, sodium cyanide, and potassium cyanide is present.

18. The process of claims 14 and 15 wherein the said at least one stage of wet grinding a gravity concentration stage is used to recover at least part of said gold in a gravity concentrate; subsequently subjecting said concentrate to amal-gamation and retorting to recover said gold in at least semi-bullion form.

19. The process of claims 16 and 17 wherein the said at least one stage of wet grinding a gravity concentration stage is used to recover at least part of said gold in a gravity concentrate; subsequently subjecting said concentrate to amal-gamation and retorting to recover said gold in at least semi-bullion form.

20. The process of claims 14 and 15 wherein a gravity concentration stage follows the production of the said finely ground product.

21. The process of claims 16 and 17 wherein a gravity concentration stage follows the production of the said finely ground product.

22. The process of claim 11 wherein the said pulp of the said finely ground product impoverished in the said gold is subjected to a second agitation conditioning circuit and the said pulp is maintained for at least 30 mins. at at least one optimum pH point in the pH range of about 8.5 to 12.0 and wherein at least calcium hydroxide is added to the pulp during said agitation conditioning and said agitation conditioning is carried out with air disseminated throughout the said pulp and in the presence of agents selected from the group consisting of dis-persants, chloride salts, lead salts, and cyanide salts to pro-duce a second conditioned pulp; subsequently subjecting said second conditioned pulp to a conventional cyanide gold dissolution and gold recovery circuit to further recover the said gold in the said gold impoverished pulp.

23. A cyanide process for the recovery of at least part of the gold from sulphide bearing gold ores comprising; sub-jecting a pulp of the ore to at least one stage of treatment wherein at least calcium hydroxide is present and the pH of the said pulp is maintained in the range of about 8.5 to 12.0 in the presence of at least one dispersing agent selected from the family of sodium silicates and in the presence of at least one cyanide agent selected from the group consisting of calcium cyanide, sodium cyanide, and potassium cyanide to dissolve at least part of said gold; and subsequently recovering the said dissolved gold by conventional means.

24. The process of claim 23 wherein in the said at least one stage of treatment, a lead salt is present.

25. The process of claim 24 wherein the said lead salt is selected from the group consisting of lead nitrate, lead acetate, and litharge.

26. The process of claim 23 wherein in the said at least one stage of treatment, a chloride salt is present and is selected from the group consisting of calcium chloride, ammonium chloride, and potassium chloride.

27. The process of claims 23, 24 and 25 wherein the said at least one stage of treatment is a wet grinding mill.

28. The process of claims 23, 24 and 25 wherein the said at least one stage of treatment is subsequent to any one grinding stage.

29. The process of claim 23 wherein subsequent to the said at least one stage of treatment, the said pulp's density is a maximum of 35% solids and flocculant has been added prior to a thickening stage.

30. The process of claim 24 wherein subsequent to the said at least one stage of treatment, the said pulp's density is a maximum of 35% solids and flocculant has been added prior to a thickening stage.

31. The process of claim 25 wherein subsequent to the said at least one stage of treatment, the said pulp's density is a maximum of 35% solids and flocculant has been added prior to a thickening stage.

32. The process of claims 29, 30 and 31 wherein said flocculant is selected from at least the group consisting of guar gums and polyacrylamides.