Things which may be of interest to members

[Deano]

Bjorn

The carbon felt is very light. It would cost very little in postage to have someone in the US mail you a couple of square metres of it.
It is widely and cheaply available in the US where it is used as a soundproofing material.

The reasons for using the carbon felt as a cathode are that the felt is not attacked by the liquors and that the surface area per unit area is immense so the layer of felt is very efficient at recovering the metals from solution.

Often there will be no or little current flow through the cell, this is generally caused by bad electrical connectors or the solution pH is close to 7 and so there are few current carrying ions if the metal level in the liquor is low. Make sure that there is good current flow at the end of the run to ensure full metal recovery, you may need to add acid to get the current flow.

I have deliberately not given current flows as they are dependent on metal in solution levels, pH levels and cell shapes.
If someone is using a cell which is half the height of the size I gave then the current will be halved.
Similarly the current will vary depending on the distance between the electrodes.
I tried to use pipe sizes which are standard in Australia and should be approximately available elsewhere.

Note that when you perform the neutralising step for iron removal you will co-precipitate any residual metals that you may have failed to electrowin.
This means that if you have used stainless rod as the anode the final liquor will have some chromium and nickel in solution.
These metals will contaminate the iron hydroxide during the precipitation step and may make the precipitates so contaminated that they cannot be disposed of in landfill.
A good reason for using either platinised or carbon electrodes.
Note that both the above will be subject to erosion if high overvoltages are used.
You would only use high overvoltages if you want to recover the iron during the electrowin and it is much cheaper to use the neutralisation trick.

Deano

 

[Deano]

Cyanide leaching of gold

There are a lot of myths about the use of cyanide.
These are usually along the line that the vapours will kill anything within a couple of kilometres.

The fact is that cyanide is a toxic poison the same as a whole lot of other toxic poisons which people use daily without any thought or concern.
The use of cyanide is safe provided the recommended operating procedures are strictly adhered to.
Problems are usually caused by familiarity breeding contempt, laziness or not having an understanding of the proper procedures.
Treat cyanide as a useful tool requiring careful handling. Use gloves and eye protection when handling cyanide pellets or solutions.

The single most important thing is to have really good ventilation. If you look at a gold mine with large cyanide leach tanks you will see workers walking on walkways across the tops of these tanks. What you will not see is workers lying dead on these walkways.
Providing the pH of a cyanide leaching solution is kept above 10.5 there is very little outgassing of hydrogen cyanide gas. Most plants are run at pH 11 to provide a safety buffer.
The above is not recommending that you use a cyanide leach at pH 11 in an enclosed area without full extraction equipment operating.

Most labs carry out cyanide leach tests in enclosed areas but they make sure that the ventilation equipment is of sufficient capacity and operating.

If I was faced with a choice of low level cyanide fumes or low level chlorine fumes I would take the cyanide. That said, I take great care not to have to make such a choice.

Cyanide leaches used in gold mines usually run 1gram per litre of sodium or calcium cyanide in water with lime added to pH 11.
The above has an actual cyanide level of around 1/2 gram per litre or 500 parts per million (ppm)
That level of cyanide will, without adding lime, give a protective alkalinity of about pH 10.5. Lime is added to increase the pH safety level to around 11.
It is not necessary to use lime for the pH adjustment step, you can use caustic soda or similar. Lime is usually the cheapest.

If the ore has sulfates present then the use of lime may cause precipitation of gypsum (calcium sulfate). If there is a lot of gypsum formed you may be forced into using caustic soda.

If there is a lot of copper present in the ore the leach can be made more selective for gold by lowering the cyanide level. This also prevents consuming a lot of cyanide as copper cyanide. The pH is kept at 11 no matter what the cyanide level is. Check the pH regularly during the leach.

If you are leaching gold from the outside of electrical components you can easily see when leaching is completed.
Remember that cyanide is a slow leach and you expect the leaching to take hours if not days depending on the cyanide level, amount of agitation and temperature.

Cyanide needs to have oxygen dissolved in the leach liquor in order to dissolve the gold. Pumping the leach solution through your bath of components will keep the oxygen level up if you have a small drop from the pump discharge line to the bath liquor level. Don't have a large drop which will splash the solution.

When the cyanide leach has finished the cyanide liquor is transferred into a bucket for gold recovery, Make sure the leached components are well rinsed before disposing of them.

At this stage I run the leach solution through a small electrowin cell to get the gold as a metal. You absolutely must have full air extraction operating during this step.
Remember that you are treating an alkaline solution so the starting liquor you put in the cell is pH 11 water.
After you have treated all of the liquor in the electrowin cell and run a pH 11 rinse through the cell you can remove the felt from the perforated pipe and dissolve the gold from the felt with aqua regia.

Recovery of the gold from the aqua regia is done by sulfite precipitation at pH 1.5 as per a previous epistle.

Things to remember

Sooner or later you will somehow accidentally add acid to your cyanide solution. This is not the immediate disaster you imagine, the hydrogen cyanide gas does not come roaring out of the liquor and kill you in 5 seconds.

The outgassing of hydrogen cyanide gas is relatively slow at low cyanide concentrations, you have time to add caustic soda to the solution to get back to pH 11 provided you have caustic soda nearby in a prepared place ready for such a situation. Make sure you have this ready and do not use it for anything else.


Deano

 

[solar_plasma]

If there is a lot of copper present in the ore the leach can be made more selective for gold by lowering the cyanide level.

Now it gets me wondering, if this also works for other gold leaching methods like iodine, iodine/iodide or thiosulfate based. Would they become more selective at lower concentrations? I think I'll reread all documents I've found, especially the charts, but maybe someone can answer this easily.

 

[Deano]

Bjorn

The reagent dilution trick does not work to a notable degree with thiosulfate leaching of copper and gold ores.
Thiosulfate leaches usually use copper in solution as the oxidant for the gold leaching stage.

I have no direct knowledge of lower tenor leach liquors being more gold specific for halide or sulfide leaches.

Certainly for chloride leaches all I got was more rapid consumption of the chlorine and the need for more frequent Eh / pH checks.
Not what you want for ease of processing.

Generally there are really good reasons why some leaches are used commercially and others are not.

Cyanide is the gold industry's leach agent of choice because of it's cheap price, robust performance and the ease of recovery of the gold complexes.
Toxicity is the main reason for non use in some areas.
If you use the Electrowinning cell I described in the post on cyanide leaching you can recycle the liquor with a cyanide addition, just filter out any carbon particles from the anode before doing so. If you are using a platinised anode you do not need to carry out the filtration step.
Gold recovery from cyanide leaches is simple, it loads readily and with high loadings onto activated carbon, it can be direct electrowon if the gold tenor is high enough or it can be zinced out.
Note that the use of zinc fell out of favour when the carbon in pulp process was introduced.
This was due to cost factors, not just the zinc but the conditioning steps required to get a clean liquor in the optimum condition for zincing. Filtration is expensive for finely milled ores.

Of the halide leaches the only one which is economical for use in some large scale processing is chlorine/chloride.
This leach either has the problem of continuous chlorine evolution in open style leaching (there are many clever ways to minimise this chlorine evolution, none really suitable for large scale processing) or if used in sealed reaction vessels it can only be run as a batch process and still has the problem of chlorine release when the vessel is opened.

The upside of halide leaching is the ease of recovery of the gold chlorides, they load rapidly and to high levels on carbon or resin after the Eh is dropped at the end of the leach.
They can also be precipitated out of solution easily with reducing agents or zinc displacement.

The downside of halide leaching is reagent cost and consumption.
Those halides which must be run acid will have a lot of ores where the ore will react with the acid and consume the acid.
Those halides which can be run at neutral pH will generally react with a lot of the ore components and thus consume the expensive reagents.

Thiosulfate leaching suffers from a major defect in that there are no adsorbents which will load the gold thiosulfate complexes to the ready high levels achieved by gold cyanide.
At regular intervals there are announcements that some or other researchers have made a new adsorbent which is the all singing all dancing product to advance thiosulfate leaching.
Note that thiosulfate leaching has not displaced cyanide leaching in the gold industry.
The most used recovery method for gold thiosulfate is still zinc displacement, this is the method which was discarded on cost grounds by the cyanide process.

Deano

 

[Deano]

Smelting of gold concentrates

A lot of research has gone into the fluxes used in fire assaying, the work done by people in the 1800s and early 1900s is impressive.

A general assumption has been made that what works for fire assaying as a flux should also work for concentrate smelting of metals.

There is a major difference between fire assay flux requirements and concentrate smelting flux requirements.
When carrying out a fire assay you are wanting to get all of the ore components into a liquid form so that all parts of the ore can be accessed for metal recovery.
When carrying out a metal concentrate smelt you are effectively wanting to form all of the metal values into a single molten unit, it is more of a metal melt than anything.
If you can get any base metal values to go into the molten flux and not be present in the molten precious metal this is a large bonus.

Effectively smelting is carried out in two types of crucibles.
These are straight clay crucibles and carbon containing crucibles.
The carbon containing crucibles generally are two types, silicon carbide and graphite.

The carbon containing crucibles are less subject to attack from liquid borax than the straight clay crucibles.
Because of this factor they are usually favoured for gold smelting for two reasons.
1 They can be reused many times.
2 They are less likely to rupture and put your metal values at the bottom of your furnace.

The carbon removed from the crucible by the borax mostly remains in the slag and can be seen as a coating on top of the slag after pouring.

What is generally not known is that the carbon type crucibles will remove much less base metals from a smelt than will the clay type crucibles, the carbon in the smelt slag interferes with the base metal removal.

If you have low base metal levels in the smelt concentrate feed the carbon type crucibles will remove very little of these base metals and they will report to the gold bar giving you a lesser purity than hoped.

I have been unable to improve in a significant way the purity and recovery % of precious metals when using a carbon type crucible.

Clay crucibles suffer from one major defect, this is the rapid attack on the crucible by molten borax. These crucibles should always be treated as single use crucibles which are then discarded.

These crucibles do have one major upside which is the recovery of higher levels and purities of gold in a smelt.

This is achieved in the following manner.

For a smelt of less than 100 grams of metal a 30 gram crucible is used.
This crucible is placed inside a crucible which is the next size up and will take the smaller crucible inside it. The smaller crucible does have to slide down to the base of the larger crucible, there must be a large enough gap between the small and large crucible so that your tongs can remove just the smaller crucible for the final pour.
The larger crucible acts as a catch vessel if the borax in the smaller crucible eats its way through the smaller crucible walls.
I presume that a lot of crucible suppliers will have crucibles of varying quality, wall thickness and size.
Try what you have available for sizes etc.

Always use face protection and thermal gloves when smelting, a full length leather or thermal apron should also be worn.
Heat the crucibles to 1250⁰C if you have a temperature readout, if not heat until white hot.
Take the crucible combo from the furnace and fill the smaller crucible with borax which has been pre-dried in a steel tray in the heat from the furnace. This pre-drying is very important as you are putting the borax in a high temperature crucible, any moisture will cause a violent eruption of molten borax in your direction.

Having safely put the dried borax in the crucible you replace the crucible in the furnace and continue heating.
The borax is slow to melt, depending on how much the crucible cooled down and how strong a heat source you have it will take from 15 to 30 minutes to fluidise all of the borax.
The molten borax is a darkish brown colour when first fully melted.
When the temperature of the molten borax nears 1250⁰C there is a noticeable colour change to a light brown colour.
Give the heating another 5 minutes and then remove the crucible combo from the furnace and pour the concentrates you wish to smelt into the molten borax, takes about 10 seconds for a 100 gram charge. Once again the concentrates must be pre-dried before adding to the crucible.

Place the crucible combo back in the furnace for another 15 minutes.
Take out and pour into a heated angle sided mould which has had the inner walls covered in a layer of white chalk, swirl the crucible before pouring.

There are three things which you must heat to absolute dryness before using as you are going to contact these three things with extremely hot material.
These are the borax, the concentrates and the pouring mould.

What can go wrong

If for any reason you do not absolutely dry your components you will be grateful you are using a head covering full face mask, long arm gloves and body apron. Molten borax is incredibly corrosive but with all the safety gear on you will survive even if scarred. This applies to any smelting operation, not just the one above.

If you go to pour the smelted gold and find that the small crucible has ruptured and most of the borax is in the larger crucible this is not a problem. Usually the rupture level is well above the base of the small crucible so the gold is still retained in the small crucible. Do not stand there looking at the crucible and wondering what happened, just do your pour before the crucible cools and will not release all of the gold.
If there is no metal or borax in the smaller crucible then the rupture occurred at the small crucible base and all of your values are in the larger crucible. Immediately pour the contents of the larger crucible into the mould before cooling occurs.

Do not reuse any crucible which has come into contact with molten borax. You will be inviting crucible failure.

If all of your crucibles are rupturing either you or your supplier have had them on the shelf for too long or the crucibles are not of reasonable quality, change your supplier.


The dried concentrates which are poured into the molten borax must be finely divided or as very thin films. Precipitates from chloride leaching are ideal feedstock provided you have not left them clumped together from the filtration step.
Generally you can expect to add another 9 on your purity by smelting as above.
Concentrates which might run say 95 % gold in a carbon crucible smelt will run 99 +%, Cons which might run say 99% gold in a carbon crucible smelt will run 999+%.

The above has really no commercial effect here as pretty well all gold produced is sold to refiners and you will pay the same refining fee if your bars are 60% gold or 9999 gold.
I just developed the method out of curiosity.


Deano

 

[Deano]

One of the more enduring myths in gold processing is that any gold reporting to the slag in a gold smelt of the cathode material from electrowinning the gold strip solution of the carbon used in a CIP circuit can be successfully processed by throwing the slag into the mill and leaching the resulting product in the cyanide circuit.

If the slag is passed through a hammer mill and the crushed slag is then passed over a wilfley table a disconcertingly large amount of fine gold particles report to the table concentrate.

If this concentrate is leached in cyanide or aqua regia only a fraction of the gold dissolves. This does not fit in with the reprocessing of slag approach used by the gold industry.

It appears that the surface passivation on the gold particles not only prevents leaching but also is the reason for these particles not entering the main body of metal in the smelt.

The only way in which these gold particles may be successfully processed is to use these particles as the feed in the previous post " Smelting of gold concentrates".

This was the method I developed for recovery of the slagged gold, the fact that it improved the gold purity was a bonus.

Deano

 

[Lou]

What do you mean surface passivisation of gold? Under what conditions???


Regarding the cathode from cyanide plating--can usually get down to 10-15 ppm Au with steel wool cathode, stainless anode, pH 12+; run until much O2 fizzing.
Incinerate the cathode (it will be brown like gold coming out) and redissolve in aqua regia; filter to remove AgCl; precipitate, wash gold, 9995%+

Lou

 

[Deano]

Lou

A lot of the material I work with is the product of mine CIP carbon stripping circuits.

Basically the electrowin is very similar to what you wrote but we always have some copper and lead on the cathode wool. Depending on the mine there may be up to 20 kg of gold per strip cycle. Not surprisingly the mine management want to get this fairly portable gold on wool product into a single less portable bar as quickly as possible. As soon as it is dried after aciding out the residual steel wool it is smelted.

All gold mined in Australia is sold to refineries such as the Perth Mint. There will be a refining charge for each bar sent to the refiner along with various other batch and sales charges.

There is the same charge per ounce for refining if the precious metal levels are 80% or if they are 9999. No mine is going to attempt higher purity in the bars when it will cost more and there is no discount on the refining charge.

The surface passivation on the gold beads is caused by a coating of lead minerals. This coating is only a few hundred angstroms thick and is not detectable by optical methods, I had to use XRF.

The coating is not soluble in any solvent in the classical chemistry line and believe me I tried them all. You get a little depressed when you boil the treated beads in aqua regia or cyanide and nothing happens. A few hundred times.

The only way I found to recover the gold from these beads is with the borax smelt in clay crucible trick.

Deano

 

[Deano]

Lou

I was rushed on the last reply and left out a broader description of carbon stripping, my apologies.

Most mines in Australia use a hot caustic cyanide strip on their carbon from the CIP/CIL circuit. If a slight overpressure is used in the strip columns with a pressure drop chamber at the entrance to the EW box then the strip time can be substantially reduced as the temperature of the leach can be run slightly above 100⁰C. The change can be as much as going from a 24 hour strip to a 6 hour strip. This effect is used to increase the capacity of a strip circuit without having to install greater power rectifiers and more EW cells etc.

We usually do a HCl contact of the carbon before the cyanide strip, this gets rid of a lot of the base metals and improves the gold recovery from the carbon.

In a lot of mines we have difficulty getting the gold on carbon level below 100 ppm. This is usually due to high levels of base metals in the ore.

Usually we are trying to get the gold on carbon level down to less than 50 ppm, preferably less than 30ppm.

In most stripping circuits the gold on carbon levels have been well established as a function of stripping current, there is a definite drop in current as the gold level drops.

There is not a great emphasis on getting all of the gold from the strip liquor as the used liquor is later used as make up liquor for the leach circuit. Provided the gold on carbon level is about right and the strip liquor tenor is less than 20ppm gold then most mines are happy.

If you put the recovery into context, you are getting say $500,000 of gold from a strip. Actually carrying out this strip might cost you $ 1,000. There is not a great incentive to be more efficient in the strip process.

Usually a fair amount of the gold will fall off the cathode and form a sludge layer on the bottom of the cell. This gold has to be drained from the cell, filtered and dried before being added to the other gold before smelting.

The steel wool is acided out from the gold cons with HCl. An incredible amount of gold is dissolved in this step, the acid solution and rinse solutions are usually poured through a plastic drum full of carbon for recovery of this gold.

The gold cons after aciding are dried, chopped up finely with a steel scoop and are ready for smelting.

There are as many methods of actually doing the smelt as there are mines.
Some mines heat the crucible to temperature before adding the flux, they get the flux up to temperature and then add the gold.
Other mines will heat the crucible to temperature and then add the flux and gold as a single charge.
Still others will put the flux and gold charge in the crucible before starting heating.
Any variation of the above is being used somewhere.

Similarly every mine has a pet flux formula, usually set by the first gold room superintendent and not changed as an article of faith ever since.
If the gold charge is split into two before smelting there is usually no difference in gold recovery between the formula flux and straight borax. Takes a bit of the mystique out of it all.

After the carbon has been stripped and a sample retained for analysis the carbon is usually run through the regeneration kiln to restore adsorption efficiency. A lot of times it doesn't need the full regeneration but is improved by a quick clean in the kiln.

Deano

 

[g_axelsson]

The steel wool is acided out from the gold cons with HCl. An incredible amount of gold is dissolved in this step, the acid solution and rinse solutions are usually poured through a plastic drum full of carbon for recovery of this gold.

Have you ever tried to precipitate this gold with SO₂ (or sodium meta bisulfite)? Afterwards the really low level gold bearing fluid could be treated in a "stock pot" as being described in many places on this forum or in Hookes book, then discarded as waste (properly).

As I'm not familiar with the running of a CIP facility, I suspect there is a (unknown to me) reason to do it your way.

Göran

 

[Deano]

Goran

Think of processing in a gold room as industrial scale but carried out by people with little understanding of the processes involved. Would you really want someone like this carrying out a precipitation and filtration stage with the tying up of the equipment when the liquor can just be poured into a drum of carbon.

This carbon is stripped with the other CIP circuit carbon when the gold level is high enough, if you have gold metal showing on the carbon you have left it too long before stripping.

The liquor from the carbon adsorption drum has effectively no gold in it and is disposed of in the tailings dam, 100 litres in 10,000 tons of water does not constitute a hazard.

The major problems in the gold room relate to processing and theft.

The processing problems are overcome by following a standard set of simple procedures, no intellectual input required.

The theft problems are minimised by having the gold in a portable form for the shortest time possible, having the gold on carbon does not rate as portable gold.

Always remember that high purity of the bullion is nice to achieve but not an economic requirement at this scale.

Deano

 

[g_axelsson]

Ahhh... the people factor. :mrgreen:

Even Hoke has a section on losses due to dishonesty.

Göran

 

[Deano]

Water sourcing

Many mines use water pumped from an underground aquifer for make-up water. This water is usually stored in a dam until required. There are often two problems with doing this.

One is the growth of algae in the dam water to the point where it interferes with the use of the water.
The algae may be removed by two methods.
The first method is to stock the dam with carp at a rate of one fish to about 200 tons of water, this rate is dependent on the nutrient load in the water, temperature and surface to volume relationship. Make sure that only one sex of fish is used if more than one fish is required.
Many authorities will not allow dams to be stocked with carp so the second method has to be used.

This involves spraying the surface of the dam with a ferric salt solution, usually ferric chloride. This must be done using all plastic equipment including pumps and spray nozzles. You are aiming to cover the surface of the water only, not turn the dam into an acid bath.
When the ferric chloride turns into ferric hydroxide in the dam water there are short lived hydroxyl radicals formed.
These radicals are toxic to algae, in fact to most small organisms including bacteria.
The spraying must be done in daylight so that the algae are present in the upper layers of the water.

The second problem with the dam water is that many aquifers have substantial levels of ferrous salts present. When the water is pumped into the dam a slow reaction with dissolved oxygen will convert the ferrous salts into ferric salts.
These salts will precipitate out slowly forming a coarse sand. Unfortunately the precipitation occurs not only in the dam itself but also in the pipework from the dam to the mine plant.
This causes blockages in the pipework and valves, a surprising amount of this sandy material is formed.

To stop the sand formation the aquifer water should be sprayed into the dam in a fan jet so that the water is saturated with oxygen from the air. This causes the ferrous to ferric reaction to happen much faster and cuts out the slow sand formation.

The above is applicable to farm situations as well as mines.

Deano

 

[Deano]

Thiourea leaching

Many people try different methods of leaching gold out of curiosity and general interest. These efforts are to be applauded, even failures add to the sum of knowledge.

Effectively there are two types of gold leachants, organic and non-organic.
The non-organic leachants comprise the halide complexes and some of the more esoteric leachants such as selenic acid.

These non-organic leachants all suffer from one or more of the following defects.
1 They must be run under acid conditions.
2 They have serious health and safety issues.
3 They are expensive to purchase let alone use.
4 Complex chemistry.

The organic leachants suffer from one or more of the following defects.
1 Toxicity.
2 Reagent degradation.
3 Complex chemistry.

When thiourea became of greater interest as a precious metal leachant back in the 1980s a lot of work was done by many groups to optimise operating conditions.
Unfortunately all of these groups failed to recognise that the major cause of leach difficulty and high reagent consumption was reagent degradation.

Thiourea or its oxidised form called formamadine disulfide, when in soultion, is attacked by ultraviolet light and degrades to elemental sulfur.
The degradation rate increases as the solution Eh decreases and/or the Ph increases.

Thiourea as such will not complex with gold, it requires an oxidant to form formamadine disulfide (FDS). This is what actually complexes with the gold.
FDS is formed when an oxidant with an Eh of greater than 300mv, preferably greater than 350mv, is added to a solution of thiourea at levels so that the Eh is maintained in the solution.

Most testwork carried out with thiourea has used ferric salts under acid conditions as the oxidising agents. All sorts of clever Eh balancing acts have been used to minimise the losses of thiourea in the leach solution, what no one has done is to carry out the experiments in the absence of uv light.

Thiourea can be used under alkaline leach conditions with oxidisers such as hypochlorite, however the thiourea is very sensitive to uv degradation at these higher Ph ranges.
If you want to run thiourea under alkaline conditions you must do so in the absence of uv.

Keep in mind that thiourea is rated as a carcinogen and as such is not recommended for general use outside of specialist labs.
The gold thiourea complex also suffers from poor loading levels on carbon and resin, it needs to be electrowon or zinced out of solution.

Thiosulfate and other organic form leaches can be made more robust and employ simpler oxidants if used under low or no uv conditions, remember that they also suffer from extraction difficulties.

Deano

 

[kurtak]

This is some very interesting reading --- thanks for posting it Deano --- & please keep posting as I look forward to continued reading 8) :!:

Kurt

 

[Lou]

Thiourea cannot be used under basic conditions at all.

I used to elute PGM and Au thiourea complexes off of a macrocylic PS resin functionalized with thiouronium groups (Purolite S920 IIRC).

The options at that point were to use borohyride and save your thiourea, or if really concentrated, you could bring up the pH and boil and destroy your thiourea. Precious metal sulfides precipitate in quantitative yield and may then be worked up in aqua regia.

 

[Deano]

Lou

The standard elution for gold from strong base resins is thiourea in sulfuric acid solution where the sulfuric acid supplies the Eh required to complex the gold from the cyanide or chloride form into the gold thiourea complex. This is usually carried out as either a multi-stage series of contacts or as an extended flow through contact. Either way you get a strong acid solution of gold thiourea which, as you pointed out, can have the gold extracted in a multitude of ways.

If you raise the pH you destroy the Eh as the sulfuric acid is neutralised. At this stage the thiourea complex is very sensitive to degradation by uv and will precipitate elemental sulfur in minutes. If the solution is not accessed by uv there is no precipitation of sulfur.

An alkaline solution of thiourea in alkaline hypochlorite solution and exposed to uv is stable only for a few minutes but will leach gold in that time. If the uv is not allowed to contact the solution then the solution is stable as a gold leachant for a long time.

I have kept some of these solutions in a uv free area for over a year, they are crystal clear when they are then exposed to uv and last only minutes before elemental sulfur is precipitated.

None of the above means that I use thiourea as a leachant for gold, the health concerns and difficulty in recovering the gold values as a concentrate on an adsorbent see to that.

Deano

 

[Deano]

Gold complex adsorbents

In industry there are two adsorbents used for collecting and concentrating dissolved gold values.
These are activated carbon and ion exchange resin.

Activated carbon is the gold industry's adsorbent of choice in that it is cheap and easily handled through the loading and stripping cycles.
The downside of carbon is that the gold loadings from mine leach tanks are relatively low. This means that the carbon must be recycled through the load and strip cycles at a fair rate to recover all the gold available, you do not want to leave the carbon in the circuit for a long time in an attempt to raise the gold loading on the carbon as the attrition losses will increase.
Carbon will also adsorb most metals in the leach solution, especially copper and lead, these can give problems requiring extra steps in the stripping cycle as well as lowering the number of loading sites for gold in the adsorbence stages.

The elution of gold from carbon in a mine situation will always leave some low level gold irretrievably locked in the carbon, this gold can only be recovered by ashing, leaching the ash with cyanide and loading the gold onto fresh carbon.

Problems with using carbon relate to attrition of the carbon resulting in the generation of fines which carry gold values into the tailings dam, the time it takes to strip the carbon and the need for regeneration of the carbon in a kiln on a regular basis. Generally the industry uses a hot caustic cyanide strip solution for carbon but many variations are available.

Ion exchange resin is really only used widely in Russia and neighbouring countries.
The resin can load to higher gold levels than carbon and generally does so more quickly than carbon.
A lot of resin manufacturers have worked on providing a resin which is gold specific and thus will not load appreciable quantities of base metals. Generally they have been relatively unsuccessful, there appears that there is a trade off with ease and level of gold eluted in the strip cycle.

Most commercial resins used for gold cyanide recovery are functionalised with quaternary ammonium groups, these are called strong base resins and are commercially stripped with a sulfuric acid /thiourea solution.
Purolite have a gold resin which uses mixed non quaternary groups, this can be eluted with caustic cyanide solution (yay).

The pros of using resin are the faster and higher gold loadings obtained and the speed and ease of the strip cycle.

The cons of using resin are the high initial cost, the need for finer screens in the gold plant, the level of gold lockup in the resins and the effect of osmotic shock on the resins.
After a number of cycles the gold level locked up in the resin is so high that this inaccessible gold represents a substantial proportion of the gold inventory in the circuit.
This locked up gold also greatly lessens the amount of gold which can be adsorbed and desorbed per cycle as there are less adsorbance sites available.
Like carbon, this locked gold can only be recovered by ashing the resin and leaching the ash.

Generally the resin beads are less subject to wear than carbon particles. Unfortunately they suffer from osmotic shock which carbon does not suffer from.
Osmotic shock is the swelling and contraction which the beads go through when the liquor in which they are immersed changes from acid to alkali or reverse.
Going from strong alkali to very strong acid conditions and back again will ensure that the beads suffer osmotic shock and will start to physically fall apart into smaller particles.
This means that resin and thus entrapped gold losses are substantial even with extra screening.


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Organics destruction

There are often requirements for the removal of organics from a water matrix. This can vary from cyanide destruction to dye oxidation and a general removal of a large number of of other organics which are wanted to be removed for health or other reasons.

The destruction of these organics can be achieved by reversing polarity of the electrowin cell detailed earlier. This means that the carbon felt will now be the anode and it will be generating very high levels of hydroxyl radicals. These radicals are short lived but very highly oxidising, the passing of the solution through the felt ensures that the organic particles will come into contact with these radicals.

There are two controls to the process, one is the flow rate of the liquor through the cell and the other is the rate of generation of the hydroxyl radicals.

There must be current flowing through the call for the radicals to be generated. This current can be carried by by acid or alkali in the liquid or by a neutral, non active salt such as sodium sulfate in which case the current is carried by water splitting at the electrodes.

Generally as long as the electrodes have gas evolution from them then there is enough current flowing to generate the hydroxyl radicals.

Due to the sheer number of unknowns in any solution such as organic type, concentration etc. I cannot give a general current level for any liquor.
I have always adjusted pH and/or salt in solution so that I had gas evolved at the electrodes, easiest seen at the now cathode in the centre of the cell.
I then adjusted the liquor flow rate through the cell so that all of the organics were destroyed in a single pass through the cell. You can use low gas evolution with slow flow rate or high gas evolution with faster flow rate.

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Organics destruction

Once again some things I left out.

The felt cell run as per previous instructions will, by by removing organics from water, remove offensive odours.

It will also sterilise the water but the sterilising process is not an extended effect. The water is sterile only until it exits the cell. At this stage it can be recontaminated.

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