Further things which may be of interest to members
- Thread starter Deano
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The process of vat leaching can be run in virtually any scale.
It is not necessary to have vats holding tens of thousands of tons of ore.
I have run small vats holding 10 tons or so of concentrates, the recovery is always superior to standard tank based systems.
Similar systems can be used for E-waste recovery of precious metals.
Deano
It is not necessary to have vats holding tens of thousands of tons of ore.
I have run small vats holding 10 tons or so of concentrates, the recovery is always superior to standard tank based systems.
Similar systems can be used for E-waste recovery of precious metals.
Deano
Thanks for the pictures, very interesting!
Göran
Göran
When gold which has been dropped from a solution which has a large range of complexed metals in it is being prepared for smelting there are some cases where using the adjust pH of the pregnant solution to 1.5 trick will still not give a clean gold drop.
These cases usually but not always involve lead complexes which do not respond to the usual lead removal steps.
Usually but not always the lead complexes are Lanarkite, a lead oxide lead sulfate complex.
The gold precipitated from these solutions does not come together well during the smelting stage and even the addition of borax will still allow the formation of small gold beads apart from the main gold pool.
The precipitated gold can be cleaned up before smelting by an extended boil in 50% HCl, the boil is kept going until the gold clumps in the liquid.
In some cases this boil can take up to 1 hour to finalise.
A 200 gram sample of precipitated gold from a known problem ore was split into 2 batches of 100 grams each.
One of these samples was smelted in a clay pot without the HCl treatment, the other sample was treated to a 1 hour boil.
The untreated sample formed what was best described as a mass of microbeads of gold, the addition of borax allowed the gold to coalesce with only a few large beads out of the pool.
The treated sample coalesced rapidly and did not require any borax as a pooling aid. A borax sample of the same weight as as added to the first sample was added to the treated sample for comparison purposes.
Smelting was done in a clay crucible retained in a graphite crucible as a catch vessel.
The graphite crucible was mounted on a piece of ceramic plate as a protection for the furnace floor. Some graphite crucibles, under heat, exude liquid which acts similarly to borax in that it bonds the crucible to the surface on which the crucible is sitting.
View of boiled gold in HCL gold bar in pure borax slag. Slag is actually colourless, the colour in the photo is reflected from the gold.
View of untreated gold bar in slag, note prills on right and the presence of base metal in the slag.
Cleaned gold bar with slag removed
Clay crucible inside graphite crucible mounted on a piece of old ceramic plate.
Gold losses into the bar from the untreated gold were 0.67 grams from 100 grams, this was recovered as prills in the slag.
Gold losses into the bar from the treated gold were 0.02 grams from 100 grams.
The solution used to boil the gold in had a surprisingly high level of gold. 100 grams of gold material in 500 ml of 50% HCl reported 50ppm gold in the solution, this was recovered for reprocessing.
Reboiling the gold in barren HCl gave 2ppm gold in solution.
A third boil in barren HCl on the same gold gave 0.2ppm in solution.
It appeared that the base metal dissolved from the gold was capable of dissolving appreciable levels of gold in 50% HCl. As the level of base metal dropped with reboils in barren solutions, so did the level of gold dissolved.
Deano
These cases usually but not always involve lead complexes which do not respond to the usual lead removal steps.
Usually but not always the lead complexes are Lanarkite, a lead oxide lead sulfate complex.
The gold precipitated from these solutions does not come together well during the smelting stage and even the addition of borax will still allow the formation of small gold beads apart from the main gold pool.
The precipitated gold can be cleaned up before smelting by an extended boil in 50% HCl, the boil is kept going until the gold clumps in the liquid.
In some cases this boil can take up to 1 hour to finalise.
A 200 gram sample of precipitated gold from a known problem ore was split into 2 batches of 100 grams each.
One of these samples was smelted in a clay pot without the HCl treatment, the other sample was treated to a 1 hour boil.
The untreated sample formed what was best described as a mass of microbeads of gold, the addition of borax allowed the gold to coalesce with only a few large beads out of the pool.
The treated sample coalesced rapidly and did not require any borax as a pooling aid. A borax sample of the same weight as as added to the first sample was added to the treated sample for comparison purposes.
Smelting was done in a clay crucible retained in a graphite crucible as a catch vessel.
The graphite crucible was mounted on a piece of ceramic plate as a protection for the furnace floor. Some graphite crucibles, under heat, exude liquid which acts similarly to borax in that it bonds the crucible to the surface on which the crucible is sitting.
View of boiled gold in HCL gold bar in pure borax slag. Slag is actually colourless, the colour in the photo is reflected from the gold.
View of untreated gold bar in slag, note prills on right and the presence of base metal in the slag.
Cleaned gold bar with slag removed
Clay crucible inside graphite crucible mounted on a piece of old ceramic plate.
Gold losses into the bar from the untreated gold were 0.67 grams from 100 grams, this was recovered as prills in the slag.
Gold losses into the bar from the treated gold were 0.02 grams from 100 grams.
The solution used to boil the gold in had a surprisingly high level of gold. 100 grams of gold material in 500 ml of 50% HCl reported 50ppm gold in the solution, this was recovered for reprocessing.
Reboiling the gold in barren HCl gave 2ppm gold in solution.
A third boil in barren HCl on the same gold gave 0.2ppm in solution.
It appeared that the base metal dissolved from the gold was capable of dissolving appreciable levels of gold in 50% HCl. As the level of base metal dropped with reboils in barren solutions, so did the level of gold dissolved.
Deano
Deano, you said "These cases usually but not always involve lead complexes which do not respond to the usual lead removal steps. Usually but not always the lead complexes are Lanarkite, a lead oxide lead sulfate complex."
So is there no recognized steps to take to get rid of most of those lead complexes? And if not, is it because they are a combination of an oxide and a sulfate? Or something more complex?
So is there no recognized steps to take to get rid of most of those lead complexes? And if not, is it because they are a combination of an oxide and a sulfate? Or something more complex?
Lanarkite is a lead mineral occurring naturally in some mines.
It is rare enough that no published work has been done on its dissolution.
The difficulty in dissolution is, as you suggested, related to the two forms of lead in the one mineral.
The few references to it in the literature offer a small range of solvents, what you find if you try these solvents is that they are only good for a partial dissolution with the bulk of the material still being residual.
The lanarkite is unusual in that it forms a layer on precipitated gold which is only angstroms thick.
Visually it is not detectable in this form on gold, its presence can only be determined by the behavior of the gold or by high intensity XRF examination.
I only found the HCl boil method by a lot of trial and error testing.
What I did find was that the HCL boil technique was capable of removing virtually all base metals from precipitated gold.
On larger scale applications any dissolved gold in the HCl solution was separated from the solution containing the base metals by organic extraction.
Usually in these cases only a single boil contact with the gold was required to clean it up enough for smelting.
The method will clean up precipitated gold from all origins including ewaste.
It can be applied in cases where you have not been as diligent in preparing a starting material for leaching as you might have been.
If your precipitated gold shows any sign of beading during smelting it can be cleaned by this process.
Deano
It is rare enough that no published work has been done on its dissolution.
The difficulty in dissolution is, as you suggested, related to the two forms of lead in the one mineral.
The few references to it in the literature offer a small range of solvents, what you find if you try these solvents is that they are only good for a partial dissolution with the bulk of the material still being residual.
The lanarkite is unusual in that it forms a layer on precipitated gold which is only angstroms thick.
Visually it is not detectable in this form on gold, its presence can only be determined by the behavior of the gold or by high intensity XRF examination.
I only found the HCl boil method by a lot of trial and error testing.
What I did find was that the HCL boil technique was capable of removing virtually all base metals from precipitated gold.
On larger scale applications any dissolved gold in the HCl solution was separated from the solution containing the base metals by organic extraction.
Usually in these cases only a single boil contact with the gold was required to clean it up enough for smelting.
The method will clean up precipitated gold from all origins including ewaste.
It can be applied in cases where you have not been as diligent in preparing a starting material for leaching as you might have been.
If your precipitated gold shows any sign of beading during smelting it can be cleaned by this process.
Deano
There are three levels of gold reclamation from e-waste.
The first is the hobby person who is interested in the chemistry involved and who wants to produce nice looking high purity product. Generally works with gold quantities up to 100 grams per batch.
The second is the hobby person who is interested in operating a larger process capable of treating material in gold batches of up to about one kilogram. Usually but not necessarily has refining stages in the process stream.
The third is the commercial operator who has pretty much sorted out the cyanide or other lixiviant leaching and recovery of the precious metals in gold batches of kilograms. May do their own refining stages or may sell to a refinery depending on what relationships have been developed with the refinery. Many of these operators are licensed to mark and sell the gold produced in their own right.
Generally the processes used by each level of processing are different but some crossover occurs.
The first level hobby operator is in it with the aim of satisfaction from a process carried out well, the value of the gold is nice but is not necessarily the main driver.
Generally the hobby operator has less commercial type equipment such as commercial type fume hoods and condensers and tends to compensate for this by being very careful with the chemicals and processes being used. These operators also tend to be less familiar with chemistry but still experiment more than the higher level processors.
Most interesting developments chemically and mechanically come from these operators, they deserve all the encouragement they can get.
The second level operators are separated from the first level operators by both the quantities of gold recovered and the processing equipment they use.
This requires a reasonable investment in processing equipment in order to treat the quantities of material involved.
Many of these operators base their process pathways on those of commercial operators, they work on the principle that these must be the the best methods because the big operators use them.
This is not necessarily correct, many of these process pathways are only viable for large scale processing.
It is usually useful for a second level operator to scrutinise their processing in detail to see if what they are doing is actually the best method for their own use.
One of the easiest mistakes to make is to not fully cost their own time and thus look for more efficient processing options to lessen their time.
One step I have previously used is in the precipitation of gold from chloride solutions with sodium metabisulfite.
If the metabisulfite is added as a solution then a greater liquid volume will need to be filtered at the end of the process, this will use up some more time.
If the metabisulfite is added as a powder then time is used when the additions are made under agitation, you don't want to over froth.
If you dampen the metabisulfite and put it into plastic moulds used for lolly making or similar you can get dry pellets about the size of apricots.
If you add these pellets to your beaker of gold chloride one at a time the pellet will start fizzing near the surface of the liquor. As the pellet loses size it will slowly drop down to the bottom of the beaker and the bubbles will act as an agitation system for the liquor.
At no stage will the pellet cause excess frothing unless the pellets have not been fully dried after formation.
It will take a few minutes for each pellet to be consumed and another added, you can be doing something else and just go to the beaker to add another pellet when convenient.
As you approach the end point the familiar brown precipitate is spread through the solution, add one more pellet and leave overnight.
The solution is filtered in the morning and the solids are boiled gently in 50% HCl until all of the solids have clumped.
Any excess metabisulfite is digested in the HCl.
This system is less time consuming than others but some operators may have other and better methods.
It may be worth while to have a thread dedicated to these types of techniques.
Deano
The first is the hobby person who is interested in the chemistry involved and who wants to produce nice looking high purity product. Generally works with gold quantities up to 100 grams per batch.
The second is the hobby person who is interested in operating a larger process capable of treating material in gold batches of up to about one kilogram. Usually but not necessarily has refining stages in the process stream.
The third is the commercial operator who has pretty much sorted out the cyanide or other lixiviant leaching and recovery of the precious metals in gold batches of kilograms. May do their own refining stages or may sell to a refinery depending on what relationships have been developed with the refinery. Many of these operators are licensed to mark and sell the gold produced in their own right.
Generally the processes used by each level of processing are different but some crossover occurs.
The first level hobby operator is in it with the aim of satisfaction from a process carried out well, the value of the gold is nice but is not necessarily the main driver.
Generally the hobby operator has less commercial type equipment such as commercial type fume hoods and condensers and tends to compensate for this by being very careful with the chemicals and processes being used. These operators also tend to be less familiar with chemistry but still experiment more than the higher level processors.
Most interesting developments chemically and mechanically come from these operators, they deserve all the encouragement they can get.
The second level operators are separated from the first level operators by both the quantities of gold recovered and the processing equipment they use.
This requires a reasonable investment in processing equipment in order to treat the quantities of material involved.
Many of these operators base their process pathways on those of commercial operators, they work on the principle that these must be the the best methods because the big operators use them.
This is not necessarily correct, many of these process pathways are only viable for large scale processing.
It is usually useful for a second level operator to scrutinise their processing in detail to see if what they are doing is actually the best method for their own use.
One of the easiest mistakes to make is to not fully cost their own time and thus look for more efficient processing options to lessen their time.
One step I have previously used is in the precipitation of gold from chloride solutions with sodium metabisulfite.
If the metabisulfite is added as a solution then a greater liquid volume will need to be filtered at the end of the process, this will use up some more time.
If the metabisulfite is added as a powder then time is used when the additions are made under agitation, you don't want to over froth.
If you dampen the metabisulfite and put it into plastic moulds used for lolly making or similar you can get dry pellets about the size of apricots.
If you add these pellets to your beaker of gold chloride one at a time the pellet will start fizzing near the surface of the liquor. As the pellet loses size it will slowly drop down to the bottom of the beaker and the bubbles will act as an agitation system for the liquor.
At no stage will the pellet cause excess frothing unless the pellets have not been fully dried after formation.
It will take a few minutes for each pellet to be consumed and another added, you can be doing something else and just go to the beaker to add another pellet when convenient.
As you approach the end point the familiar brown precipitate is spread through the solution, add one more pellet and leave overnight.
The solution is filtered in the morning and the solids are boiled gently in 50% HCl until all of the solids have clumped.
Any excess metabisulfite is digested in the HCl.
This system is less time consuming than others but some operators may have other and better methods.
It may be worth while to have a thread dedicated to these types of techniques.
Deano
upcyclist
Well-known member
I like that--thanks for the tip, Deano! I'm definitely at your "stage 1", so I might make mine a bit smaller, like a teaspoon or so.Deano said:
I'm tempted to try it with copperas, too
anachronism
Well-known member
- Joined
- May 31, 2016
- Messages
- 3,169
Nah - copperas works best dissolved in hot water and a dash of HCl to clear it up to an emerald green colour.
There is a lack of detailed knowledge around the adsorption of gold on to activated carbon.
My experience covers both lab scale and plant scale adsorption.
All carbons are not created equal despite what the marketing boys say.
The cheapest and easiest to use are coconut shell sourced carbons.
Every so often there will be articles promoting coal sourced carbons, these are great if you really want to have problems downstream of the adsorption stage. Poor filtration and retention of gold on un-ashable solids are the major problems.
As with most processing there are trade-offs.
The adsorption rate of gold cyanide is controlled by the surface area of the carbon per unit volume of liquor.
Thus a fine carbon powder will adsorb the gold complexes faster than coarser particles under the same conditions of liquor tenor and conditioning, liquor volume, agitation and carbon addition rate.
The trade off is that when the carbon is ashed the coarser particles of carbon will ash more completely than the fine powder.
The fine powder will tend to form an insulating blanket layer which prevents oxygen access to the un-ashed powder below this layer.
This then requires either rabbling during the ashing stage or a re-ashing of the un-ashed carbon after digestion of the ash.
So you have a faster adsorption but a slower ashing stage.
If a coarser carbon is used then the ashing stage is usually completed in one pass but the adsorption stage is slower.
For lab work I use 4 x 8 mesh Pica carbon.
I have found that usually the Pica carbon is fairly fresh, it has not sat in a ware-house for months.
Activated carbon will go off just by sitting in a bag in a ware-house.
I buy it in 25kg plastic bags which I open and put into 2 x 20 litre wide mouth black poly ethylene drums with screw on lids.
The first thing I do when using a sample from the drum is to put however much I want to use into a plastic kitchen hand sieve, 2mm openings.
This is vigorously shaken under a strong stream of water from a tap to remove fines and round off weak corners.
If the carbon is fresh there will be a strong fizzing of the carbon in the sieve, the pH will be raised to above 10, usually above 10.5.
You have two options for the adsorption step.
The least appealing is to stir the carbon in a container of the cyanide liquor.
Not only is this slow but you will generate fines which you either discard or slowly filter out for separate processing.
The preferred method is to flow the solution through a column or cartridge containing the carbon.
This method will minimise fines generation.
Fresh carbon in good condition will easily load gold cyanide to around 15 grams of gold per kg of carbon.
If there is extended contact between the carbon and the liquor , loadings of 30 grams per kg can be obtained.
Always keep in mind that some or most of the gold loading sites on the carbon can be taken up by base metals such as copper if high levels of these metals are present, this will lower the gold loading possible.
I ash the dried carbon in an electric muffle at 650C overnight using a ceramic dish.
It is important to have slow heating and cooling cycles so as not to crack the dishes.
In other words, do not open the door wide on a cold morning after ashing and expect the dish to remain in one piece.
The ash is digested in aqua regia and precipitated out as per usual techniques.
The precipitated gold is then given a boil in 50% HCl until all of the gold comes together in large clumps.
It is then dried and smelted.
One of the major points for using carbon for small scale lab processing is that it will adsorb gold in virtually all complexed forms through out the pH range whilst not being affected by the chemical conditions itself.
Deano
My experience covers both lab scale and plant scale adsorption.
All carbons are not created equal despite what the marketing boys say.
The cheapest and easiest to use are coconut shell sourced carbons.
Every so often there will be articles promoting coal sourced carbons, these are great if you really want to have problems downstream of the adsorption stage. Poor filtration and retention of gold on un-ashable solids are the major problems.
As with most processing there are trade-offs.
The adsorption rate of gold cyanide is controlled by the surface area of the carbon per unit volume of liquor.
Thus a fine carbon powder will adsorb the gold complexes faster than coarser particles under the same conditions of liquor tenor and conditioning, liquor volume, agitation and carbon addition rate.
The trade off is that when the carbon is ashed the coarser particles of carbon will ash more completely than the fine powder.
The fine powder will tend to form an insulating blanket layer which prevents oxygen access to the un-ashed powder below this layer.
This then requires either rabbling during the ashing stage or a re-ashing of the un-ashed carbon after digestion of the ash.
So you have a faster adsorption but a slower ashing stage.
If a coarser carbon is used then the ashing stage is usually completed in one pass but the adsorption stage is slower.
For lab work I use 4 x 8 mesh Pica carbon.
I have found that usually the Pica carbon is fairly fresh, it has not sat in a ware-house for months.
Activated carbon will go off just by sitting in a bag in a ware-house.
I buy it in 25kg plastic bags which I open and put into 2 x 20 litre wide mouth black poly ethylene drums with screw on lids.
The first thing I do when using a sample from the drum is to put however much I want to use into a plastic kitchen hand sieve, 2mm openings.
This is vigorously shaken under a strong stream of water from a tap to remove fines and round off weak corners.
If the carbon is fresh there will be a strong fizzing of the carbon in the sieve, the pH will be raised to above 10, usually above 10.5.
You have two options for the adsorption step.
The least appealing is to stir the carbon in a container of the cyanide liquor.
Not only is this slow but you will generate fines which you either discard or slowly filter out for separate processing.
The preferred method is to flow the solution through a column or cartridge containing the carbon.
This method will minimise fines generation.
Fresh carbon in good condition will easily load gold cyanide to around 15 grams of gold per kg of carbon.
If there is extended contact between the carbon and the liquor , loadings of 30 grams per kg can be obtained.
Always keep in mind that some or most of the gold loading sites on the carbon can be taken up by base metals such as copper if high levels of these metals are present, this will lower the gold loading possible.
I ash the dried carbon in an electric muffle at 650C overnight using a ceramic dish.
It is important to have slow heating and cooling cycles so as not to crack the dishes.
In other words, do not open the door wide on a cold morning after ashing and expect the dish to remain in one piece.
The ash is digested in aqua regia and precipitated out as per usual techniques.
The precipitated gold is then given a boil in 50% HCl until all of the gold comes together in large clumps.
It is then dried and smelted.
One of the major points for using carbon for small scale lab processing is that it will adsorb gold in virtually all complexed forms through out the pH range whilst not being affected by the chemical conditions itself.
Deano
I have to make a comment on Deanos threads.
If anyone is interested in gold recovery please take the time to fully read and understand what he is trying to teach, we are sitting at the feet of a master of the trade, some of the processes will not cheaply or easily translate to small scale recoveries but many are scalable and can be done with a little ingenuity and by building equipment oneself but to be able to do so you need to understand what is been taught.
If anyone is interested in gold recovery please take the time to fully read and understand what he is trying to teach, we are sitting at the feet of a master of the trade, some of the processes will not cheaply or easily translate to small scale recoveries but many are scalable and can be done with a little ingenuity and by building equipment oneself but to be able to do so you need to understand what is been taught.
kurtak
Well-known member
nickvc said:
Thanks for the comment Nick --- AND - I 'fully" agree
In fact I have been doing exactly that for the last week or more now - spending like a couple hours a day often with multiple tabs open to other threads Deano has posted to - in order to cross reference between threads & thereby truly gain a FULL grasp of the knowledge & info Deano has been so generous in sharing for our benefit
It has so far been an awesome "study" - which I am clearly not done with yet
That said - I would like to say --- "Thank You Deano" for the time & effort you have put into sharing the vast knowledge of your experience with us :!:
It is much appreciated :!: & truly :G
Kurt
Earlier I mentioned in several posts the adjusting of gold chloride solutions to pH 1.5 before precipitating the gold.
What I did not do was to say that the pH reading was 1.5 by instrumental reading as per any pH meter.
Due to the chloride levels affecting pH meters, the actual pH is around 2.5 even though the reading on the meter is 1.5.
My apologies for the oversight.
Deano
What I did not do was to say that the pH reading was 1.5 by instrumental reading as per any pH meter.
Due to the chloride levels affecting pH meters, the actual pH is around 2.5 even though the reading on the meter is 1.5.
My apologies for the oversight.
Deano
One of the things I forgot to put in the post about vat leaching is that the flow of liquor aimed for is approximately 10 parts liquor to 1 part carbon by weight.
So if you have a ton of carbon in the adsorption column you can pass up to around 10 tons of liquor per hour through the column.
Note that this is not necessarily the optimal amount for a particular vat it does give a good starting approximation for the flow which can be increased or decreased as necessary for maximum adsorption of gold from the liquor on to the carbon.
Deano
So if you have a ton of carbon in the adsorption column you can pass up to around 10 tons of liquor per hour through the column.
Note that this is not necessarily the optimal amount for a particular vat it does give a good starting approximation for the flow which can be increased or decreased as necessary for maximum adsorption of gold from the liquor on to the carbon.
Deano
Most people fail to understand that many chemical reactions are equilibrations or are slow acting or both.
These reactions are not like a switch which is either fully on or fully off.
This means that under the conditions specified the particular reaction may appear not to be happening even though chemically there is a reaction.
Often a change in conditions will drive the equilibrium further to one side of the equation or speed up the reaction or both.
The effect is best compared with evaporating water.
If you apply enough heat to boil the water then the water rapidly and visibly forms water vapour as steam.
Think of a bucket of water which very slowly dries out over a time of several weeks.
Unless the temperature of the water is low enough to freeze the water there will still be a loss of water as vapour, just that it may be so slow that it is not visible.
A similar effect occurs with potassium ferro and ferri cyanides.
The formation of cyanide complexes from the ferro-cyanide occurs under much less severe conditions than the formation from ferri-cyanide.
This means that under the same conditions you will expect to get a greater conversion of the ferro-cyanide complex to cyanide complexes than you would from the ferri-cyanide complex.
This does not mean that under moderate conditions such as sunlight there will only be conversion of the ferro-cyanide complex to cyanide.
There will be some conversion of the ferri-cyanide complex to cyanide but this will be substantially less than the ferro-cyanide conversion.
Keep in mind that you have two effects occurring here.
The first is the degree of conversion is relative to the amount of uv energy being provided.
The second is that both of these conversions are equilibrium reactions and if the uv is removed then the cyanide will revert to ferro and ferri-cyanides.
So you need the uv to start the conversions and you have to keep the uv going in order to maintain the free cyanide.
If you have thin gold plating on an object then the ferri- cyanide will still produce enough cyanide under sunlight to deplate the object, it will just do it very much slower than ferro-cyanide.
The gold cyanide complexes formed will load onto activated carbon exactly the same as gold cyanide complexes formed from sodium cyanide.
Theoretically if you converted ferro or ferri-cyanide complexes to cyanide by applying uv and then leached gold with this solution you should be able to remove the uv source and have any unused cyanide revert back to the ferro or ferri-cyanide form.
In practice the presence of other materials in the leach may lock up some or all of the ferrous or ferric ions and prevent complete reversion, this effect would depend on the feed supplied for leaching and the composition of the water.
The entire process could be run in a darkroom fitted with switchable uv lamps and the cyanide levels tested to see if, with that particular feed, there was complete reversion.
If the reversion was effectively complete then zincing out of the gold should also be complete.
Deano
These reactions are not like a switch which is either fully on or fully off.
This means that under the conditions specified the particular reaction may appear not to be happening even though chemically there is a reaction.
Often a change in conditions will drive the equilibrium further to one side of the equation or speed up the reaction or both.
The effect is best compared with evaporating water.
If you apply enough heat to boil the water then the water rapidly and visibly forms water vapour as steam.
Think of a bucket of water which very slowly dries out over a time of several weeks.
Unless the temperature of the water is low enough to freeze the water there will still be a loss of water as vapour, just that it may be so slow that it is not visible.
A similar effect occurs with potassium ferro and ferri cyanides.
The formation of cyanide complexes from the ferro-cyanide occurs under much less severe conditions than the formation from ferri-cyanide.
This means that under the same conditions you will expect to get a greater conversion of the ferro-cyanide complex to cyanide complexes than you would from the ferri-cyanide complex.
This does not mean that under moderate conditions such as sunlight there will only be conversion of the ferro-cyanide complex to cyanide.
There will be some conversion of the ferri-cyanide complex to cyanide but this will be substantially less than the ferro-cyanide conversion.
Keep in mind that you have two effects occurring here.
The first is the degree of conversion is relative to the amount of uv energy being provided.
The second is that both of these conversions are equilibrium reactions and if the uv is removed then the cyanide will revert to ferro and ferri-cyanides.
So you need the uv to start the conversions and you have to keep the uv going in order to maintain the free cyanide.
If you have thin gold plating on an object then the ferri- cyanide will still produce enough cyanide under sunlight to deplate the object, it will just do it very much slower than ferro-cyanide.
The gold cyanide complexes formed will load onto activated carbon exactly the same as gold cyanide complexes formed from sodium cyanide.
Theoretically if you converted ferro or ferri-cyanide complexes to cyanide by applying uv and then leached gold with this solution you should be able to remove the uv source and have any unused cyanide revert back to the ferro or ferri-cyanide form.
In practice the presence of other materials in the leach may lock up some or all of the ferrous or ferric ions and prevent complete reversion, this effect would depend on the feed supplied for leaching and the composition of the water.
The entire process could be run in a darkroom fitted with switchable uv lamps and the cyanide levels tested to see if, with that particular feed, there was complete reversion.
If the reversion was effectively complete then zincing out of the gold should also be complete.
Deano
kernels
Well-known member
Fantastic post and excellent thread Deano, thanks for taking the time to share your knowledge. I tried some Potassium Ferricyanide this week and it happens pretty much exactly as you describe. Relatively slow process that is very affected by the level of sunlight hitting the solution. When moving the solution to a dark location after the leaching, I had some yellow powder precipitate, so very likely the cyanide reverting back as you suggest.
That is so true, and the minute you realize the truth in the above statement you start to get paranoid as a refiner. Whatever you do there is always some gold dissolved when you don't want it and remaining when you don't want it. A 100% yield is never possible.Deano said:
For more on that topic...
http://www.chemguide.co.uk/physical/equilibria/kc.html
https://en.wikipedia.org/wiki/Equilibrium_constant
https://www.youtube.com/watch?v=cHAjhM3y3ds Equilibrium
https://www.youtube.com/watch?v=xfGlEXWDRZE Equilibrium constant
Göran
To add to Gorans comments I think most members who have or do refine in quantity know that even with karat scrap there is always some values in the system somewhere that cannot be recovered straight away, once you enter the world of e scrap or ore recovery and refining this becomes even more true due to the complex chemistry that occurs or doesn't with the mix of metals and elements, there is also a cost else,emt to add to the mix when it becomes uneconomic to chase any remaining values.
