Things which may be of interest to members

[Deano]

I have been involved with gold processing in Australia on a commercial scale for over 40 years.
There are many facets of processing which I have developed over that time and which may be of interest to members, I am retiring from commercial activities and would not like to see this knowledge wasted.
Please understand that I have access to all the equipment and use it on a routine basis, this includes Atomic Adsorption Spectrometry (AAS), pH meters and redox meters for measuring Eh.
I mention the above not for any reason but that having the equipment makes the processing easier and in some cases it is necessary to have the equipment to efficiently carry out the processing.
I will need to put the information in a series of posts due to the sheer scale of the information. I will attempt to keep each post on a single topic.
Often I will be on a remote sits and will be not contactable for days at a time.
I felt that the chemical processes site was the most suitable forum for these posts even though some of them are about smelting processing.
If it is felt that these posts are not suitable for any forum I will not be upset about it.

Topic 1

Chlorine leaching

Equipment required - pH meter, Eh meter

Textbooks say that gold leaching solution requires an Eh of 1000mv to leach the gold.
This is not quite correct, what the solution requires an Eh of 1000mv for is to retain the gold in solution. A lower Eh will still dissolve gold but there will be an equilibrium established such that as more gold is leached a matching amount of gold precipitates out of solution.
If all the gold is dissolved and the solution is filtered then the gold will generally stay in solution for at least hours and depending on how clean the solution is it may be several days.

What the above means is that you want a high Eh for leaching and removal of the leach liquor from any residual solids by filtration when the leaching is finished.

The major problem with chlorine leaching is the outgassing of chlorine from the leach liquor. This will lower the Eh of the leach solution below the 1000mv required to keep the gold in solution.
If the pH of the leach solution is kept between 3.0 and 3.5 the outgassing of the chlorine is kept to a minimum, an odour of chlorine is noticeable but not really strong.
This is not to be taken as an invitation to do this type of leaching where there is no fume removal system in place, even low levels of chlorine will cause health problems.

What you end up doing with the leach is a continual sequence of adding HCl to adjust the pH and then adding hypochlorite to adjust the Eh, hence the need for the meters.
Some form of agitation is required to ensure complete leaching, for large scale a peristaltic pump with suitable hosing can be used, for small scale a magnetic stirrer is used.
A thick walled rotating plastic pressure vessel can be used for batch operation, as there is no outgassing of chlorine in this format there is no need to continually check the leach conditions after they have been initially set.

Note that you should always have a catch tray under all of your processing equipment.

Next post

recovery of gold from acid solutions

Deano

 

[g_axelsson]

Hi Deano and welcome to the forum. (Although you have been a member since February.)
I'm for one finds it interesting to hear about gold refining and reclamation from a little more scientific point of view. Looking forward to the coming postings.

Is chlorine leaching used commercially anywhere today?

Now we have one Deano from New zealand (NoIdea) and one from Australia. 8)

Göran

 

[Deano]

Chlorine leaching

A couple of things I accidentally left out of the last post.

If you are wanting to dissolve a substantial quantity of gold in a chlorine leach you should add 4% sodium chloride to the leach liquor.
This ensures that there are enough chloride ions available to form high levels of gold chloride and also speeds up the leach reaction.

Most gold leaching reactions are a trade off between reagent concentration, temperature and time, more of the first two will speed up the leaching rate.
Generally the leaching rate doubles for each 10° Celcius increase in temperature.
Provided you do not run out of reagent or freeze the leach then you will get all the gold in solution in time.

Each leaching batch is looked at from the point of what is most suitable to the operator.
Reagents and heat cost money, if you are time poor then your time also costs money.
Work out for yourself what is the cheapest leach type for what you are doing, generally small batches of less than 100 grams of gold are run hot with stronger reagent concentrations.

Deano

 

[Deano]

Goran

Chlorine leaching is used commercially only for specific leaches, generally where the gold is distributed throughout a metal matrix and the whole matrix must be dissolved to access the gold.

I started this thread with chlorine leaching as a lot of the posts I read used chlorine leaching for base metal removal ( hydrogen peroxide - hydrochloric acid leaching).
If you are looking for a good way to dissolve gold and have it reprecipitate as ultra fine particles then the H₂O₂ - HCl leach is what you use.

Note that these fine particles are so fine that they will pass through micro filters, they will not be caught on any paper filters.

Another version of this type of leaching occurs when you have a leach liquor containing HCl and iron chloride, formed when you contact iron with HCl.
The Eh of the leach will be about 600 mv, enough to dissolve gold but not enough to retain the gold in solution.
Worse still, a leach consisting of copper chloride in HCl will also form these micro particles of gold.

What the above means is that if you use a HCl based leach for base metal recovery you will be putting some of your gold into a form where the gold can only be recovered by leaching with stronger reagents such as chlorine or nitric acid in the leach liquor.
Usually chlorine as hypochlorite is cheap and readily available so why would you risk gold losses without carrying out this leaching step.
You do not need to do all of your leaching using hypochlorite in the HCl leach, just a final step at the end of the base metal leaching stage.

Deano

 

[Deano]

Recovery of gold from acid solutions

Gold in acid solution is in the form of gold chloride for the most common forms of leaching.
It can be in other forms but you need to have a strong need to leach the gold in these other forms to bother doing so.
Generally leaching gold with other than straight chloride reagents is either more expensive or more toxic than the chlorides.

There are two main methods for recovering gold from acid solutions. Both have their strengths and weaknesses.

The first method is precipitation with a reducing agent.
This method is very simple, it has two steps.

Step 1 is to get the pH of the solution between 1 and 2. I usually run the precipitation at pH 1.5.

Step 2 is to add the cheapest sulfite chemical available as the calculated weight + 20% and allow to precipitate overnight.

Step 1 methods.
If the pH of the solution is greater than 1.5 I bring the pH down to 1.5 with HCl. This is usually needed for a chlorine leach liquor where you have leached at pH 3.
If the pH of the solution is less than 1.5 I bring the pH up to 1.5 with urea and only urea. This is usually needed for an aqua regia leach liquor.
The urea used must be clean, this cuts out most garden and agricultural versions. I use Harnstoff technical grade urea from Germany, the cost is not significant when set against the gold recovered.
Using this method also forces you to keep your leach liquor volumes as small as possible so that the amounts of pH adjusting chemicals are the least possible.

Step 2 methods
I use sodium metabisulfite to get the greatest bang for my buck, you may be able to access cheap batches of sodium sulfite or sodium bisulfite. Food grade as used to clean out beer or wine fermentation vats is OK.
Allowing precipitation overnight is very important if you do not have much gold in the leach liquor (this is called low tenor gold in solution).
The gold precipitated at the start when there is a high gold level in solution will form large particles on the bottom of the beaker.
As the gold level drops the gold precipitated will form smaller particles, towards the end the gold particles are extremely fine.
This means that you will need to recover the gold on a microfilter, I use a 0.45 micron cellulose filter in a tall form micro filter cup when doing larger quantities of gold.
This allows me to filter off about 1000 grams of gold at a time.

An alternative to microfiltration is to use a floc agent to bind the gold particles together to allow use of much larger filter papers.
The best floc agent is Cytec N300 superfloc. This is a food grade neutral floc agent which works beautifully in the strong acid pH range.
It needs to be premixed at the rate of 1 gram per litre of water. Use hot water and a magnetic stirrer, the mixing takes a long time.
I use 100 ml of the premix per kilo of gold on the bottom of the beaker. Add the premix to the beaker with the gold in and use a plastic stirrer such as a paint stirrer in a drill to beat the absolute pants off it for 10 seconds.
The liquor goes all brown for a short time from the gold in suspension but then becomes clear as the gold flocs precipitate.
These flocs can be filtered on any 15 micron or finer filter paper, eg Whatman no 1.

What can go wrong when using a floc agent.
You try to use a beaker which is too small to contain the leach liquor and floc premix under strong agitation.
You think that you will make the premix stronger than 1 gram per liter. Two reasons why this is not a good idea.
1 If you make it a lot stronger you will not be able to stir it and dissolve the N300 in the water.
2 The dilution from adding the premix helps in the floc action.
All the information about using floc agents says to gently stir for a long time to form large flocs. You think that this is what the product experts recommend so you try it. It does not work well.

With all the above methods in place you will get 999 gold from the smelt

Organic extraction of gold from acid liquors

I use di butyl carbitol as the extractant, the full name is di ethylene glycol di butyl ether. Also has a trade name of Ansul E444.
Why do I use this instead of the other extractants available?

At room temperature.
1 I can use it on undiluted aqua regia after gold dissolution.
2 I can load it to the point where it sinks in the liquor, this makes it easier to use in a separatory funnel.
3 When used with the normal ventilation requirements there are no side effects on health after 30 years of use.
4 recyclable indefinitely with oxalic acid reduction of the gold or vacuum distillation recovery of the gold.
5 It is not objectionable to use from an odour perspective.

Disadvantages of using it.
There is a low solubility in the leach liquor, this requires a scavenge cycle contact to recover any gold values associated with this dissolved organic.
This is not a great disadvantage as I always carry out a triple contact sequence to scavenge any gold values.
If the gold loading in the organic is low then the gold recovery from oxalic acid will be poor and difficult to recover as films on walls etc.
There is a low solubility of the organic in the liquor which means reagent loss. Note that the solubility figures quoted for this reagent are not usually achieved in plant practice.

How do you use organic extraction

Have a filtered solution of gold chlorides
Add what is theoretically the required organic volume + 10% and agitate strongly with a plastic stirrer such as a paint stirrer in a drill for 1 minute.
Remove the stirrer and place all the liquor in a separatory funnel.
Allow to sit for about 1 hour for full phase separation.
Unless you really screwed up on your volumes you should have an orange layer on top of the liquor.
If you have a yellow layer rather than an orange layer you have used way too much organic. A yellow layer will cause problems at the oxalic acid stage as the gold grade is way low.
The only way to get out of the yellow layer problem is to contact the organic with more gold chloride solution until an orange layer is achieved.
If you have an orange layer at the bottom of the liquor you have used too small a volume of organic. The upside of this is that you will have easy separation of the organic and high recovery of the gold at the oxalic acid step.
Either way you then put another lot of organic in the leach liquor after the separation step has been completed and repeat the cycle.
The organics from the second and third cycles are kept to be used as the first and second contact organics on the next batch of leach liquor
The first cycle organics are contacted with 50% HCl for base metal recovery and again separated in the separatory funnel before the organic phase is placed on the oxalic acid solution.
The oxalic acid solution is made up as a concentrated solution and the organic is left on it overnight at room temperature (20^C)
Next day the organic layer is recovered in a separatory funnel ready to be the third contact organic on the next batch of liquor.
The gold is filtered and smelted
Done properly this will give you 999 gold.

Deano

 

[nh6886]

Thanks for sharing Deano,

I look forward to reading your contributions.
I have a question for you or anyone who cared to answer. What I have to work with at this time is,
HM Digital ORP-200 Waterproof ORP Meter, -999 to +1000 mV ORP Range, 1 mV Resolution, +/- 0.5%
I was not familiar with "Eh" before I read this, so my question is whether this meter is sufficient for this process.

All the Best,

John

 

[Deano]

John

Eh is like chemists shorthand for a specific ORP (which is an acronym for Oxidation Reduction Potential).

Eh means that the potential is measured against the standard hydrogen electrode, this is the international reference standard.

There are several other standards still being used, in particular in the US the Ag/AgCl reference electrodes still get an outing.
I can only presume that this is so because the users of these electrodes have used them for a long time and are reluctant to update to international standards.

Usually an ORP meter calibrated in mv is reading Eh values, any(rare) which are not reading Eh will have a prominent announcement on the literature.

The bottom of the actual gold chloride stability range is 1000 mv, generally in a chlorine leach you will have an upper limit of 1300 mv, usually in the mid 1200s.
Your meter will tell you if the Eh is high enough for gold leaching but will not tell you how your Eh is varying in the leaching range.
It is not a good idea to be have to wait until the Eh has dropped below 1000 mv so that you can read it before taking action to get the Eh back up over the 1000 mv mark.

I would say to get a meter which reads at least to 1300 mv to make chlorine leaching easier.
The meter does not have to be an expensive model, there are some very good cheap ones around.

When using the meter in a leach do not leave the electrode in the leach solution permanently, you will shorten the life of the electrode a lot.
I usually have the pH and Eh electrodes in a beaker of water along side the leaching vessel and only place the electrodes in the leach when I want a leach reading.
After doing a couple of leaches you get a good idea how long the intervals should be between readings.
The reading intervals can be affected by impurities in the water, however these impurities will not affect the actual leaching as long as the Eh and Ph conditions are maintained.

Deano

 

[Deano]

A further missed point on the recovery of gold from acid solutions

If you have a commercial quantity of base metal such as copper in your solution then you can electrowin the metal after the gold recovery step.
It is difficult to carry out individual metal electrowin runs even with all of the right equipment, it is easiest to carry out a bulk electrowin of all metals.
Usually not much iron comes out in these electrowin runs as it tends to bounce between the valence states of +3 and +2 at the electrodes rather than plate out on the cathode.

The best form of EW cell for small scale electrowinning uses carbon felt as the cathode and a carbon rod as the anode. The carbon felt is sold as soundproofing material in large rolls.

Cell design

Get a piece of plastic pipe about 100 mm in diameter and about 300mm long. Glue in place one end cap.
Get someone to drill as many 3mm holes as possible in the pipe and wrap a layer of carbon felt around the outside making sure that at the end with no end cap you extend the felt past the end of the pipe by 30mm to allow for electrical connection.
Hold the felt in place with teflon thread tape.
Note that you cannot use this trick with gold cyanide solutions as the gold cyanide will dissolve the teflon.

Get a second end cap and cut out a hole in the centre of the cap which will just fit a 25mm diameter carbon rod which is long enough to go from the glued end plate to about 50mm outside the not glued end plate.
The diameter of the rod does not need to be 25mm, it is just what I use. Any diameter between 20 and 50mm will do, depends on what you can get cheaply.
Drill a second hole near the carbon rod hole of a diameter suitable to take a small pump suction hose.
I use peristaltic pumps so there is no contact between the acid liquor and the pump. These pumps are self priming, it won't hurt them to run dry.
There are some cheap peristaltic pumps available with built in speed control, the AQUA brand from Italy has been reliable.

Stand the piece of pipe upright in a container which is about the same height as the pipe and about 100mm greater in diameter than the pipe.
This container can be made from a piece of 200mm diameter plastic pipe with one glued end as the bottom. Standard blue pipe glue works fine.
Drill a 12mm hole in the wall of the outer pipe as close to the top of the wall as possible and glue a short piece of 12mm OD plastic pipe in the hole so that the pipe projects at least 30mm outside the wall.
Use a length of 12mm ID plastic hose placed over the plastic pipe to direct overflow from the large pipe back to a plastic container which holds the leach solution which is to be stripped.

What we now have is a plastic bucket or whatever containing the leach solution which is to be stripped in the cell.
A peristaltic pump pumps the liquor from this plastic bucket into the leaching cell between the two pieces of pipe, a wooden clothes peg holds the discharge line in place.
As the leaching cell fills with the leach liquor some of the liquor will pass through the carbon felt into the inside perforated pipe.
When the leaching cell is full of leach liquor the liquor will start to return to the plastic bucket via the 12mm overflow hose .
This means that the leaching cell cannot overflow without some creative stupidity being applied.

A second peristaltic pump set to run slightly slower than the first pump is used to suck the liquor from between the perforated cell wall and the carbon rod and discharge it into a tailings bucket.

The carbon rod is used as the anode and the carbon felt is the cathode, current is supplied from a constant voltage power source. Usually around 2.5 volts but depends on the actual cell dimensions and the liquor conditions.
Bubbling from the electrodes escapes from the hole that the peristaltic pump suction line passes through.
I use standard spring clamp connections on the anode and cathode. I then connect the power leads to these clamp connections with the clamp connections attached to the power leads.
This way when the inevitable happens and the leach liquor wicks its way up the felt and the carbon rod it can only contact the first clamps, treat these clamps as consumables and get the cheapest available.

When starting a run through the cell, the cell is first filled with an acid solution at pH near that of the leach solution to be stripped.

The power to the electrodes is turned on and then the peristaltic pumps are started.

The cell is run until until the liquor bucket is empty and the liquor level in the cell is down to near bottom.
At this stage a couple of litres of acid water , same as the start water, is added to the bucket to flush the last of the liquor through the cell.

When the flush water has passed through the cell a couple of litres of plain water are run through to remove the acid from the cathode carbon felt.
Now turn off the power and stop the pumps.
Undo the teflon tape and remove the felt containing the metals from the leach solution.

Depending on your recycler you may be able to sell the loaded felt as is or you can use it as an anode in a flat electrode electrowinning cell with stainless steel or other metal cathode.

This form of cell is incredibly efficient and will remove most metals from solution. Iron is the major exception to the effective removal. It can be removed but high overvoltage is required.

When using the cell for the first few times there are some things you need to find out for your solutions.
I usually run pretty standard solutions through the cell and I know what actual voltage and liquor flow rate through the cell I need for these solutions.
When I am about to run a liquor where I am not sure of the metal levels I have the luxury of running the solutions through AAS and adjusting the cell flow rates to suit the levels.
If you do not have this option and most people don't then you have to find out the maximum flow rate of the liquor through the cell by trial and error.

Start running the cell with a flow rate of 1 litre per hour. Run a batch all through making sure to save all the liquor.
Remove the felt and replace with a new piece. Check the level of metal deposit on the felt.
Now rerun the saved liquor, save once again.
Remove the felt and check for metal deposits, compare with the first felt.
If there are metals on the felt you need to slow down the rate of liquor through the cell and rerun the test with fresh leach liquor.
If there are no metals on the second felt then you can speed up the rate of liquor through the cell and rerun the test with fresh liquor.

Generally I can tell how the electrowinning is going by the amps the cell is drawing.
This is much harder to do when you are running batches where the acid concentration and metal levels in solution vary wildly.
In this case I would err strongly on the side of caution and run a much slower flow rate of liquor through the cell. Remember that the power draw for the whole setup is very low, doubling the electrowinning time is not going to hit you finnancially.
In my case where I am running large batches of liquor containing high levels of metal the cost of doubling my cell time would hurt, this is why I use the ASS. It is just a very useful cost effective tool.

If you don't change your felt often enough or are running solutions with a lot of metal in them then you can actually load the felt with so much metal that the liquor cannot flow through the felt.
You generally find this out when you go to check if the processing is complete and you find your second pump is sucking air because no liquor is getting through the felt.
When this happens take the felt off and wash in water to neutralise any acid residual. Give it several rinses in fresh batches of water. All of this wash and rinse water will have to be put through the cell when the new felt has been fitted.

Things that sooner or later will go wrong.

You will clog the felt with metal as above.

You look at the felt after a run and there is little or no metal on it.
Usually you are running a solution with little or no metal in it to start with.
Other causes are poor electrical connections to the electrodes, running the solution too quickly through the cell and having a liquor where in the leaching step you have used up all the acid, this acid carries the current in the cell. No acid, no electrowinning.
Always treat the electrode connectors and felt as consumables, it is false economy to try and get a few more runs out of dodgy materials.

Note that the dimensions given above are not set in stone. You can vary them to suit your own circumstances.
If you are running small batches the pipe can be shorter and of smaller diameter but make sure that there is a minimum gap of 20mm between the carbon rod and the pipe so that you are not drawing liquor through the felt in one area.
The carbon rod can be of lesser diameter but the erosion rate will be much greater, not a concern if you have access to a cheap batch of them.

Deano

 

[g_axelsson]

Very interesting. One advantage of electrowinning that I see is to remove metal salts as a step of treating waste water. Even if it isn't worth a lot in metal value, just making it easier to process the waste is worth a lot.

A lot of what we do in refining is in chloride based solutions. Is there a lot of chlorine given off during electrowinning?
Have you tested using platinized titanium instead of a carbon rod?

Göran

 

[kurtak]

A lot of what we do in refining is in chloride based solutions. Is there a lot of chlorine given off during electrowinning?
Have you tested using platinized titanium instead of a carbon rod?

Göran

I had the same questions

I tried running an experiment winning cell to win the copper from a copper nitrate solution (after cementing silver from silver nitrate) & the carbon anode deteriorated before the winning was done

As a test it was a small cell (1 gallon) with a 6' long X 1" diameter carbon rod as the anode & a 1" diameter copper pipe as the cathode --- it did win some metal before the anode was destroyed

I was then going to try using a titanium anode but read where it needs to be platinum plated or it would end up forming a titanium oxide layer thereby hindering &/or stopping the winning - So :?:

The chlorine question is a good question as well

Kurt

 

[Deano]

Goran

There is always evolution of chlorine when electrowinning from a chloride matrix.

I can only repeat my earlier warning about always carrying out leaching and recovery processing in a well ventilated area.

I use either a fume cupboard or a room with excess extraction capacity and make sure that the extractors are running.

Platinised titanium is the best anode to use, it is also the most expensive anode to buy but in the long term when you are doing a lot of electrowinning it is cheaper than having to keep buying carbon rods.

There are a lot of different grades of carbon available as carbon rod, unless you specify that you are using the carbon rod as an anode you will probably be sold the cheapest grade which is also the softest grade and the fastest to fall apart.

If you can get a really good deal on softer carbon rod and you do not do a lot of electrowinning then these would be my preference.

Using any metal rod as an anode apart from the platinised titanium will either form a non-conductive layer on the surface or consume the anode.
Keep in mind that the anode metal in the latter case will now be in your cell discharge stream so your metal removal efficiency is lowered.
It is possible to configure the cell so that the liquor is discharged through the cathode felt and any metals dissolved from the anode are removed at the cathode, apart from the iron.
I was trying to keep the cell at its simplest and cheapest form, it is a much more expensive build and requires more expensive pumps and a platinised titanium anode in sealed format if you want to use it as cathode discharge type.

I automatically run leach solutions through a cell after gold recovery just to remove the base metals and make the solutions ready for final clean up.

I take the tailings liquor from the cell discharge, check the base metal levels on the AAS to make sure I removed them all and then neutralise the solution with the cheapest alkali I have available to pH 7.
The solution will go brown as all the ferric ions are converted to insoluble ferric hydroxide.
Use the N 300 flocculant solution with the vigorous agitation step to bring the ferric hydroxide into a layer at the bottom of your reaction vessel ( bucket or whatever).
You can filter these flocs out on any paper filter.

As pure ferric hydroxide these flocs can be dried and disposed of in landfill, at least here in Australia.
The solution remaining is neutral water with salt and can be disposed of in the sewer, at least in Australia.

Deano

 

[g_axelsson]

Thanks for the answers Deano.

Actually there is at least one electrowinning process that can be run without chlorine evolution, at least theoretically. 2 CuCl -> CuCl₂ + Cu and the resulting copper(II) chloride can be used to leach copper. It might be hard to develop this into a practical leach method but we discussed it on the forum a little over a week ago.
http://goldrefiningforum.com/phpBB3/viewtopic.php?f=34&t=21454&start=20#p221877
One problem with this method is that you can't increase the voltage too high or you would start to release chlorine. That would put a limit on how fast you can run the cell I guess.

Good to hear about the platinized titanium, as I bought one this autumn but haven't had time to test it yet. I will probably count my production of copper in kilos at most, but since it's a hobby for me, slow is okay.

It is good that you keep the cell design simple so it will be practical for most people. I think I know a few of our members that will test your design so we will probably be seeing reports in a while. At least I hope so. 8)

Göran

 

[solar_plasma]

Even though I also tried high overvoltage, I couldn't smell any chlorine (though I could have been cheated by the fume hood) and there was no visible gas evolving at the anode, only at the cathode, if I remember correctly. I hope I get some time to try it once again soon and let it run for a longer time.

Setting, as far as I remember:
saturated CuCl₂ solution
saturated the CuCl₂ solution with copper
cathode: copper
anode: gold plated non-magnetic base metal alloy (assumed as white copper alloy)
voltages tried: 0-24V, left for some minutes at low voltages 0-2V

In that short time the brown (CuCl) electrolyte did not become green (CuCl₂).
There were visible fluid turbulences near the anode.
Gold plating got dissolved (probably got reduced back somewhere) and visible erosion at the cathode occured.
Gas evolved at the cathode, which I assumed to be H₂.
Copper metal sponge formed at the cathode.

 

[solar_plasma]

@Deano
If you could draw a fast sketch of your cell, it would be easier to understand all details for those of us, for whom English isn't the first language. Thank you for all the great information!

@all
Does anybody know a source for carbon felt in Germany?

 

[Deano]

Bjorn

I hope that the attachments work, let me know if they do not and I will try another method.

Deano

 

[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