# Further things which may be of interest to members



## Deano

For the last few weeks I have been processing an electrowin sludge from a gold/copper mine carbon strip plant.

The sludge is formed from the cathodic reduction of metals solubilised from the stainless steel anodes in the electrowin cell.

Major metal components of the sludge are manganese, nickel and chrome.

The sludge contains around 1% gold with some silver, no PGMs.

The sludge also contains high levels of chloride from the strip circuit water.

As an electrowin product, the gold is encapsulated in the manganese and is not in the main accessible for cyanide leaching.

The dry product is a fluffy brown solid which becomes a brown slime, virtually unfilterable, when wet.

As there are a few hundred kilograms of this material the operator was quite interested in recovering the gold values.

If the sludge is placed in acid solution and the mangaqnese dissolved around 70% of the gold also instantly dissolves as a gold chloride.

If placed in a pressure filter I could only filter 5 litres of liquor in 24 hours from this pulp.

The first requirement was to improve the filtration of the pulp.

I did this by furnacing the dry sludge for 8 hours at 500C.

The operating temperature range was 450 to 550C, furnacing outside of this range caused further problems.

The furnaced material was then digested in HCl and enough HNO3 was added to drive the gold dissolution to maximum values.

An attempt was made to put the gold values from this pulp onto carbon however the manganese swamped the carbon and allowed only low gold adsorption.

I then took the pulp and added NaOH to a pH around 10.

All of the metal values including the gold precipitated out.

I then added cyanide to the pulp at 3 times stochiometric for the gold values and solubilised just the gold and silver.

As the gold had precipitated rather than been electrowon it was accessible to a cyanide leach.

The pulp was then pressure filtered and the liquor was contacted with carbon and the gold loaded thereon.

The carbon was later ashed with the ash digested in aqua regia and the gold recovered with metabisulfite.

Note that the gold could not be recovered from the filtered liquor by electrowinning as there was a high chloride level in the liquor which would lead to high chlorine evolution levels.

The gold levels at all stages were monitored by AAS and residues were assayed by aqua regia digest.

It appeared that the initial sludge was an artefact of attack on the stainless steel anodes by the copper rich strip solution.

The treatment process I developed is not the only one possible but is what I felt was the most cost effective.

If the above can help anyone with a similar problem feel free to use it.


Deano


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## g_axelsson

I don't think I'll ever have an opportunity to test this procedure, but it was an interesting read.

Good work!

Göran


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## Deano

I was a bit rushed when I posted the last effort.

What I was wanting to say but missed doing so was that the post was really about the effect in a leach of having manganese in solution.

For most people this involves an acid leach, having manganese in solution in alkaline solution is not an everyday procedure.

Having manganese in solution means that any gold present will not precipitate out until the manganese has been itself reduced to Mn2+.

You can be adding precipitating agents in industrial quantities to achieve this depending on the manganese levels.

The idea of adding a whole lot of reductant such as metabisulfite to the solution usually results in the dropout of brown manganese complexes as well as the gold, there is a small window where only gold is precipitated.

If someone has a solution which has known gold dissolved but which will not drop this gold when the correct amount of precipitant has been added then the presence of manganese in solution should be investigated.

The presence of manganese in solution also interferes with solvent extraction both in the loading cycle and the phase separation stages.

Manganese in solution will usually be sourced from having stainless steel present in the leach step.


Deano


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## Lou

I process considerable gold obtained from stripping of manganese containing materials and have never had an issue in selective gold precipitation or solvent extraction or resin with Mn(II/IV) in acidic/oxidizing milieu. Usually my big demon is Fe(III).


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## Geo

Could the manganese carry over to gold recovered by the copper(II) chloride leach? I know that the nickel substrate is mostly left intact from AP solution. I believe that manganese is used in the Kovar alloy.

Fe	Ni Co C Si Mn
balance	29%	17%	< 0.01%	0.2% 0.3%

Would this be enough to cause problems with the precipitation with SMB?


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## Deano

The material I was working with was the sludge from a carbon strip electrowin cell from a gold circuit.

A quick analysis of the sludge showed around 80% Mn, 10% Ni and 10% Cr with 1% Au.

I have had considerable experience with gold recovery from acid liquors containing low levels of these metals and have had few problems.

It was only when I worked with high tenor leaches of these metals that I had problems, filtration being the worst.

Multi cycle aqua regia leaches still left a problem residue.

I don't know if it was the manganese per se that was the problem or if it was the combination of Mn and the other metals.

I did run a diluted leach and had little problems in recovery from such.

Running more dilute liquors on the total was not an option due to the volume of liquor involved. I really don't want to spend my retirement on this one project.

The main message is to avoid dissolving stainless steel in your gold leaches, low levels are probably OK for small scale processing.

Deano


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## Deano

Table optimising

Many years ago I was involved with a project to maximise gold recovery from a milled ore by using a small (3') wilfley table to clean up concentrates and middlings from a bank of larger tables.

Due to environmental constraints chemical processing was not allowed on site.

A short period of operation revealed that running the table in its standard format was leading to losses of fine gold where the particle size was less than 50 micron screen size.

The gold particles were flattened from the milling and appreciable quantities were also lost from the plus 100 micron screen sizes.

Despite doing all of the standard optimisings of adjusting feed size screening, feed rate onto the table, table side tilt and water flow I could not get a major recovery improvement.

I went into the literature and found that the largest number and most useful papers were those from the British tin industry.

Basically they said that keeping a tight sizing on the feed was vital, I was already doing that so OK there.

Making sure that the table was level on the longitudinal axis was a fundamental which I was also doing.

Feed rate was best when a loose bed was set up along the table, don't put too little feed on the table nor try to put too much feed on the table.

Keeping the feed rate constant was also very important, I was feeding from a wet sump with a screw feeder so OK there also.

Side tilt was to be such that clay fraction particles were washed over the side of the table but the tilt was to be little enough that a middlings product could be readily separated at the end of the table.

Even when I had the table set up to cover all of the above parameters I was still losing fine gold.

The only area where I did not have full control was the side water coming onto the table, no matter how much I tried I could not get a perfectly even flow across the table.

I decided that I needed to improve the delivery of water to the table.

I did so by running a length of 3/4" copper pipe suspended about 4" above the side of the table where the water exited the original flow boxes.

The pipe was blanked off at the table exit end and was connected to a hose and ball valve at the feed supply end.

Ever 1" along the pipe was drilled a 1/8" hole.

When the water was turned on a curtain of water sprayed down onto the top edge of the table and delivered an even flow of water which could be easily adjusted for flow rate with the ball valve.

Even this change only partly improved the recovery, it was evident that there needed to be a difference in the water flow rate supplied to various parts of the table.

After a lot of testwork I settled on having the water delivery pipe as two pipes.

The holes remained the same size and spacing but the delivery was split into two parts.

The first part was as above but only extended 2/3 of the table.

The second part covered the last 1/3 of the table, each part was as a separate length of pipe so that adjustments could be made to either part without affecting the other part's flow rate.

In order to keep the water holes at the 1" spacing the pipe for the last 1/3 of the table was fed from the bottom of the table and the pipe ends almost touched.

Each pipe length had its own ball valve for separate flow adjustment.

The side wash water pipes were fed from an overflow overhead tank so that a constant head was maintained.

This was important on a mine site where valves were being opened and closed in other parts of the circuit, this would affect the pressure to the wash pipes.

This setup allowed recovery of free gold down to 25 microns, the disappointing part was the low weight of the 25 to 50 micron gold recovered, it looked a lot as a sheet like paint on the table but weighed far less.

On the plus side there was a substantial improvement in the plus 50 micron gold which did weigh well.

If run from a municipal water supply the overhead header tank may not be necessary depending on the vagaries of the particular supply.


Deano


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## g_axelsson

This thread just gets better and better, one nugget of information at a time.

Thanks for sharing!

Göran


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## Finn from Ecuador

Thank you Deano

Your exellent article about fine tuning Wilfley table would be very good to have in mining and ore section too.

Salud

Finn


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## kazamir

Hi Deano
Do you see potential gold recovery from electronic scrap with a Wilfley table.
Regards
Kazamir


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## Deano

I started off these threads in the chemical processes section because they originally were mainly chemically focused.

I am keeping them there as I expect that the bulk of material to come will also be chemically focused.

If someone wants to post these more mechanically mining parts in the actual mining section I am happy to have them do so.

I can see potential for scrap treatment with a table but only in the separation of metallics from non-metallics.

This would require a consistent milling format for the material, with the variations of plastic types and thicknesses I see this as the hardest part to get right.

If the material can be milled so that the metallics are freed from the plastics then I would expect a very good separation to be achieved.

Note that some awfully cheap tables are available from China, the drives are really good but the actual table surface is pretty ordinary and should be replaced in total.

Deano


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## Finn from Ecuador

Dear Deano and Moderators

May i add a link to your Wilfley table article to another gold prospecting forum (GPEX) as well?

Salud

Finn


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## FrugalRefiner

It's OK with me.

Since Deano's threads tend to cover a lot of subjects, you might want to go ahead and start a new thread.

Dave


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## FrugalRefiner

My mistake.  

When I read Finn's post, I thought he was intending to add a link to something on the forum he mentioned that might have additional information. Now that I've re read it, I believe he's asking if he can link to this article from the other forum. Finn, I'm sorry for any misunderstanding, but I don't have the authority to answer your question.  

Dave


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## g_axelsson

I can't see any problems with linking from outside sources. I link heavily from my wiki and gogle links to everything.

Linking between sites is what makes the web what it is today.

Göran


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## Deano

As far as I am concerned the decision is not mine to make, I have no objection to the proposal but once I have posted information it is not my position to say what links may be put in place. Refer the point to the site ownership.

Deano


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## goldenchild

Few hundred kilos of the material with 1% AU? So there were several pounds of AU precipitated?


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## Geo

600 avoirdupois pounds at 1% is 7.29 troy pounds.


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## Deano

Sometimes when precipitating gold the product is fairly slimy and difficult to filter.

One way to improve the gold floc size is to put some silica flour in the precipitating liquor.

About a teaspoon in a litre of the precipitation liquor is a good working amount.

Stir it to distribute it evenly.

The silica will report to the slag when the gold precipitates are smelted.

Deano


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## kazamir

Hi Deano
Would you care to make some comments regarding the M44 Process.

http://goldrefiningforum.com/phpBB3/viewtopic.php?f=37&t=23030#p241735


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## Deano

I work in industrial scale with precious metal recovery.

Unless I have exceptional circumstances I will use cyanide as my solvent of choice for a whole range of reasons.

If I cannot use cyanide then my first response is; can I sell the material to someone else who is set up to handle this type of material.

My time I regard as too valuable to engage in small scale projects for a few grams of metal.

Having said that I do not want in any way to denigrate the efforts of people who do run small scale projects.

I am impressed by the perseverance, ingenuity and curiosity shown by these people.

Most of them have much greater familiarity with the techniques used in their projects than I will ever have.

I can readily understand what and why they do some things but virtually none of these things are applicable to my processing needs.

The M44 process fits under this area, I see it as another variation on a well worn theme.

It is, however, a variation which will be exactly what someone needs for a certain project and so should not be put aside on general principles.

The proof of any process is whether that process is the most suitable for a certain need.

This decision can only be made by people who have used the process and can pass practical judgement.

From my perspective the process is too involved to be useful to me but that does not mean that the process is not exactly what someone else needs.

I run this thread because I have some knowledge which may be of use to some of the members.

This knowledge should not be hoarded for just myself but should be put out for any one to use if it will help them.

This does not make me any type of guru in the precious metal processing field, there are many members who have greater experience and skills than I have.

What I am trying to do is to reveal some of the tricks of the trade which many members possibly are familiar with but do not recognise as being more than just useful to themselves.

Deano


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## Deano

Operating a jig

There are probably as many myths and sources of poor information about using jigs for separation as there are for all other methods combined.

Pretty well all jigs are of twin hutch design with hutch sizes from 150mm or 6" square for final clean up to 1.1 metres or 42" square for a full scale primary jig.

Jigs come in two main types, the end pulse and the undertyre pulse.

The end pulse is the cheapest and easiest to operate.

Some operators dislike end pulse jigs as they have dead zones in the corners where the pulsation is less pronounced.

Realistically the dead zones make little difference to the efficiency of the jig.

The undertyre jigs of which the Pan American is the best known example do not have dead zones and so are preferred for secondary or clean up operation.

A primary jig will handle one tonne per hour of minus 10mm (3/8") sized feed per hour per 0.1 sq metres (1 ft sq) of surface area.

Just enough water is used to wash the feed onto the jig such that the feed will move easily along the jig.

Many operators have a flood of water running over the jig and then complain about losing fine gold.

The settings for the jigs are simple.

A primary jig has a 25mm (1") stroke length at 100 rpm.

A secondary jig has a 12mm (1/2") stroke length at 200 rpm.

A tertiary jig has a 6mm (1/4") stroke length at 400 rpm.

Hutch water is fed in at a rate so that if your hand is placed flat on the ragging there is a strong suction pulling your hand down through the bed.

Generally a single hutch of a secondary jig will handle the concentrates from two twin hutch primary jigs.

The area where most of the myths originate is the screens and ragging.

Most operators use wedge wire or woven mesh screens with what is called ironstone ragging, this is the heavier fraction of the natural feed.

The first problem with this screen/ragging combo is that the screens will blind with ragging and require frequent cleaning.

The second problem is that the ragging bed cannot close fully on the downstroke and so a lot of lighter material will pass through it and report to the cons.

Operators have used many weird systems to overcome the above problems, lead shot and ball bearings among them.

None of them are really successful as they all still lead to a bed which cannot close fully.

The problem can be overcome by using punched stainless steel plate for the screens and using the punchings themselves as the ragging.

When the punching is done the punchings go oversize and will not fit back in the punched holes.

Hole sizes are around 5mm (3/16") for the primary jigs, 3mm (1/8") for the secondary and tertiary jigs.

It takes a 20 litre steel bucket full of punchings to fill a twin hutch primary jig.

It means that you will need to get your screens and punchings from the same source of manufacture.

Many operators have a small section of riffles in the tray leading from the wash down tray to the jig.

These are used to get the coarse gold out of the system so that there is not the need to clean out the ragging on the jigs so often.

It also simplifies security in that you only need to put a lockable mesh over the riffles rather than over the full jig beds.


Deano


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## Deano

Chloride leaching

Further to previous posts on chloride leaching I have developed some useful information on oxidation potentials.

The most efficient form of chlorine for leaching is hypochlorous acid, it will supply an ORP or Eh (think oxidation strength, these both mean the same thing) in excess of 800 mv.

This is high enough to leach gold but not high enough to keep the gold completely in solution.

The leached gold may be kept in solution by adding common table salt to the solution.

The level of gold which can be kept in solution increases with increasing salt concentration up to about 20% salt.

Above 20% salt the effect of increasing the concentration diminishes.

As salt can only be dissolved in room temperature water to 35% the use of 20% salt solution allows a fair margin for evaporation etc.

So now you have the 20% salt solution ready and you are ready to set up the hypochlorous acid level and pH.

The important point to remember with pH is that the meter reading is affected by the presence of high salt levels.

In the area of pH 7 a salt addition of 10% will lower the pH reading by about 1 unit, a 20% salt solution will lower the reading by about 1.1 units.

This means that if you are aiming for a solution with 20% salt at pH 7 you must add acid or alkali to adjust the instrument reading to pH 5.9.

The ratio of hypochlorous acid to hypochlorite is pH controlled, each pH unit change gives about a 10 times change in the ratio.

At pH 6 the ratio is around 25

At pH 7 the ratio is around 2.5

At pH 8 the ratio is around 0.25

At pH 9 the ratio is around 0.03

The ratio has an increase below pH 6 but there is a fair quantity of free chlorine starting to be evolved at these lower pHs, the quantity gets higher as the pH lowers.

Note that what is generally referred to as chlorine leaching starts at pH 5 and is more efficient down to pH 3, below this pH the free chlorine has a strong preference to evolve as a free gas and not stay in the leach liquor unless under pressure.

All the above says that if you want to carry out chloride leaching you have the best conditions with a 20% salt solution (do not use iodised salt) at actual pH 7 or a meter reading of pH around 6.

The obvious question is why not run the leach at actual pH 6 where more hypochlorous acid is available.

At actual pH 6 you are starting to get noticeable evolution of free chlorine, you can smell it.

At actual pH 7 there is virtually no evolution of free chlorine but there is still enough energy in the leach to dissolve and maintain in solution high levels of gold.

This gives you two advantages, from a safety aspect there is virtually no toxic free chlorine, from an economic standpoint you are using virtually all of the available chlorine in your processing and not wasting it as free off gassing chlorine.

So the lower performance of the actual pH 7 leach is not actually much lower than an actual pH6 leach but has advantages.


Deano


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## Deano

When going over an old notebook of things which interested me even though I could see little prospect of using them I found the following.

I have never tried the process and have absolutely no further information than what appears here.

An old jeweller said that he stripped gold plating from metallic articles by first making a leach solution of 1 litre of sulfuric acid with 1 ml of glycerine stirred through it. No information as to the sulfuric acid strength.

This leach solution was put in a ceramic dish with a piece of lead partially submerged in the solution to act as a cathode.

The piece of metal you want to strip is placed in the solution as the anode.

A battery charger is used to deliver power to the circuit, he used a 10 amp unit which could be run at 6 or 12 volts.

The connectors were stainless steel spring clips, he stressed the use of stainless steel.

The gold formed a black powder on the floor of the dish, he claimed that it was fast and had no fumes.


Deano


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## FrugalRefiner

The glycerin is not necessary.

Dave


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## g_axelsson

That is the well known procedure of "reverse plating" or "sulfuric cell", you will find a lot of information and even videos if you do a search with these terms.

The sulfuric acid has to be at least 90% strength and the temperature must be kept down or base metals will be attacked.

Göran


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## kurtak

Deano

Here is "one" of the best threads on the forum about building just such a gold stripping sulfuric cell

:arrow: http://goldrefiningforum.com/phpBB3/viewtopic.php?f=40&t=22603

Kurt


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## upcyclist

Me, I'll stay away from the organic compounds as much as I'm able. I can't help but think 

"Hey, you put your nitric acid in my glycerin!"
"Hey, you put your glyc--- BOOM


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## Deano

When I previously wrote about how to use a flocculation agent I had assumed that the processing was being performed at acid pH ranges.

In order to cover the rest of possible processing methods there are the areas of neutral and alkaline pH ranges.

For most ore types these ranges are best processed by using a cationic floc agent.

This is made up and used as per the previous instructions.

In the US I would recommend approaching Tramfloc, they are quite liberal with samples.

They also have a range of cationic floc agents some of which may be more effective on particular ores, generally I find that their standard floc is the most efficient.

If you are wanting to dewater these flocs you can either use a fine mesh vibrating or rotary screen to get the last of the free water out after concentrating the flocs in a tangential cone settler.

The flocs can be made more mechanically stable by the addition of alcohol to the concentrate from the settler, depending on the pH and ore type used, about 1% alcohol is a good starting level.

Deano


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## Lou

Tramfloc is what we use. They are a good company. 

Anionic flocculants--basic conditions
Cationic flocculants--acidic conditions
Neutral floccualnts--metal fines

sulfonated flocculants, very acidic conditions


Best advice it to call their technical support. I've only run into a few things that can't be handled with their charge neutralization/bridge approach. 

@ Deano, we're using the flocs going into a filter press or 36" Buchners.

Lou


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## Deano

Lou

Thank you for picking up my mistyping, I meant anionic but typed cationic.

For strongly acid conditions such as 1% sulfuric acid or greater I have always used Cytec N300 as the floc agent because of its availability in Australia.

Apart from its good performance the N300 has the rating of being food quality, this removes all issues for disposal.

I have no doubts that Tramfloc make similar products of similar high quality.

I pretty well do all of my volume filtration in a 20 litre pressure filter, they were put out by AMDEL in the 80s and 90s as filters for metallurgical labs but can be used for almost any separation.

I lost my enthusiasm for large buchners many years ago, about the time I first saw one of the AMDEL filters.

If I am wanting to carry out very large volume separation I go to using a cone settler with a tangential inlet and clear water overflow with continuous solids drain from the inverted cone apex.

It is really nice to put your pulp into the mixing tank, check the liquid levels and then press a button to start a continuous process run.

You have to be processing a lot of large volumes to go from the 20 litre filter to one of these setups.

Deano


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## Deano

Vat leaching

The simplest and certainly the cheapest method of leach processing gold ores is by vat leaching.

There are many opinions on how to design and operate vats, most of these operate satisfactorily for small scale operators.

The usual flaw in most of these vat designs is that the ore is not fully wetted and thus some of the ore is not leached. Any ore zones which are not fully wetted at the start of the leaching process will not miraculously become wetted during the leaching cycle.

The following is the only method I know of how to build and operate a vat and recover all of the cyanide soluble gold.

Note that the dimensions and weights given are directly scaleable for smaller vats, it is usually only the pipe diameters and the pump sump dimensions which are retained where possible.

The vats I have run are a maximum of 20,000 tons as a single cell.

Vats for small scale mining are usually configured around a 5,000 ton maximum ore parcel per cell.

A large, say 50,000 ton vat will be subdivided into 10 x 5,000 ton cells.

The reason for this is twofold, it is easier to control the process requirements in a cell of that size and it also simplifies the material handling aspects.

Most leaching vats are around 100 metres long by 50 metres wide and divided into 6 cells

Generally the cells are 25 metres wide by 2.5 to 3.5 metres high by 30 metres long. Fall of the length of the base is dependent on the ores being treated, a coarse free wetting ore may be in a cell with fall of 1 cm ( 1/2") per metre, a less permeable ore can have a fall of half this but the 1:100 fall is pretty standard.

There is a fall across the cell, generally around 1 in 25 so a 25 metre wide cell will have a 1 metre fall across it.

These sizes have proven optimal both in leaching performance and in materials handling.

The vats are usually dozer formed with a bobcat finish.

A pumping sump is dug at the lowest corner of the cell, it is usually configured as a 1metre cubed hole .

An access hole is maintained to the pump sump, usually a series of cyanide drums welded end to end for short lived projects or steel pipe 500mm (1.5') or greater diameter for extended use, 200 litre (55 gallon) drums are also often used. This pipe must be bedded on a series of large flat stones in the pump sump to avoid the pipe rupturing the liner. The top of the pipe projects at least to above the wall height of the cell.

The cells are plastic film lined, usually 1.5mm HDPE which is heat welded. Glueing is viewed with suspicion due to failures attributed to salts in the process water.

A cheaper alternative is to use 0.2mm PVC film as a one piece installation.

This does require a lot more bed preparation to flatten out all spiky bits of rock with a bobcat or loader with a lot of weight in the bucket and constant wetting of the working surface.

Overall it is the cheapest option even allowing for the extended bed preparation and if the bed preparation has been done thoroughly it will work as well as the heavier films.

Many of these PVC film vats have operated without any problems for several years.

These vats are usually closed down due to a gypsum build up in a layer just above the graded gravel bed rather than materials failure.

The type of liner used will most likely be mandated by local mines department regulations, the trend has been to the thicker liners.

Drainage inside the vat is provided by 50mm (2") slotted agricultural drain hose which is teed at 1.5 metre intervals off a 75mm (3") PVC pipe.

The 75 mm pipe runs from the pumping sump along the length of the cell at the base of the wall with the greatest depth.

The end of this 75mm pipe furtherest from the sump has an end cap glued on to it.

An elbow is fitted to the 75 mm pipe at the inside of the sump and the pipe then runs up the access hole to a tee piece at the top of the sump.

The horizontal leg of the tee leads to a diaphragm pump, usually diesel powered, the same units as are used on construction sites for dewatering.

This pump will deliver the pregnant cyanide solution to the carbon column, flow rates for the 30 x 25 metre cells are around 10 metres cubed per hour.

The vertical leg of the tee piece leads to a header tank which is at least 6 metres (20') above the level of the top of the vat.

A constant head arrangement is achieved by running the header tank in overflow mode from a separate water source.

A ball valve is installed adjacent to the tee piece on each of these two lines.

The 50mm ag pipes run across the bottom of the cell to the base of the side wall, note that these ag pipes are running uphill on the steep sloping base.

At the base of the side wall the ends of the 50mm ag pipes have a piece of 10 to 12mm poly pipe taped into them.

These lengths of 10mm poly pipe run up the wall to about 1 metre above the wall, ball valves are fitted to the upper ends of these pieces of pipe.

Filter fabric socks are wrapped or slid on all of the ag pipe lengths.

The entire floor is now covered with sized clean creek bed gravel, the usual sizing is 0.5mm (25 mesh) to 5 to 6 mm (1/4").

The gravel bed depth is at least 300 mm (1') and will cover the ag pipe by 250mm (10").

This gravel bed is crucial to the operation of the vat, if the gravel cannot be gotten in the required quantity then the vat will not operate properly.

Before the gravel is put into place a length of 50mm ag pipe is coiled inside the pump sump and connected to a length of 50 or 75 mm PVC pipe which rises up the access hole to tee into the pumping line just after the ball valve on the 75mm line adjacent to the tee on that line. This line is referred to as a scavenging line.

This pipe also has a ball valve fitted so that it can be isolated from the main pumping line.

This scavenging line is used only when a cell is being drained and recovers the liquor which does not report to the main pumping circuit.

The ag pipe part of this scavenge line is also fitted wjth a filter fabric sock and has the sump end capped.

None of the cells have drain pipes installed through the walls to empty out the ore after leaching. These drain pipes always leak solution along the outside of the pipe, despite a lot of trying no one has come up with a leak proof version.

The ore is loaded into the vat as a mixed material. It is vital that the charge is well blended and machine mixed. This ensures that the clay and other fines are evenly distributed and that zones of less permeable material are avoided.

You should never feed pulp directly into a vat from a mill, doing so will almost guarantee leaching problems due to the formation of fines zones.

The usual practice is to place the mill discharge into a dam and when the dam is full and the water has been siphoned off to use an excavator to mix the material well before placing it in the vat.

A large (40 ton+) excavator is used to place the ore into the vat, the boom length is such that the vat can be filled without having machinery drive over the surface of the placed ore

The vat is filled to about half a metre from the top of the walls.

When the vat has been filled the ore is covered with a cyanide solution which is allowed to slowly saturate the ore.

When the leach solution is present in the pump suction line the process of hydraulic lifting can start.

This consists of backflowing water from the header tank through the ag pipe drains to totally saturate the ore charge with water.

By selectively closing the valves on the pipework the water is limited to only flowing backwards through the ag drains.

At this stage the valves on the ends of the 10mm pipes on the ag pipes are all open.

The aim of the procedure is to have the water introduced as a slow moving front which saturates all of the ore.

It will take two or more days to totally fill the vat with water, the process is controlled by the constant head from the header tank.

The process will not work if a pressure pumping filling is used.

What you are trying to do is to remove all air from the ore solids in the vat and replace it with water.

This does not occur in the initial filling of the vat with cyanide solution, it can only be achieved by the hydraulic lifting process.

As the lifting water gradually moves through the cell, water will sequentially begin to come out of the 10mm pipes on the top of the walls.

These pipes act as air bleeds for the ag pipes.

If one of these pipes has no flow or a lesser flow than the other pipes then the valves on the adjacent pipes are closed so that the water flow is increased through the ag pipe and air bleed pipe in question.

When the bleed pipes show similar flows with no air bubbling they can be closed off at the valves as the water front has then passed them and there is no reason to have water discharging from them.

When all of the bleed valves are closed off then the ore parcel is totally saturated.

The lifting water is allowed to continue to be flowed into the vat until there is a layer of water about 300mm (1') deep over the surface of the vat. 

The leaching stage can now proceed.

The valves on the pipework are adjusted so that the pump sucks only from the 75mm PVC pipe connected to the ag drain pipes.

The pregnant liquor is passed through 3 x 1 ton carbon columns in series.

These columns are often made from what materials are at hand but generally around a metre in diameter x 4 metres high

The reason for the three columns is that the first column does most of the gold adsorption, the second column acts as a scavenger and the third column is used when the first column is being changed over after loading so that there are always two columns in series doing the adsorption.

More importantly the third column is there to handle the high gold tenors in the leach liquor when a cell is either first started or restarted after resting.

These gold tenors can reach over 5 grams per cubic metre and will overwhelm a two column setup.

The barren liquor exiting the carbon circuit is returned to the rest of the liquor covering the ore, it is usual to run spray bars of some type to oxygenate the return liquor.

If the pump operates such that it can pump more liquor than the pump feed line can deliver, the pump product will include some air which has been sucked in through the ag lines.

This air will also help re-oxygenate the liquor but the carbon adsorption columns must be designed to accomodate this air, you do not want carbon floating out the columns.

Solution tenors are checked daily and the vat is run until the solution tenor going into the carbon circuit is the same as the exit tenor. 

If the tenor is above 0.5 grams per cubic metre at equilibration this usually indicates that the carbon is loaded to the level where it requires stripping and this carbon is sent to the strip circuit.

It also is possible that the carbon has reached its equilibrium loading level for that tenor liquor and that the gold content of the cell is so low that the gold tenor will not increase without some other actions being taken.

This is generally around 0.2 grams per cubic metre and may take several months of leaching to achieve.

In this case the return water from the carbon circuit is sent to another cell and the first cell is allowed to completely drain, this is where the scavenge line comes into play.

After a couple of weeks drying out the cell has all of the ore dug and mixed by an excavator, this is referred to as fluffing up the ore.

The aim is to break up zones of fines which will have formed during the leach cycle.

The fluffed ore is allowed to aerate for at least a month before it is levelled out and the entire cycle is repeated.

This repetition of cycles is continued until it is no longer economical to do so.

It is not surprising to have initial liquor tenors of 2 to 3 grams per cubic metre on a restart for the first day or so.

When the vat is finally regarded as exhausted it is drained and after a couple of weeks drying out the ore is removed, replaced with fresh ore and the cycle started again.

Note that the exhausted ore is usually removed by trucking it to a tailings heap, vehicles can usually drive safely on the tailings after a couple of weeks drying.

Deano


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## Deano

Vats with liquid on top are often a magnet for aquatic birds, this is not a good thing for cyanide liquors.

A triple benefit method for keeping birds from the liquor is to install high density shade cloth as a layer over the liquor.

The shade cloth is supported on lines of 200 litre (55 gallon) plastic drums which are about 2/3 filled with water and strung in lines with poly rope.

Enough tension is applied to the shade cloth so that it is clear of the liquor and thus presents a dry surface which is not wicking up liquid.

The shade cloth will also substantially cut down on evaporative water losses, this means that the cells can be run deeper into summer than would normally be the case.

The shade cloth also cuts down the degrading effect of UV rays from sunlight on the cyanide complexes, this lessens the amount of make-up cyanide required.

Deano


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## g_axelsson

Thanks Deano, now I can build a leach plant.

... only thing I need now is a couple of thousands of tons of ore. :mrgreen: 

Göran


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## mahalas

Hi everybody .
It is my first post on this great forum. I read the forum for awhile, and I learned a lot. Many thanks .
Hope that one day I will contribute something to that mass of knowledge.
For now I have some questions for Deano.
On earlier post you mentioned the phenomenon of overmilling of sulfides. I try to do some cyanide leaching of sulfides concentrate from tailings. What is industry standard milling for sulfides? I plan to do test on -200 and -325 mesh.

Also could you give some more information about using undissolved starch instead of activated carbon.

Any answer will be greatly appreciated. Adam.


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## Deano

One of the strange things in metallurgy is that gold values which can be cyanided in some head ores cannot be cyanided in concentrates.

It is standard practice to carry out a combination of bottle roll cyanide and aqua regia digest tests on both head ores and sulfide cons and tails.

This tells you what gold recovery is possible with cyanide and what gold distribution is present in your ore.

There are many ores where sulfides are present but the gold in the ore is carried in the quartz fraction and the sulfides are either barren or have low grades. The above test regime will show you where the leachable gold is.

Assuming that there are viable cyanide grades in the sulfides you are now looking at what milling size is required for the best recovery of leachable gold.

This is not necessarily the same size which gives you the best sulfide recovery from the ore.

Assuming that you are working with a tailing which was previously treated for sulfide recovery, the values in the tailings will be fairly low.

This means that the tailings most likely will not be viable to remill if required to improve the sulfide recovery so you are stuck with what you can recover from the as is tailings.

However if the tailings were generated before flotation became established there may be enough sulfides present in coarse ore particles to justify a regrind.

This is because the pre flotation ores were milled as coarsely as possible consistent with reasonable recovery, this gave the best gravity recovery.

However, let us look at what you have in the sulfides available.

The first test is a few aqua regia digests to establish what gold grade is present in the sulfides.

Next you mill representative samples of the sulfides for bottle roll testing.

Not knowing what size your sulfides are, I will assume that they are fairly coarse as in gravity cons.

In this case milling samples to 100% minus 100 mesh (150 micron), 100% minus 200 mesh (75 micron), 100% minus 300 mesh (50 micron) and 100% minus 600 mesh (25 micron)would show how cyanide recovery is likely to be.

The bottle rolls should be run at a solids to liquid ratio of 1:3 with 2 grams per litre sodium cyanide for 24 hours.

You are just trying to get the best recovery regardless of processing cost at this stage, you can always optimise the reagents and ratios later if warranted.

There are good reasons why industry uses a certain process path, usually tied up with it works repeatedly without needing specialist input, ie robust and is cheap for the process.

This is why activated carbon is the gold industry standard recovery method.

There have been many reports of groups, usually university based, who have developed novel gold adsorbents.

None of these have displaced activated carbon as an adsorbent in industry.

Many of these novel adsorbents show great promise in the laboratory but have some characteristic which means that they are either not suitable in an industrial process or are more expensive than carbon.

Deano

There is no standard milling for sulfides, each ore is different and will respond differently to different millings


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## mahalas

Hi Deano.
Thank you for your replay. My tailings are pre flotation ones. Coarse gold was recovered by mercury and the rest of it is encapsulated in sulfide matrix. Tailings are pretty rich from 9 to 16 ppm depending on the mine . Have access to a few.

Unfortunately I do not have a sophisticated lab, but do fire assay on daily bases. Tailings concentrate done on M - 7 table ( table made by Action Mining ), 100% sulfides - mostly pyrite and galena, showing from 180 to 450 ppm depending on the mine, but also on the adjustment of the table.

Test for cyanide ( 300 g concentrate in 1 l solution pH 11, 1 g per liter NaCN, 24 hours in beaker on magnetic stirrer)
showing me 60 % recovery, but my concentrate is coarse - 80 mesh (177 micron).

Bigger amount of NaCN up to 3 g per liter I usually use when I have higher content of silver. In that case content of silver is low ( gold after cupellation has slightly green color ).

Could you comment on it , suggest improvements ? Cheers Adam.


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## Deano

The grade of the tailings is high enough to look at milling and leaching the entire tailings, this would be my first testing program.

The only reason I would not do this is if I could not get cyanide permits.

If I could not get cyanide permits there is a high enough grade in the tailings to stand trucking them to some-where you can use cyanide.

Assuming that you either can get the permits or can truck to a permitted location, you need to find out what recovery can be gotten from the tailings with cyanide.

Carry out the milling sizing tests on the tailings at the sizes previously given for the sulfides only tests.

Use riffle splits of a single sample which is reasonably representative of the tailings, spend time with an auger etc to get good samples.

Filter and dry the milled products, riffle split the dry material into quarters.

First do an aqua regia digest on several splits of one of the quarter splits for each size fraction, 25 gram samples are fine.

You will probably have a fair bit of bounce in the assay numbers from the coarser samples, plan on doing at least 6 samples to get results that are meaningful.

Then do cyanide leach tests on a second split so that you can see what percentage of leach available gold is cyanide leachable for each milling size.

Then do fire assays on the third split to see what the total gold is showing in that milling size fraction.

The fourth split is retained in case repeats are needed.

You are probably thinking something along the lines of " this is expensive to do, is it really necessary."

If you have only small tailings quantities, less than a hundred tons total, then it is not absolutely necessary as you will be approaching the treatment as a hobby project, albeit one with hundreds of thousands of dollars of gold at stake.

You can table them and accept the losses which will occur even though you will be having to have a small leach plant to cyanide the table cons.

Quite simply, I would not carry out gravity separation on a sulfide ore unless the ore was of so low a grade that it was not viable to either process as a whole or to cart to a site where it could be treated.

If you are going to table the tailings you must factor in the cost of picking up the tailings, pulping and sizing them, tabling them and then depositing them in a suitable dam.

If done as a hobby project you will still face reasonable costs to carry out these actions.

You still will not know if milling the tailings will give you a better gravity recovery of the sulfides unless you do the testwork to find out where the gold is and how much of the total gold is in liberated sulfides.

In other words, just do the tests, get the information you need and do a better job.

If the tests show that there is an optimal milling size for the tailings for cyanide recovery and that this milling size also demonstrates good sulfide liberation then you should give consideration to a straight milling and vat leach of the entire tailings.

It will certainly be the simplest and cheapest method of treatment.

A vat leach will give a recovery at least equal to an agitated leach test.

The proportions given for your cyanide leaches are OK.

Deano


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## mahalas

Deano
Cyanide should not be a problem here. Now I am in Columbia on small mine site and they expect to get all permits shortly. This year is the last one, when Colombian miners can use mercury without jail penalty.

I have 1000 liter stainless vat with mechanical stirrer, but before I fill it up I want the process to work perfectly on a bench scale. To recover gold from solution so far I do zincing but plan to switch to carbon. Could you give me a hint how to strip gold from activated carbon without ashing ?

Yes it looks like a hobby project but there is room for something bigger.
Cheers Adam


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## Deano

I am probably not explaining this clearly enough so I will give it a last try.

Fire assay will give you total gold, it will not tell you where the gold is in the ore or what percentage of this gold is available for leaching.

If you do only fire assays then you have no idea what is possible in treating your ore.

It is usually hard enough to viably mine gold without trying to do so without the information you can get by carrying out the basic tests.

Sulfide losses from gravity recovery are usually high due to the sliming of the sulfides during the milling process.

If you carry out gravity recovery of sulfides you have only two choices.

The first is to run the tailings in total over a table and accept that that your losses in the fine sulfides will be very large.

The second is to size the tailings and run the coarse and fine fractions separately over the table.

This is much more efficient but the tabling of the fines is a very slow process and still has large losses.

Assuming that you can find a milling size which gives good cyanide recoveries and you will still have to find this milling size out whether you treat the tailings as a whole or just the sulfides themselves, the most efficient way to treat the tailings is by vat leaching.

I have run vats sized from 50 to 20,000 tons, size is no problem.

I realise that you already have a small leach tank for sulfide leaching and thus wish to use this in your processing but you should look at whether this is the best way to treat the tailings.

If you are happy to have needless losses then by all means go for the tabling option.

Small scale gold recovery from cyanide liquor is easiest done by zincing, keep the gold liquor tenor up and the losses are minimal.

If you recirculate the zinc stripped liquor then the gold losses should be near zero.

Basically carbon is stripped with a 1% caustic/1% cyanide solution at near boiling temperature after the carbon has been hydrochloric acid soaked.

The major controller is temperature.

The strip solution is recirculated through the stripping column and in series with an electrowin cell.

The EW cell has stainless steel perforated anode plates and wound steel wool cathodes.

Depending on the cleanliness of the strip liquor and the whims of the operator the voltage applied is anywhere from 3 to 6 volts.

The process may be sped up by addition of hydrazine hydrate or similar to the strip liquor to get the ORP down below 200 mv.

Depending on the cleanliness of the carbon regarding base metals, strip times are usually from 12 to 24 hours.

If you want to carry out your own carbon stripping you need to see how someone else does it, there are a lot of little tricks for small scale operators to learn.

Deano


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## mahalas

Deano
Thanks for the answer. I know that fire assay gives me the total amount of gold, but with table combination it gave me also gold distribution. Because of that I collect slime for separate leaching ( close to 30 % of total gold ). 

The poorest part are middlings , mostly quartz 1 - 1.5 ppm Au. Tabling is fine but is rather a slow process, especially on 
M - 7. Next week I am going to get a small continuous ball mill. Plan to use it like tromel to separate slime, regrind the rest and -following your advice - leach everything . 

Looks like carbon striping would require additional investment in equipment. I think I should stick to zincing.
What about pretreatment in a leaching tank ? Something like aeration 24 hours or addition of soluble salt of lead ?

Cheers Adam


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## Deano

Fire assay gives you gold distribution but it does not give you leach available gold distribution.

You can only get this with aqua regia / cyanide leaches.

If you have a ball mill and are planning to leach the total feed you only want to mill the tailings to the size which size testing says will get you the best cyanide recovery.

It really is nice to know what size you should be milling to before you start milling.

Doing pre-sizing etc in the mill circuit may seem a good idea but you get no advantage over just milling the unsized feed in total.

The final mill discharge is placed straight into a settling dam or series of dams. When the dam is full and the free water has been removed from the surface the material is ready to be machine mixed with a backhoe or excavator before being placed into the leaching vat.

Because the vat leach will have several cycles of wetting and drying there is occurring a lot of surface oxidation reactions on the surface of the ore grains.

In most cases this means that pretreatment is either not required or is an expensive waste of time for little gain.

Deano


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## mahalas

Deano
I am sorry but I mixed something up. ( English is not my first language ) If I am correct vat leaching for you means something like percolation closer to heap leaching. I try to focus on tank leaching with stirrer . It gives me more possibilities to control the process.

My first attempt with percolation ( 30 kg conc in the bucket ) showed me 33 ppm total gold left in tailing after leaching.
I assume that was because conc was to coarse. Now I am in the process of doing second try with finer batch.

My goal is to recover 90% and up of total gold, not only what is available for leaching. I am not sure if it is possible with my tailings, but at least I am going to try.

I am afraid that very fine grinding will clog percolation circuit that is why I prefer the tank with stirrer.
During leaching my ORP shows around - 50 mV , is it OK ? Does it make sense to measure ORP during leaching with cyanide?

Cheers Adam


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## Deano

If you run the vat as detailed in my earlier post you will not have any clogging of the tailings in the vat, this method I consider important enough to be available to all users.

The advantages of this method are recovery of all leachable gold both in the sulfides and in the quartz fraction and the simplicity of the circuit and operation.

Leaching of sulfide cons in a tank will not give you the recovery level that you are hoping for, it will always give lower recoveries compared to vat leaching.

You are also missing out on the values in the non sulfide fraction.

Running the sulfide cons in a tank will give you less control over the process than running a vat.

In a vat the entire ore is leached so the average grade of material in the vat is lower and the process requirements are correspondingly lower.

Tank leaching of sulfide cons requires the use of a leach promotor as well as the usual cyanide and lime if you are to get really good recoveries.

The leach promotor will give you problems in the zincing stage.

I have no idea how you expect to get the non-leachable gold, a leach circuit will only recover leachable gold.

During cyanide leaching you are really only interested in pH, cyanide level and dissolved oxygen level(DO).

Those are the things you need to monitor throughout the leach cycle

Deano


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## mahalas

Hi Deano.
I had to digest it. The more I think about it, the more I am convinced that you are right. Does not make sense inventing the wheel. Will do a few more tests to check what size gives me the best response for cyanide, and I will leach everything in vat.

Especially that on the mine site there is a 10 000 liter plastic tank I can use to build the vat for leaching.

Oryginally I had idea to mill the tailings as fine as possible and do the leaching in tank with stirrer. Additionally in pre treatment stage I planned to do aeration with ozone in hope to oxidize sulfides and liberate all gold for leaching .

You mentioned about leach promotor. How is it used and how does it work?

And thanks again for sharing your knowledge . It is priceless.

Cheers Adam


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## Deano

Leach promotor is a powder formulation designed to speed up cyanide leaching and break down the components of the ore to allow cyanide more access to the gold.

Think blends of peroxides and chelates.

In Australia the most used brand is Leachwell, others are available both here and overseas.

These promotors are expensive and do not give you any better gold recovery than a vat leach will give you.

Note that a vat leach will run for months whereas a tank agitated leach with a promotor will be days at the most.

Like all things in mining the use of a vat leach is a compromise between time and cost/recovery.

The vat will take longer but will give the highest recovery at the lowest cost.

This gives you a problem if you want to use your plastic tank as a vat.

It will hold roughly 20 tons of material only and will then be tied up leaching this 20 tons for several months.

This is why only concentrate vats are small, vats for run of ore material are much larger to allow economies of scale.

My advice is for you to add up all of the tailings you have access to so you have a total in tons.

Halve this and you will have an approximate vat size in cubic metres to handle all of your material in one go.

This vat is built as a two or more cell vat so that the drying and rewetting cycles can be practiced.

Please realise that your tailings will be only part of the leachable material which can be leached in the vat, there will be large quantities of waste rock which will not have a high enough grade to have been viable to process when the mine was being first worked but which will be a large resource for your vat leaching.

In many mines the waste piles are a greater resource than the tailings heaps.

In vat leaching you are going to be making money at 0.5 grams per ton leachable gold, this will be your cyanide leach assay value you are looking for.

Most of these waste dumps are best treated by running only through a jaw crusher before being placed in the vat, the cost of finer milling is not recovered in the leach cycle.


Deano


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## Deano

When using distillation processes there is always the possibility of sucking the distillate back up the tube and into the heating bowl. This generally does not end well.

I have always used what is called a soft plumb section between the end of the condenser and the container of distillate.

The soft plumb section can be made of just a piece of rolled cloth tied to the end of the condenser such that the inside of the cloth roll acts as a pipe for the distillate.

The end of the rolled cloth dips into the distillate in the receiving container so that the distillate can wick its way up the cloth and form a fully wetted tube of cloth from the end of the condenser to the distillate.

If something unexpected happens during the processing then the distillate cannot be sucked back up the cloth tube as this tube collapses in on itself.

If distilling aggressive liquids the cloth can be replaced with very finely woven soft polypropylene or ethylene cloth or even high density thin carbon felt.

Deano


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## Deano

Many people who chase alluvial gold are frustrated by the gold being in cracks in the bedrock.

Most of these people have not the permits or expertise to use explosives to break up the bedrock and thus gain access to this trapped gold.

There are non-explosive alternatives available to break up the bedrock which usually do not require either permits or expertise.

Systems such as Cardox and Nonex whilst not being classified as explosives do require some expertise to be used successfully.

The expansive mortar systems are different in that the operation requires a hole to be drilled and a container of grout to be mixed and placed in the hole.

It can all be done with a battery operated hammer drill and stirrer.

There are many brands of these systems available such as Expando, Novatech and Dexpan, Google "chemical rock breaking" for web sites.

These systems really do work silently and safely, the only noise is the hole being drilled.

Deano


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## kurtak

Doesn't sound anywhere near as much fun as Kinestick or Kinepac though :lol: (both of which I have worked with)

I have heard tell that the old timers would drive (dry) wooden wedges into the bedrock cracks & they would split the crack open by hydraulics as the wood expands from soaking up the water

Kurt


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## Deano

When recovering gold from filter papers there are two main options.

The first and most used is to digest the filters up in aqua regia and thus get the gold as gold chloride in solution.

Commercial plants will usually substitute cyanide for aqua regia.

The second is to ash the filter papers and to then digest the ashed residue.

Some gold is always lost in this ashing no matter how careful the operator is and what ashing system is used.

A scrubbing system is not really viable for the small scale operator.

A simple method of stopping these losses is to ash the filter papers in porcelain crucibles where a covering layer of activated carbon is placed over the filter papers.

The carbon layer thickness must be at least two particles thick, preferably three particles thick.

Only half fill the crucibles with papers before covering with the carbon.

The crucibles are placed in an electric muffle and furnaced overnight at 650C.

After cooling the ash can be digested in the solvent of your choice for gold recovery.

Any gold which was volatilised from the papers is immediately adsorbed on the activated carbon and is retained there until the gold chloro complexes are broken by temperature.

This occurs before the carbon is ashed so there is only particulate gold on the carbon during the ashing process.

Deano


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## Deano

I am presently processing metallic pieces resulting from large scale smelting of a high lead/silver accessory mineral type gold ore.

The pieces range in size from 5 to 500 grams each in weight and are generally 5 to 10mm thick, they appear like flat gray stones.

Metal levels are approximately 40% silver, 50% lead and 10% gold.

I have tried all possible combinations of commonish type solvents but have not been able to prevent the formation of passivation layers with one simple exception.

Not wishing to spend all of my time carrying out separate leaches in sequence in order to separate the metals I have opted for a 10% NaOH solution with 2% NaCN added.

This leach will slowly dissolve in total all of the metals, agitation is by an overhead stirrer with a plastic stirrer.

The pregnant solution is continuously pumped through a carbon column using a peristaltic pump, the return liquor is plunged into the leach bucket to maintain dissolved oxygen levels.

When the gold level in the return liquor is near the same as that in the leach solution the carbon is saturated with metals.

The carbon is then changed for new carbon and the loaded carbon is ashed overnight in porcelain dishes in an electric furnace.

The ash from the carbon is then digested in aqua regia and allowed to cool to precipitate out the lead and silver chlorides.

The lead chloride in the photo is white, it appears yellow from the gold solution, note that the silver chloride is the first to precipitate.

After filtration and rinsing the lead chloride is separated from the silver chloride by hot water rinsing.

Deano


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## Deano

When the gold industry changed from Zinc recovery of the gold to CIP, there were several reasons for the change.

CIP is cheaper and simpler to run as a process than zincing.

Theft of loaded carbon is less attractive than theft of zinced gold.

More importantly the recovery of gold by CIP is usually greater for most ores than zincing.

The reason for this is tied up with re-adsorption and solution tenors.

Gold complexes in a leach solution will both adsorb onto suitable materials in the ore and displace onto other suitable materials.

The degree and rate of these occurrences are mainly controlled by the solution tenor of the gold complexes.

The greater the gold tenor in the solution, the faster and more complete are the adsorption and displacement reactions.

Maximum gold losses occur when the gold tenor is high.


When zincing for gold recovery the gold from an ore must all be solubilised before the filtration stage.

This gives two options.

1 Maintain as high a pulp density as possible to minimise filtration and zincing costs, this results in a high gold tenor in the leach liquor.

2 Lower the pulp density to give a lower gold tenor in the leach liquor but have an increase in filtration and zincing costs.

Option 1 will ensure that maximum losses of gold from adsorption and displacement will occur, this is balanced by the cheaper operating cost.

Option 2 will ensure minimum losses of gold from adsorption and displacement but will have higher operating costs.


With CIP there is a continuous removal of gold complexes from the leach solution so that the solution tenor remains low compared with a zincing solution tenor.

This minimises the adsorption and displacement losses.


A similar technique can be applied to the leaching of gold from scrap material.

This involves the running of a continuous bleed stream of leach liquor through a gold adsorbent column or electrowin cell such that the gold tenor of the leach liquor is maintained at a low level.

If gold is being leached from scrap which has surfaces of copper or zinc to name just two common metals then there will always be competition between the leach solubilising the gold and the other metals displacing the gold from solution.

By keeping the gold tenor low the displacement reactions are minimised and the maximum gold recovery is attained.

Deano


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## Deano

Images which may be of interest are porcelain dishes containing loaded carbon which has been ashed in an electric furnace at 650C.

The left hand dish was sourced from carbon fully loaded from a 50% lead, 40% silver, 10% gold leach solution under high pH 13, high(1%) cyanide conditions.

The right hand dish was sourced from carbon fully loaded from a 70% gold, 20% copper, 10% silver leach solution under standard pH 11, 0.1% NaCN conditions. The red colour is from the copper which loaded on the outer carbon surface at a faster rate than did the gold.




This is a view of the carbon from the left hand dish before ashing, the blue/white material is lead reduced on the carbon.

The carbon from the right hand dish appeared to have no deposit on it before ashing.

Deano


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## anachronism

To add to Dean's already excellent ashing pictures here's some I did yesterday. The raw product was type 43 plated connectors and the ash contained mostly gold with traces of copper and nickel. Ashing really is the way to go for recovery from leach solutions over dropping with zinc IMHO.


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## patnor1011

Do you have picture how it look like after ashing?


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## anachronism

That is after ashing Pat.


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## nickvc

Looking good have to tell us the yield.


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## Deano

Following the write up of how a vat leach is set up these photos are a mix from two sites which together show the various stages.
It appears that I can only place and comment on 5 photos at a time so I will have to do several posts to fit them all in.
Note that the plastic liners used here are the old 0.1mm thick ones, now days you need to run 2mm thick liners which need heat welding.




Initial cuts for vat




Forming up walls, lengthwise slope from right to left, crosswise slope from left corner to pump sump immediately behind photographer.




Forming up walls




Start of liner placing, pump sump in position, crossfall from right side to left side, sump at lowest point.




Unrolling and placing liner, drum type sump in position, agricultural drain lines ready for fitting.


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## Deano

Two vats side by side, nearest one filled with ore prior to leach solution addition, second vat in process of being lined.




Left hand vat in leach cycle, right hand vat being constructed. In the right hand vat a holding layer of screened creek gravel is being placed over a white filter fabric sock covered agricultural drain pipe.




Close up of the drain covering in previous photo.




Placing plastic liner and filling vat with ore. OK for tracked machinery to operate on placed fill but not wheeled machinery due to compaction.




Lining and filling vat as a continuous process.


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## Deano

Lining and filling vat as a continuous process.




Filled vat immediately prior to leach addition.




Operating vat on left, second vat being lined and filled on right, levelled area for third vat in background.




Operating vat




Adjusting hoses while under bird netting.


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## Deano

Active leaching vat on right, note sump pump and carbon column at end of vat. New vat being constructed on left. Salt lake in background.


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## Deano

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


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## justinhcase

Yet again an amazingly informative set of post's.
Much thanks as always.
There are some spoil heaps from 1800's that I may have to take an other look at.


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## g_axelsson

Thanks for the pictures, very interesting!

Göran


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## Deano

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


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## Smack

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?


----------



## Deano

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


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## Smack

Thanks Deano.


----------



## 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


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## upcyclist

Deano said:


> 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.


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.

I'm tempted to try it with copperas, too


----------



## anachronism

Nah - copperas works best dissolved in hot water and a dash of HCl to clear it up to an emerald green colour.


----------



## Deano

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


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## nickvc

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.


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## kurtak

nickvc said:


> 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,



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


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## Deano

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


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## 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


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## 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


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## kernels

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.


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## g_axelsson

Deano said:


> 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.


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.

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


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## nickvc

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.


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## Deano

It should be noted that with a ferro cyanide leach the uv conversion to cyanide is, assuming enough uv is present as sunlight etc, pretty well completely to the cyanide side of the equilibrium.

In the case of the ferri cyanide leach under the same uv conditions the conversion to cyanide is minimal.

However this is an equilibrium condition and any effect which removes cyanide from the solution will cause some more of the ferri cyanide to convert to cyanide.

Locking up some of the cyanide as gold cyanide will effectively remove that cyanide from the equilibrium and some more of the ferri cyanide will convert to cyanide in order to maintain the equilibrium.

Thus a solution of ferri cyanide will have most of the ferri cyanide as ferri cyanide with only a small amount converted to cyanide.

This means that there is a reserve of ferri cyanide in solution available for the production of cyanide if any of the already converted cyanide should be locked up as gold cyanide.

If you have a lot of time available and wish to be using the safest possible cyanide type leach then use the ferri cyanide leach.

If you are willing to trade off some of the safety buffer for time then use the ferro cyanide leach.

Deano


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## Deano

I have never made any secret of my preference for leaching of ore in a vat rather than agitated tanks.

One further point in the favour of running a vat leach is the treatment options available for cyaniding sulfide ores.

Most easily treated oxide ores have been found and mined, what is still available are sulfide ores.

Many sulfide ore deposits have been not worked because the deposits are not large enough to cover the very large costs of CIP circuits in the cases where they can be successfully leached in cyanide CIP.

If it is decided that the ore should be concentrated into the sulfide fraction and then to treat that fraction there are losses and extra costs you will run into.

Unless you are extremely lucky you will be unable to get high recoveries of the sulfides by gravity separation.

This leaves flotation as the best option but will require a float plant and suitable milling.

Even an extremely well run float circuit will have sulfide losses at a level which hurts a small operator.

Big operators can use economies of scale to offset these losses, by definition small operators cannot.

A further problem is that most sulfide concentrates do not respond well to cyaniding without some form of pre-treatment such as roasting or bacterial oxidation.

This requires further plant expenditure and operator expertise.

If the head ore is run as a vat leach there are unexpected benefits. 

There are no concentration losses as there is no concentration.

Milling does not need to be tightly controlled, usually a jaw crusher followed by an impact mill of some form will do the job nicely. This is as cheap a milling circuit as you can get both in capital and operating costs.

It also changes the mining cycle requirements in that you no longer are attempting to have a steady supply of blended feed to the mill in order to maintain consistent pulp density and gold tenor through the CIP tanks.

Instead you can stockpile ore on the ROM pad in dry weather periods ready for milling before wet weather.

This often means that you can have the stockpiled ore toll milled on site by a large trailer mounted jaw/impact mill combo on an as needed basis, you do not have to purchase the milling circuit upfront.

Once the ore has been milled and blended it can be left in a holding dam until required to fill a new or rotated vat.

If the head ore does not respond well to straight cyaniding then a high pH caustic cyanide leach can be used to break down the sulfides and allow leaching of the gold.

Because this breakdown is slow, requiring weeks to months, it cannot be employed in a tank type circuit.

As you are going to be running the vat for several months anyhow the sulfide degradation time overlaps with the normal leaching cycle.

Because you are running a vat leach there is no loss of liquor from leakage in the tailings dam as you do not have a tailings dam as such.

The liquor is recycled through the vats as required.

The main extra expense with a vat leach is the cost of fitting rain shedding covers over the vats to prevent flooding of the vats in wet seasons.


Deano



Cyanide recovery is often acceptable without further treatment of the ore


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## Deano

One other thing I forgot in the vat leach post is regarding the pumping system.

Regardless of what type of pump is used there will at some stage be entrained air passing through the carbon column.

This means that you need to have a top screen to retain the carbon in the column.

Usually a 1mm hole diameter punched 316 SS screen will do the job well.

Deano


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## Deano

A further omission in the vat leaching post concerns the pumps used.

Many operators use self-priming centrifugal pumps, these work well but are not designed for extended use if a suction blockage occurs.

I have always favoured a diaphragm pump because they are almost bullet - proof.

However if using a diaphragm pump the liquor is presented to the carbon column in a pulsing flow pattern, this tends to attrition the carbon unnecessarily.

The flow pattern can be smoothed out by placing a vertical air vessel in the discharge line next to the pump outlet.

Generally the air vessel has a height 5 X the outlet pipe diameter and a width 3 X the outlet pipe diameter.

So if the pump discharge is, say, 50mm diameter then the height of the air vessel will be 250mm and the diameter will be 150 mm.

These dimensions are approximate but work well if the vessel is made from a piece of 150 mm diameter pipe.

If the pulsations in the liquor are still greater than you would like you can place a second air vessel in the line about a metre from the first vessel.


Deano


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## 4metals

I agree that a diaphragm pump is bullet proof and can run dry without any issues. Try that with an impeller pump! I have had issues with some particulates fouling the diaphragms and stopping the pumps, always solved with appropriate in-line strainers. But I do most of my work with recycled values not natural ores so I have more control over what crud may be sucked up. You have to deal with whatever nature throws at you in large outdoor pits. 

So your chamber is a pressure chamber which will pressurize as the trapped air is compressed. How do you accommodate air in the feed? There will come a point where the "pulse tank" will have too much air in it. Do you allow for pressure relief, either automatic or manual?


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## Deano

I really did not phrase the location of the air vessel very well.

For a horizontal pump discharge line the air vessel is fitted vertically upwards from a 90 degree tee piece in the discharge line.

This means that the air in the vessel is permanently retained in the vessel.

Air in the feed simply bypasses the air chamber as the chamber is already full of air.

If the feed has a lot of air slugs in it these can be gotten rid of by running the discharge line in a single upwards peaking wave shape and fitting an air release valve at the top of the wave.

The horizontal and vertical components of the wave are designed on the flow rate of the pump.

This will not get rid of all the air but will get rid of most and minimises its effect on the carbon.

Deano


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## justinhcase

Absolutely top notch.so much useful information in so few post's.
Thank you for introducing us to such a practical and fascinating's process as vat leaching .
Could you recommend some further reading? such as environmental impact reports or any well put together study.
Have you any published paper's your self?
As this is a relatively unknown process in my part of the world having some case study's to use in discussion would be of help.
The councillors in Devon are all in there sixties and need information in triplicate from several different sources before you can make any headway with them in discussion.
Much Thanks
J


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## 4metals

So kind of like this;





This will also allow a diaphragm pump to supply a constant flow rate at a constant pressure if you use a pressure switch in line. I have used systems like this in aquaculture setups to keep the column of packing where the denitrifying bacteria live aerated (by a constant flow of fresh water, not necessarily with air) and freely moving in the column. Never did this with carbon but it makes perfect sense. 

Thanks Deano, we truly appreciate your perspective from the mining world. You don't stumble over facts and tips like these every day, even when you do this for a living.


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## kurtak

4metals said:


> Thanks Deano, we truly appreciate your perspective from the mining world. You don't stumble over facts and tips like these every day, even when you do this for a living.



Very well said 4metal & I absolutely agree :!: 

Deano - your posts are absolutely :G 

Thank you Sooooo much for taking you time to share your VAST knowledge with us :!: 

It is much appreciated 8)  :!: 

Kurt


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## anachronism

I have recently had the extremely humbling experience of working with Dean for ten days. 

Three days in a cyanide carbon stripping plant and a week at his own lab. Whilst it was a long way to go from England, it was worth every dollar. The man's a genius. 

I could never thank him enough for the opportunity. 

Jon


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## kurtak

anachronism said:


> I have recently had the extremely humbling experience of working with Dean for ten days.
> 
> Three days in a cyanide carbon stripping plant and a week at his own lab. Whilst it was a long way to go from England, it was worth every dollar. The man's a genius.
> 
> I could never thank him enough for the opportunity.
> 
> Jon



This reminds me of another friend I have --- he lives in southern California - he calls me up in the middle of the winter when its like 20 below zero with wind chills of 35 - 45 below zero here in Wisconsin & tells me how he is out riding his Harley in a Tee shirt :lol: :lol: :lol: 

I am JEALOUS as hell mate :!: 8) :lol: :mrgreen: 

Kurt


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## Deano

I have always tried to keep pressure switches etc. out of my pump circuits.

They always give trouble at the most inconvenient times.

You can usually design a circuit which is totally basic, little or no electronics involved.

It may not look as pretty as the fancy circuits but the reliability is far superior.

Deano


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## Deano

There must be some papers around on vat leaching but I have never chased them as the method or running a vat I practice is in many aspects different to standard vat leaching practice.

That is the reason I put out these posts on vat leaching is to spread the knowledge.

Usually people running vat leaches are either too busy or not inclined or not capable of putting together a study.

Deano


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## justinhcase

Deano said:


> There must be some papers around on vat leaching but I have never chased them as the method or running a vat I practice is in many aspects different to standard vat leaching practice.
> 
> That is the reason I put out these posts on vat leaching is to spread the knowledge.
> 
> Usually people running vat leaches are either too busy or not inclined or not capable of putting together a study.
> 
> Deano


O well,If you are trying to put a proposal together it is always of help to have some one go before.
Trying to communicate the details of the process and answer proactively environmental concerns with out a successful model to work from will be interesting.
The first step in Europe would be to employ a specialist to conduct an environmental impact study and hope to high heaven that no one else funded a counter argument..it is very important to get the right specialist.Do you not have to produce such ground work in Australia.
Much thanks
Hope all is well in the land down under.
Regards
J


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## Platdigger

anachronism, that must have been soooo good! Do you think you learned anything? 
I bet you are wishing you filmed or somehow recorded every minute.


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## anachronism

Platdigger said:


> anachronism, that must have been soooo good! Do you think you learned anything?
> I bet you are wishing you filmed or somehow recorded every minute.



I learned an absolute boatload, starting with the fact that I actually didn't know very much :idea: 

The joys of modern tech mean that I was able to take a lot of pictures!


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## Deano

Vat leaching has been practiced in Australia in large scale for many decades.

The main requirements are vat design particularly in relation to liner type and properties and wall construction.

Other requirements are similar for vat or tank leaches.

Deano


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## 4metals

I understand that the quantities of material processed by vat leaching can greatly exceed the quantity processed in an agitated vat but how does one ascertain that all of the contents of the vat are effectively leached. Is it possible that a design can eliminate the channeling effects often seen when flowing a leach over a granulated pile of scrap material?


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## Deano

Channeling effects in a vat are eliminated mainly by the hydraulic lifting approach to flooding the vat.

If you have been diligent in blending the feed then your main concern is to wet all ore particles, hydraulic lifting does that for you.

When the vat is being rested the ore is fluffed up with an excavator, this removes any problem areas which have formed during the leaching cycle.

The cycle of leaching and resting a vat is maintained until there is not enough gold recovered from a leach cycle to maintain viability.

Testwork carried out on exhausted vats register residual leachable gold levels of well less than 0.1ppm.

Deano


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## 4metals

So it is a multiple leach process, flood, drain, fluff, test and flood again if necessary. I'll bet the job of driving an excavator through a damp pit of cyanide soaked ore isn't one people are lining up to do. No sipping coffee on that job!


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## Deano

The fluffing step is carried out after the pit has been drained for a week or two.

The excavator does not actually go into the pit, it always stays on top of the ore.

The cyanide level is correspondingly low, you have to work hard at even smelling its presence.

Keep in mind that OHS regulations must be adhered to, no one wants their project closed down through bad work practices.

In my years of vat leaching I have never even heard of anyone suffering accidental cyanide poisoning.

You respect cyanide but you do not fear it, you base your work practices around that approach.

Deano


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## Deano

An unexpected source of base and precious metals is often found in copper slag.

The early copper miners appeared to have two flaws in their smelting.

The first was that they did not know the best practices to follow when they smelted.

The second was that that they took the attitude of "there is plenty of rich ore available, if we do not get all of the metal recovered from the smelt it does not really matter, there is plenty more available".

This means that many of the early slags contain high levels of copper prills, most starting at 1% with some going up to 5% and more by weight.

Depending on the ore used, many of these prills had high gold levels up to 200 ppm.

The gold equalled the copper in value.

The cheapest way to recover these prills is by milling and gravity separation.

Generally it is not worth trying to mill to finer than all passing a 2mm screen.

Milling is done by sequenced jaw crushers with the final size obtained by choke feeding the last crusher.

Most of the prills are spherical but will still separate into a heavy concentrate on a well operated table.

Best table recovery is achieved if the milled slag is sized at 2mm and 0.5mm, these fractions are tabled separately.

Any prills left in the slag are not worth the cost of finer milling needed for liberation.

Ball or rod milling alters the prill shape and lowers the gravity recovery levels as well as costing more.

A general idea of what is available in a slag can be gotten by hand pulverising and then panning.

Deano


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## 4metals

> Ball or rod milling alters the prill shape and lowers the gravity recovery levels as well as costing more.



This is a descrete fact that can save someone starting out a lot of time and money, both on equipment and in time. It makes sense that a flattened bead would behave differently on the table, but to have first hand experience having doccumented that fact..... Well it is valuable!

Deano, you Aussie's are OK! Thank you.


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## Deano

While I was aqua regia digesting some ashed carbon I became curious as to exactly how much gold was co-precipitated with both the silver and lead present in the leach.

Both the silver and lead were, as expected, present as chlorides in solution in the aqua regia.

As the de-noxed aqua regia was allowed to cool down the silver and lead chlorides precipitated out of the solution and formed a layer on the bottom of the beaker.

There was enough of a density difference between the lead and silver chlorides that with extreme care a separation could be done on a super panner.

1 gram samples of each chloride were dissolved in separate fresh aqua regia solutions and the resulting warm solutions were analysed for silver, lead and gold.

The silver:gold ratio was 1:200, the lead gold ratio was also 1:200.

The gravity separation was surprisingly good in that the cross contamination between the lead and silver fractions was also in the ratio area of 1:200.

I suspect that there were also co-precipitation effects between the silver and lead chlorides.

A new artificial 1:1 blended sample of the gravity separated chlorides was treated to a dissolution of the silver chloride in sodium thiosulfate, this indicated the silver to gold ratio was around 220:1.

The lead chloride residue from the thiosulfate leach was analysed and showed a lead:gold ratio of 180:1.

It appeared that some of the gold released by the sodium thiosulfate dissolution of the silver chloride was adsorbed onto the lead chloride complexes.

A hot water contact between separate metal chloride samples revealed that both lead and silver chlorides desorbed gold complexes in the ratio of 10:1 metal:gold when related to the minor dissolution effects of the metal chlorides.

In other words there was a flush of gold from the hot water contact in a quantity in excess of that expected from a dissolution of the chlorides.

Aqua regia dissolution of the residue from the hot water flush gave a metal chloride:gold chloride ratio of 300:1 for both silver and lead.

This appears to indicate that the gold and silver and lead chlorides will co-precipitate from a chloride solution but the greatest mode of attachment is by adsorption of the other still dissolved chloride complexes onto the precipitated chloride complexes.


Deano



This indicates that the metal chloride mix


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## Deano

Let me start by saying that I am not interested in getting into the old argument about whether fire assay in ores reveals all of the gold values.

It has been my experience that in many ores where the fire assay values are zero it is possible to recover ultra fine gold by gravity means.

This is done by hammer milling ore samples and panning a < 100 micron screen undersize.

The best pan for such an exercise is a blue non-riffled pan, the gold is so fine that the agitation from a riffle will cause most values to be lost.

The pan concentrates can be fire or aqua regis assayed and will show values which were not revealed in an assay of the head ore.

A common constituent of such ores is iron in most of its forms.

The above is the reason for many small old workings on areas in shed zones from iron rich formations.

The original prospectors did all of their sampling by milling and panning, usually they were testing quartz stringers in the hope that they would thicken and contain high gold levels.

Modern sample testing does not reveal gold values which the original prospectors were following.

There are many examples of old workings where hundreds of shallow pits were dug but no major excavations occurred.

Most people assume that the old prospectors were either stupid or that they got all of the available gold.

The fact is that neither occurred, the grades were not high enough to be viable but they were present. No body digs holes for fun, these people depended on finding enough gold to live on. They found enough values to encourage them to look further but not enough values to actually mine.

Many such areas respond to metal detecting for small nuggets, most of the eroded values are found only in shed zones where the water flow is so slow that these values can be deposited, these zones can be a long way from the original source ore.


Deano


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## 4metals

> It has been my experience that in many ores where the fire assay values are zero it is possible to recover ultra fine gold by gravity means.



While I do not doubt this statement based on your life experience with the mining industry, and I do not in any way wish to cast doubt on what you have said. But there was a former member named Dr Poe who claimed that colloidal gold was everywhere and while it couldn't be assayed by conventional means, he could recover it. You, sir are by no means a Dr Poe, and if you say it is quantifiable, I believe it.



> The pan concentrates can be fire or aqua regis assayed and will show values which were not revealed in an assay of the head ore.



The problem I have with a statement like this is not that gold which escapes detection by classical fire assay techniques is actually in some ores, it is that members who believe to have ore which is valuable will have yet another excuse not to bother with a fire assay. Your lab, Deano, enables you to crush and sieve the material, perform a classic fire assay, and pan and assay any suspect on your AA. Most assay labs wouldn't go that extreme once a classic fire assay, fluxed properly for the ore composition of the sample, results in a non detectable answer. 

It does beg the question however, what range of PPM's gold concentration have you detected by this method? I can understand it being detectable and not being economically recoverable on a large scale. Are you talking about levels of Gold which can be profitably extracted?


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## Deano

The head grades of the ores which respond only to gravity separation of gold from the milled ores vary from 0.5 to 3 ppm with the majority in the bottom end of the range.

It is very much dependent on the ore tested, generally ores from a single formation carry similar grades for separate samples but similar looking ores from a separate formation may have wildly different values.

I have not seen a highly profitable ore body of this type, most would fit in the marginal at best category.

It has to be kept in mind that the ore has to be mined, milled and then gravity processed before the concentrates are treated as a separate parcel, even with economies of scale there are no cheap processing steps involved.

An indication of the difficulty of gravity processing this type of material is given by my inability to recover this type of gold from creek bed formations which would usually be regarded as guaranteed gold traps, this gold is only found in very low flow speed situations.

Very few people have the panning skills to recover this gold, the top size of the gold particles is around 5 microns with the bulk being much lower.

I wrote about this ore type because of an argument some prospectors were having about large areas of obvious prospect holes in places where gold mining was not existing either now or in the past.

Many of these areas have been extensively rock chip sampled by mining companies and the gold grades were elevated in the low ppb range but not to such a level that it was thought to be worth further testing.

The sheer amount of effort involved in the digging and testing of these sample holes was such that it was impossible to believe that it had been done without the encouragement of seeing gold in the samples.

In many of these areas there were many hundreds of these prospect holes in a few acres, this represented months of work.

Most of these holes were dug in the mid to late 1800s so the only technology available to the miners was pick and shovel followed by dolly pot and pan.

If you treat the ores in the manner of the original miners you could see gold in the pan, carry out modern sample handling and assay methods and you get nothing.

You still have to do the assays on the pan concentrates to get a head ore grade and these assays can only successfully be done on a really clean pan concentrate. It really gives me much admiration for the panning skills of the original miners.

I have always suspected that most of these early miners were hard rock tin miners from Cornwall in England, they would have been among the few people who had the panning skills to even see this gold.

Deano


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## 4metals

Back in the day I was using a dredge in Ecuador to recover placer gold from streams. Before I traveled to Ecuador I visited Keene Engineering to purchase a dredge and have it shipped to Guayaquil so it would be there when we arrived. The salesman convinced us to select the model which was designed to also recover flour gold. Back then flour gold was a term I had never heard as I came from the secondary refining industry but I took his advice and purchased a 6" dredge which had riffles to collect "flour gold" as well as placer nuggets. 

I must admit the difference between moderately successful and very successful was flour gold. When cleaning up the dredge, the fine powdery sandlike gold appeared when panning. It appeared like the finest beach sand I have ever seen except it was gold in color. 

Apparently the flour gold that I recovered were mini boulders compared to the gold that Deano is talking about. When you consider the flow through the riffle, which wasn't hard enough to blow the flour gold off the dredge, it is hard to imagine just how small the particle size is on the values Deano is talking about. And I thought the dredge made to trap flour gold got the smallest of the small! Now I have to wonder if any went off the south end of that dredge. 

Oh well, let it go, we still made money!


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## cosmetal

4metals said:


> Back in the day I was using a dredge in Ecuador to recover placer gold from streams. Before I traveled to Ecuador I visited Keene Engineering to purchase a dredge and have it shipped to Guayaquil so it would be there when we arrived. The salesman convinced us to select the model which was designed to also recover flour gold. Back then flour gold was a term I had never heard as I came from the secondary refining industry but I took his advice and purchased a 6" dredge which had riffles to collect "flour gold" as well as placer nuggets.
> 
> I must admit the difference between moderately successful and very successful was flour gold. When cleaning up the dredge, the fine powdery sandlike gold appeared when panning. It appeared like the finest beach sand I have ever seen except it was gold in color.
> 
> Apparently the flour gold that I recovered were mini boulders compared to the gold that Deano is talking about. When you consider the flow through the riffle, which wasn't hard enough to blow the flour gold off the dredge, it is hard to imagine just how small the particle size is on the values Deano is talking about. And I thought the dredge made to trap flour gold got the smallest of the small! Now I have to wonder if any went off the south end of that dredge.
> 
> Oh well, let it go, we still made money!


I know that this can be a controversial subject, but, here it goes . . . 

Back in the early ‘80’s I had the good fortune to live in Placerville, CA eight miles from Coloma, CA where the gold was found that started the 1849 gold rush.

I had four placer claims on the South Fork of the American River. My partners and I dredged these claims for years (not full time – we found good gold but not enough to support three families (gold hit $850/oz. in 1980 and fell to $599/oz. by the end of 1981).

I’ve always been interested in “micron”. “flour” or “”float” gold – however it was defined. When we cleaned out our suction dredges (panned by the river), we could see this very small gold but it didn’t seem to be worth chasing even though we produced a lot in our concentrates.

One winter came and I decided I was going to build a concentrating table to be housed in my garage where we could process these fine gold concentrates at our leisure and in comfort.

It ended up being approximately 46cm wide x 183cm long. Similar to a miller table except mine was also equipped with (here it is!) a 46cm x 46cm copper plate electroplated with silver that I would coat with mercury! Yes . . . mercury.

Before this post is condemned, please know that I am not encouraging the use of mercury. I was also raised by ex-copper miners and geologists who taught me how to handle mercury and I knew about its toxicity. My table also had a final riffle section at the very end where I had mercury traps and miner’s moss riffles.

The table consisted of controllable water flow, an upper section coated with chalk board paint, an upper miner’s moss and riffle section, the mercury coated plate, a lower miner’s moss, riffle and trap section, and the final catch basin that was suspended in a large water basin before the water was allowed to drain onto my pasture area.

Not only did the table cut out our river side panning, it caught a multitude of gold sizes - even the Deano sizes. In fact, when I started it the first time to adjust the water flow, I did not run any concentrates for a while as I fine-tuned everything. Maybe two hours or so later, I was amazed to see an extremely fine sheen of gold over all of the coated plate! It wasn't a optical illusion as I could actually move it with my gloved finger. Yes, there is colloidal gold suspended in water – at least in El Dorado County, CA there is! Fortunately, i never had any mercury escape. 

Again, please do not use mercury. As Deano says, the smallest gold can be panned but it takes practice, practice, practice and patience, patience, patience! :|


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## anachronism

Thanks for the post cosmetal. I think it's good information, and let's face it, it's a pointless exercise to condemn talking about a method that was the used in the past just because today's standards are different.


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## snoman701

anachronism said:


> Thanks for the post cosmetal. I think it's good information, and let's face it, it's a pointless exercise to condemn talking about a method that was the used in the past just because today's standards are different.


I find it interesting to review history. Sometimes it's relevance isn't that far off.


Sent from my iPhone using Tapatalk


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## 4metals

Back when I was recovering placer gold, those things were called plate amalgamators. The silver plating made the mercury more adherent than a plain copper plate which helped to prevent the mercury from breaking up into tiny balls and being lost. 

We had a plate amalgamator in Ecuador as well as a mercury retort to recover the mercury. Mercury was sold in 50 pound dewars and widely used back in the early '80's. 

Mercury will form an amalgam with most metals with the exception of aluminum and iron. The ability for mercury to form an amalgam with gold but not iron may be the reason it can separate the gold from iron rich black sands if they are crushed enough. 

Mercury was so overused in placer mining in Ecuador back then that it was not uncommon to see small pools of mercury near the outwash of the mills. It had to take an environmental toll. We are better off today to steer clear of its use.


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## cuchugold

This is a great thread to read and reread. 
Q for Deano: Any thoughts on the use of resin instead of carbon for the recovery of the liquor?. My experience is only using small 20 Ton vats, and found that while resin is more expensive, it is reusable and more effective.


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## cosmetal

4metals said:


> Back when I was recovering placer gold, those things were called plate amalgamators. The silver plating made the mercury more adherent than a plain copper plate which helped to prevent the mercury from breaking up into tiny balls and being lost.
> 
> Mercury was so overused in placer mining in Ecuador back then that it was not uncommon to see small pools of mercury near the outwash of the mills. It had to take an environmental toll. We are better off today to steer clear of its use.


*"The silver plating made the mercury more adherent than a plain copper plate which helped to prevent the mercury from breaking up into tiny balls and being lost."*

Much easier . . .

*"Mercury was so overused in placer mining in Ecuador back then that it was not uncommon to see small pools of mercury near the outwash of the mills. It had to take an environmental toll. We are better off today to steer clear of its use."*

We would find small pools hiding in the cracks within the bedrock. They had been there so long that they were usually very pregnant with gold flakes and gold dust.

James


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## cuchugold

Like it or not, most of the gold from Brazil and Venezuela come from amalgamation in one fashion or another. It's just very effective enhancing gravity methods. It's very common "charged mercury", which is an amalgam of sodium, which keeps the surface very soapy and clean, facilitating capturing the gold flakes. In an operation I supervised, we retrieved more mercury from tailings processing than we used.


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## snoman701

cuchugold said:


> Like it or not, most of the gold from Brazil and Venezuela come from amalgamation in one fashion or another. It's just very effective enhancing gravity methods. It's very common "charged mercury", which is an amalgam of sodium, which keeps the surface very soapy and clean, facilitating capturing the gold flakes. In an operation I supervised, we retrieved more mercury from tailings processing than we used.


If you were in the us, you could charge for environmental remediation!!


Sent from my iPhone using Tapatalk


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## Deano

Re use of resin instead of carbon for gold recovery from solution.

Resin will do a better job of adsorbing the gold complexes from cyanide solution than will carbon. It adsorbs faster and to greater completeness.

The latest resins from large producers will also strip quickly and easily.

So why are not all of the worlds gold mines using resin rather than carbon.

The initial resistance is price, resin costs multiples of carbon cost.

The second is osmotic shock in those resins still using a sulphuric/thiourea strip, the degradation rate can be fierce.

The third is inventory lockup of the gold. Resins are notorious for locking up gold which cannot be eluted, the longer these resins are used the greater the lockup level.

This means that apart from the lower adsorption levels per cycle there is an increasing amount of gold which has been processed but which cannot be recovered until the decision is taken to recover this gold by ashing the resin.

A further strike against resins is that, as with carbon, the initial and fastest adsorption occurs on the outer surface of the adsorbent particle.

The organic functionalities of the resins are subject to uv degradation and these functionalities will degrade to higher forms which are more difficult to strip and/or will degrade to forms which are less effective as adsorbents.

So with time the loading kinetics of the resins drop markedly as the surface sites are no longer active as adsorbent sites.

With carbon the greatest problems are associated with attrition of the carbon and the loss of carbon and gold values as fines.

Most of these losses occur during regeneration of the carbon and so the gold losses are minimal as the carbon has already been stripped.

Deano


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## cuchugold

Thanks a lot Deano!. :G

edit to add: Can you point to any links or post any knowledge regarding the recovery of gold from ores in regions with little or no water available?. I understand there are several desert/ near desert regions in Australia with plenty of good material.


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## Deano

There are many gold bearing areas in Australia which are in desert or semi desert conditions.

Usually the main problem is not availability of water but availability of reasonable quality water.

Often any better quality water is only available in small quantities so it is apportioned initially for drinking/ camp use and any surplus is shandied with lower grade water in an attempt to improve the plant water quality.

It is usually not viable to attempt to improve the water quality by RO etc.

Deano


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## cuchugold

Thank you Deano.

I'd very much appreciate your input with regards to recovery in this situation:
http://goldrefiningforum.com/~goldrefi/phpBB3/viewtopic.php?f=44&p=278491#p278491


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## Deano

Judging by the queries from forum members there is a need for basic gold analytical information regarding ores.

It should be noted that there are always specialised exceptions to the general rules but these exceptions are unlikely to be encountered by the average prospector.

Most queries start along the lines of how much gold is in the ore sample.

An experienced person can estimate the grade of visible free gold in an ore sample. They know what gold in ore looks like and from having previous samples assayed they can estimate the grade.

If gold is not obviously present as coarse free particles then it is necessary to call in a professional assayer.

There are three types of assay for gold which are applied according to what the prospector wants to know.

Keep in mind that most labs charge a fairly hefty batch fee to cover the cost of doing paperwork so it pays to look at pricing to see if it is near the same cost overall to do more than one assay type on an ore sample.

The first assay type is the standard fire assay. This will tell you how much gold is present in your ore and is the basic assay format. Unfortunately it does not tell you anything more than the gold grade. An experienced assayer should be able to look at an ore sample and identify the presence of sulfides and anything else obvious which may affect the fire assay and use the appropriate assay flux accordingly.

The modern trend is to use only two flux types covering ores with and without sulfides present, this has been brought about by the shortage of assayers able to formulate fluxes and the high cost of doing these formulations compared with just taking a scoop of premixed flux from a commercially purchased container of such.

That said, the standard flux formulations will give you a go/no go result for possible further work on a sample.

Always keep in mind that human nature is to take the best looking samples and have them assayed so your first samples assayed should be treated as indicative only, the results are not necessarily applicable to the whole deposit.

So now we have done the basic fire assay and registered gold in viable quantities. The next thing we want to know is where in the ore sample is the gold; is it in the quartz matrix for example or is it associated with sulfides or is it all as free gold.

Most labs will offer you a range of acid type assays along with reasons for having them done on your samples.

For the average prospector just wanting some basic information these assays are bewildering and expensive.

What is required is a basic aqua regia digest assay. This assay type will show gold which is either available as free or exposed gold particles or is in a matrix such as sulfides which can be dissolved by the aqua regia.

The aqua regia values are usually slightly less than fire assay values due to fire assay being able to access the gold locked in ore particles which the aqua regia cannot access.

If we have a large difference in values between fire assay and aqua regia digest then the usual reason is that the bulk of the gold is encapsulated in a matrix which aqua regia cannot dissolve such as quartz.

In the above case we now have the information that there is a strong possibility that fine milling of the ore may expose enough gold values to allow viable treatment of the ore.

What the above tests do not tell you is if the gold is free or is in a matrix which aqua regia can dissolve, both the fire assay and the aqua regia assay will show you the above gold values.

Often gold in sulfides is in the form where it is as small particles down to atoms locked inside the sulphide crystal matrix. You are not going to recover much of this gold without either smelting the sulfide in the presence of a collector metal or by chemically destroying the sulfide matrix.

The third form of assay has two uses; it tells you what gold recovery can be expected with cyanide and it tells you what level of the gold is locked in sulfides or similar by comparing the results with those of an aqua regia assay.

This assay form is a cyanidation assay, it is usually offered in scale from a few grams to several kilograms.

The smallest scale is often a few grams of ore stirred in a beaker but it is usually run as a bottle roll of 50 grams plus of ore run for 24 hours.

Because the bottle roll is an alkaline leach it has little dissolution effect on most but not all sulfides.

If you have highish gold values from fire and aqua regia assays but low gold values from cyanide then it is a fair assumption that the gold is locked in sulfides and that recovering the sulfides alone will recover most of the gold. This will give you a concentrate which can be sold or treated in a separate stage.

None of the above is totally definitive but is the basis for test work to set out an operating plan.

Deano

It will at least let you know if doing further work on a deposit is warranted and will do it at the lowest cost.


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## g_axelsson

Thanks for the run-through of ore assay.

I wonder where in this scheme does tellurides and other sulfosalts of gold end up? Is the aqua regia solution strong enough to break down tellurides?

We have an interesting mine close by (about 150 km, suitable for a day trip).
- Kankbergsgruvan (Kankberg Mine) where most of the gold is locked up in 10-15 um large grains of sylvanite and calaverite in a matrix of mostly topaz and quartz, the ore is grayish white to pale blue from topaz. Visible gold is really rare. Boliden have a secret process to process this ore that also recovers the tellurium. It was a great thing a few years back and they increased the world tellurium production with 10% but since then the tellurium price have crashed.

Göran


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## Deano

Somehow when I wrote about recovering gold from slag I left out the treatment of the table tails.

By jaw crushing and tabling the minus 1mm jaw product you will recover all of the coarse free gold prills.

There will still be substantial quantities of both fine gold prills and superfine gold prills in the table tailings.

The fine gold prills can be recovered by ball milling the table tails to 100% minus 100 microns and then running this product slowly over the table with minimal wash water.

Depending on the slag used, the recovery of fine prills usually runs between 300g to 1 kg of prills per ton of slag tabled.

The tailings from the second tabling are really only worth leaching as a gold recovery method, any further milling will require running the table as a slimes table with the very slow feed rates that come with that format.

When looking at what leach type you are going to use on the final table tails you must always keep in mind that the reason the prills are present in the slag in the first place is that the gold was dirty when smelted.

Certainly the slag will remove substantial quantities of metallic impurities from the gold during the smelt and thus allow the cleaned gold particles to coalesce into a pool.

There will, however, be a substantial quantity of gold which becomes coated with the metallic impurities and cannot coalesce into the gold pool.

Acting under a range of physical restraints, this coated gold forms prills which are distributed throughout the slag.

The size of these prills ranges from several millimetres to well under 1 micron.

This means that in order to recover all of the gold present in the slag you are looking to mill the slag down to single micron size.

This will require the use of either a stirred mill or an air mill or similar.

This type of equipment is both expensive to buy and use so you are looking at how you can get the best recovery for the cheapest cost.

Jaw crushing is the cheapest method of size reduction with minimal amounts of shape deformation, thus allowing fast tabling with good recoveries of any coarse prills.

Ball milling allows you to reduce the slag size to the point where you can still table liberated fine sized prills because the deformation of these fine sized prills is minimal.

The prills present in the slag after ball milling usually run a gold grade of 200 to 500 grams per ton of slag.

You can carry out milling and leaching tests to see if it is worth carrying out the more exotic milling methods before leaching.

Unless you have access to large quantities of slag, the further milling is usually not worthwhile.

It must be kept in mind that the prills are only present because they are coated with base metals, most of these base metal coatings are resistant to cyanide leaching.

Leaching is usually only carried out on the slag after all gravity recoverable gold has been recovered.

As a general rule the coatings can be dissolved in chloride liquors.

This allows for the dissolution of the coatings in HCl, a filtration step with a water wash, pulping the residue with lime or caustic and then carrying out a cyanide leach.

This sequence is going to be cheaper than carrying out an aqua regia or chlorine leach and safer than both aqua regia or chlorine leaches.

Deano


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## 4metals

Another option could be smelting the slags with copper in a rotary furnace and producing copper anodes for electrolytic recovery. This eliminates milling but often employs equipment not available on site. (Electrolytic copper cells) 

With the quantities of gold you listed, 300g to 1 kg per ton, it may be easier to install a rotary furnace and produce assayable copper bars to be refined elsewhere. As with all gold containing feeds, the ability to quantify the gold content on site by producing the copper based bullion will yield the maximum payout.


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## Deano

A lot of fairly smart metallurgists have attempted gold recovery from slag using molten copper in various configurations.

Basically they all failed, gold recovery was minimal.

It appeared that the coatings on the gold which prevented the gold from coalescing into the original gold pool also prevented the prills from being adsorbed into molten copper.

Thus the use of physical methods to get a concentrate which can be attacked by a chloride matrix in a cleaning step before further smelting, the residue from fine milling and HCl attack and then leaching still contains substantial gold values.

Deano


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## Deano

Further to the treatment of slags I was recently involved in the setting up of a small vat leach system for treating the slag after the ball milling step mentioned in a previous post.

We could not get enough further values by air milling or stirred milling down to single micron size to justify anything bar a ball milling to 100% minus 35 microns.

The milled product was leached in a small vat 1.5 metres diameter and 0.5 metres high. This vat size took a charge of 500 kg milled slag.

A 2.5 metre length of 19mm grey water dispersion hose was used as the collection hosing, this was covered with polypropylene filter fabric which was retained with cable ties.

After tying off, the ties were cut neatly so that the hose could lie flat on the base of the vat.

The blind end of the hose had a poly plug siliconed in and a 90 degree 13mm poly bend was siliconed in at the other suction end.

The hose was positioned on the floor of the vat and a length of 13mm poly hose was attached to the poly bend, no clamps were used as the pressure on the join is negative.

Note that no metallic components were installed in the wetted area of the vat, this allowed the use of a hypochlorite leach if necessary.

The hosing on the floor of the vat was then covered with 150mm of concrete sand and a coarse filter fabric was laid on top of this sand before the milled slag was placed in the vat.

The function of the filter fabric on top of the sand is to act as a demarcation lining so that when the slag was being emptied from the vat after leaching it was easy to stop shovelling when the fabric was reached and no disturbance of the sand and collection hose occurred.

The 13mm poly hose suction line was fed to a speed controlled peristaltic pump which then fed the liquor into a vertical carbon column.

A tapped off take point was installed in the line after the pump and immediately before the carbon column to allow for sampling to monitor the gold tenor of the liquor before carbon adsorption.

By having a loose coil of the poly pipe between the pump and the carbon column it was possible to minimise the pulsation of the pump flow, only a slight movement of the carbon was observed.

The liquor overflow from the column was returned to the leach vat via a 20mm poly pipe, the flow was plunged into the vat liquor.

Provided a bubble trail extended from the return stream entry point as shown in the photo, the dissolved oxygen level in the vat liquor was kept high for efficient leaching.

The carbon column had 7 kg of washed carbon placed in it after the circuit had only water pumped through it for 24 hours as a leak check.

The flow rate of the liquor up the carbon column was set at 70 litres / hour and the first few minutes of flow was run to waste to get rid of any carbon fines generated by the placement of the carbon in the column.

A cyanide level of 1 gram/litre sodium cyanide was used and the vat was run for 4 weeks. At this stage the gold in liquor levels had dropped from an initial 60ppm to 1.2ppm and the daily return was approaching break even stage.

The liquor was then pumped into a drum for further use and after two rinse runs each of three bed volumes a hypochlorite leach in 20% salt was introduced into the vat. The calcium hypochlorite level was 10 grams per litre at pH 6.5.

The carbon was changed between the two runs to prevent reagent wastage.

After 1 hours pumping the gold tenor in the leach liquor had raised to 7 ppm and it maintained around this level for two weeks before tapering off to just under 1ppm after 3 weeks.

The Eh of the solution remained around 980 mv for the entire time, pH was adjusted with HCl

The liquor was then pumped to storage, the vat subject to a rinse run, the vat solids shovelled out and a new batch of material placed inside and the cycle was started again.

The method of leaching is capable of being scaled up or down depending on what application is proposed.

By keeping the wetted surfaces non-metallic either form of leach could be run with no problems.

Deano



Suction hose placed on floor of vat with pump suction line attached.



Suction hose covered with sand and filter fabric sheet in place immediately prior to filling vat with slag.



Vat partially filled with suction line to pump



Peristaltic pump with speed controller


----------



## Deano

Further vat leach photos



Carbon column being leak tested




View of circuit, note that the pump and speed controller kept weatherproof in shed while vat and carbon column are in open air.




Sampling point at base of carbon column



Return liquor line from carbon column, note bubble trail showing high dissolved oxygen levels in vat liquor

Deano


----------



## Platdigger

This is just soo good Deano, Thanks!


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## 4metals

Thank you for posting these descriptive photo's. A picture always clarifies things! 

I assume the milled slag has been screened. Was the gold recovery from the oversize substantial or is the lion's share of the value small enough to pass the screen?


----------



## Deano

Recovery per ton of slag.

Oversize beads and strange shapes >1mm. = 1.3kg

Tabled beads < 1mm = 1.8 kg

Leached in vat, milled to 100% < 100 microns, cyanide leach = 280g, followed by CaOCl leach = 180g.

Other slags have similar grades of gold, much depends on the ore type and the processing at the mine including the smelting flux formulation.

Deano


----------



## g_axelsson

Very interesting setup, that's on a scale that even I could handle.  

Last week I binge-watched the TV-series "Aussie Gold Hunters" season 2. In episode 10 (originally aired a couple of months ago) one team constructed a heap leaching pad and it was shown in quite good detail. One of the problems they encountered was (according to them) the water quality, a lot of crystals formed in the pipes, clogging the system. The water source was a deep well at the site (about 40 km outside of Kalgoorlie) and they suspected that it contained too much dissolved minerals and that interfered with the cyanide.
At that point they shut down the system, ashed the carbon and smelted the ash to recover the gold.

Now to my question... How does the water quality affect the cyanide leaching as long as the pH is controlled? Does hardness of the water affect it? What substances in the water can interfere with the cyanide leach.

Of what I remember from my brief visit in Kalgoorlie 15 years ago was that the deep ground water in the area was heavily contaminated by arsenic and salt. All fresh water was brought in via a pipeline or by road.
Maybe their problems was only evaporation and the crystals formed was from ordinary salt. 

Göran


----------



## anachronism

Kalgoorlie was one of the highlights of my trip to Australia last year Goran. What an incredible place it is, although I'm sure it may have changed somewhat in the 15 years since you went. 

ps I've actually leant against that exact shed wall with a cig in one hand and a bottle of beer in the other discussing how little I actually know about refining with Dean. 8) 

Something I will never ever forget. 

Jon


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## 4metals

OK I can imagine members drooling on their screens from these numbers. At $1320 gold we are talking about a recovery worth $151,000 from a ton of slag!

So to put this into perspective, how many tons of ore are processed (and approximate assay of unprocessed ore) to generate one ton of slag. From a refiners standpoint I'd be saying screw the raw ore let me process slags for a living! 

Surely the slag generating 114 ounces of gold as in this example was from an ore body representing a much larger quantity of gold. Do you have any idea of the quantity of gold generated from the original ore resulted in slag of this quality?


----------



## Deano

The ratio of slag to bar metal is very operator and mine dependent, but a very general approximation is 2:1 without any allowance for metals transferred into the slag such as iron and copper.

So look at needing slag from a high producing mine to generate appreciable quantities of slag.

If you accept that the gold grade into the plant approximately equals the amount of recovered metal then a 20kg bar will be recovered from processing 4,000 tons of 5 gram ore.

At a through rate of 100 tons/hour you are looking at 40 hours processing, call it 2 days worth.

So 10 days per 100 kg metal or 50 days for 500 kg metal = 1 ton slag.

There is very little romance in the treatment of slag, it is a dirty, dusty process you go through only because the rewards are good.

This pretty well sums up gold mining in general.

Deano


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## Deano

One of the main problems in gold mining areas, notably Kalgoorlie, is that the ground water is often contaminated with very high levels of salts, especially sulphate ions.

This means that using lime for pH control in such areas will result in the generation of gypsum and precipitation of same.

Some clever systems of lower pH cyanide leaching were developed to overcome this problem, many mines will shandy the poor quality water with just enough higher quality, read more expensive, water to allow efficient processing.

Smelting the ash from carbon used in ore processing will lead to large losses of metal values, the most efficient method is leaching the ash.

Deano


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## cuchugold

I have only seen it once, but perhaps you are familiar with a device, that I shall call "air-sluice", a series of 3 meter plexiglass pipes, of increasing diameters, with small riffles inside. It concentrated better than 20 to 1. with 98% recovery super fine, super dry dust of < 100 microns. Where water-based gravity recovery failed, vacuum or pressured air worked perfectly. I imagine it would work even better with a tight sized feed. :G


----------



## Subverted

cuchugold said:


> I have only seen it once, but perhaps you are familiar with a device, that I shall call "air-sluice", a series of 3 meter plexiglass pipes, of increasing diameters, with small riffles inside. It concentrated better than 20 to 1. with 98% recovery super fine, super dry dust of < 100 microns. Where water-based gravity recovery failed, vacuum or pressured air worked perfectly. I imagine it would work even better with a tight sized feed. :G


You would not happen to have a photo of that "air sluice" would you? Ive been doing some research on dry concentrating methods and this is one that I have not come into contact with. Very interesting concept.


----------



## cuchugold

Subverted said:


> cuchugold said:
> 
> 
> 
> I have only seen it once, but perhaps you are familiar with a device, that I shall call "air-sluice", a series of 3 meter plexiglass pipes, of increasing diameters, with small riffles inside. It concentrated better than 20 to 1. with 98% recovery super fine, super dry dust of < 100 microns. Where water-based gravity recovery failed, vacuum or pressured air worked perfectly. I imagine it would work even better with a tight sized feed. :G
> 
> 
> 
> You would not happen to have a photo of that "air sluice" would you? Ive been doing some research on dry concentrating methods and this is one that I have not come into contact with. Very interesting concept.
Click to expand...


If you keep the sizing tight, i.e. -200 mesh/+400 mesh, -400/+800, etc. You only need one diameter of pipe, and an adjustable speed vaccuum cleaner with a 55 gallon drum for the receiving chamber (or blower for production applications). At high enough speed everything gets airborne, at zero speed everything settles, the right speed is that which retains all heavies in the pipe, and zero gold flakes in the receiving chamber. A transparent pipe is almost a necessity, IMO.


----------



## cuchugold

Deano said:


> Recovery per ton of slag.
> 
> Oversize beads and strange shapes >1mm. = 1.3kg
> 
> Tabled beads < 1mm = 1.8 kg
> 
> Leached in vat, milled to 100% < 100 microns, cyanide leach = 280g, followed by CaOCl leach = 180g.
> 
> Other slags have similar grades of gold, much depends on the ore type and the processing at the mine including the smelting flux formulation.
> 
> Deano


Hi Deano: What is the rationale for doing a cyanide leach followed by a chlorine leach?. And how much gold was left after all of that?. TIA.


----------



## Deano

A cyanide leach is the simplest and easiest to run in a vat format so if you are time poor you do this first.

The hypochlorite leach is more efficient than the cyanide leach at allowing the leach access to the gold through the coatings on the gold particles. It does , however, require more supervision than the cyanide leach as the hypochlorite will react with more metals than will a cyanide leach.

Usually there is a lot of iron in the slag from the electrowinning stage of processing. This iron will not interfere with a cyanide leach but will consume highish levels of hypochlorite.

By removing what gold grades you can by using a cyanide leach you minimise the contact time required for the hypochlorite leach and thus lessen the supervision time needed.

Assaying the residual slag is difficult, it is not fire assay's finest moment.

Aqua regia digest will often (read usually depending on the flux formulation) form a gel and be unreliable in the values.

The best method I have found is to cover the slag sample with 10% HCl and let it sit for several months. Keep topping the liquid level up with water.

After leaving the brew to stew for as long as your impatience will let you, you can filter off the residual acid solution and do any of the standard assays.

This generally shows a gold level in the residual slag of 100 - 400 ppm.

It all comes down top how much slag you can get hold of and how long you can leave it in the HCl soak as to what level of recovery you get.

You will certainly recover more gold than the plant operators who just throw the slag in the mill expecting total recovery.

Deano


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## cuchugold

What is your opinion about diluting the slimes/gel with water (or HCL as you suggest) 1:4 or more and using a vibrating thickener (basically a vibrating cone-shaped VAT) as a concentrator for these superfines?.


----------



## Deano

The role of the HCl is to provide a chemical removal of metals from the gold particles to then allow gold solvents such as cyanide and aqua regia to access the gold.

It is not there as a diluent.

The size of the gold particles left in the slag after gravity separation on a table is effectively all less than 10 microns, good luck in getting any form of concentrator to efficiently operate on them.

As a special bonus the gold particles at this sizing have a strong bonding effect with the slag, they are not floating around as free particles.

You thus have a lot of slag particles with ultra fine gold attached, this is a leaching proposition not a gravity separation proposition.

Deano


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## cuchugold

This explains a lot. Thanks Deano. So, for example one could have a 100 gr per Ton (about 3 Oz) gold in ore, and if the gold is in minus 10 microns sizes, it'd be both difficult to detect properly in classical fire assays, and even more difficult to recover, especially if a lot of clay is present. Correct?.


----------



## Deano

Do not confuse the gold residual in slag with similarly sized gold particles in an ore.

The gold in slag was formed in an environment of heavy base metal contamination during the smelt and was thus contaminated with metal coating.

The gold in an ore was formed in a much less metallic environment, even ores with high metal levels do not have the concentration range of metals in slag.

It is obviously possible to have an ore which is full of base metals which will greatly affect the fire assay but over the centuries methods have been devised to modify these assays for better gold recoveries.

Unfortunately there is now financial pressure on assay labs so that to do any assay apart from a standard fire assay is very expensive and the knowledge base is shrinking rapidly.

Carrying out these assay methods on slag will not get you much return, the damage has been done during the smelt.

Recovery of less than 10 micron sized gold particles is purely a leaching proposition.

You will always recover some gold particles by gravity no matter what method you use, what becomes important is what percentage of gold you have recovered.

You do not find any of the gravity device sellers offering a bulk cyanide leach on your gravity tailings so that you can get precise numbers on your percentage recovery.

Deano


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## cuchugold

Once again, thanks for your valuable comments Deano. :G


----------



## Deano

In a totally different topic which may have relevance with the northern hemisphere summer approaching, here is an experiment which someone may wish to try.

Get access to a hedge or similar of English lavender, what you are wanting is a location where you can put a pot plant and have the pot plant surrounded with the lavender, it is important that the pot plant is immediately adjacent to the lavender, not one or two metres away.

Plant strawberries in the pot plant, the varieties appear to not matter.

When the strawberries are ripe they will have a pleasant subtle taste of lavender.

I presume that the taste comes from scent molecules transferring from the lavender to the berries, not from cross pollination by the many bees present.

Deano


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## Deano

An area which people who are performing electrowinning are often not familiar with is that DC power sources can give an incorrect readout under some conditions. These conditions usually, but not always, involve the presence of chelates in the electrowinning liquor.

The digital readouts can be up to 2v below true voltage across the plates, around 0.5 volts is common.

The only way to ensure that the wanted voltage is actually applied is to use a multimeter as a check source.

Note that when the voltage is reading incorrectly the amperage will also read incorrectly, you have to actually put the multimeter inline to get the true amperage value.

Nothing worse when thinking you are applying 1volt to the plates and you see massive bubbling from the plates, this indicates water splitting on a large scale and is not going to happen at a true 1 volt output.

Deano


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## 4metals

Thank you Deano for this observation.

Would cyanide be among the chelants demonstrating this effect?

Seeing as electrical circuitry is not my strong suit please allow me to ask for clarification. This is a typical sketch off the internet of an electroplating circuit. Electrowinning may exhibit more anodes and cathodes but the principle is the same. 



Where would you place the multimeter on this sketch for the check?


----------



## Deano

There are no hard and fast rules regarding the types of chelates which show this effect, you actually have to check each solution type in use as the effect also has variability according to the metals present.

Most of the commercial type EW power sources are designed to avoid this effect giving false readings but some of the cheaper units are susceptible.

Smaller lab type units are much more susceptible, even more so if EW direct from a pulp is practiced.

In the sketch shown you just have to put the multimeter probes across the anode and cathode bus bars and compare the reading with that shown on the power source readout.

Deano


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## Deano

One of the more useful bits of lab equipment is a vacuum filter.

The one I use was put together by a commercial engineering workshop but anyone who has access to a lathe or is willing to part this out can put together a unit similar to this.

Mine was made from a length of 200mm diameter 316 stainless steel pipe and the flange was made from a piece of 20mm thick 316 ss plate.

These dimensions were chosen because the bulk of my filtrations use a 250mm diameter filter paper.

Smaller diameter pipe can be directly scaled for smaller diameter filter papers however filters larger than 200mm pipe will require a centre support for the stainless punched plate.

I use this filter for all filtrations except where active aqua regia or hot slow filtering hydrochloric acid liquors are involved.

10 years use has shown little corrosion however I do take care to rinse out the base with water after using chloride liquors.

The main thing about this unit is the ease of use, you can change filter papers in seconds and the upper section which contacts the pulp can be simply rinsed off under a tap.

There are many designs which can be used to locate the upper section centrally over the bottom section, I prefer the 4 pillar system but others have used angled outer locaters with success.

If anyone is interested I will put the dimensions in another post.




Side view of the unit showing the vacuum system from the filter through a vacuum flask and a vacuum pump.




General side view




Closer side view showing the slope of the bottom section.




Top view of the bottom section showing the o ring and circle of ss punched plate 




View of the base with the ss punched plate removed


Deano


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## Deano

I have no idea why the photos of the vacuum filter appear inverted, however if you click on one it will appear the correct way up.

Deano


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## g_axelsson

Deano said:


> I have no idea why the photos of the vacuum filter appear inverted, however if you click on one it will appear the correct way up.
> 
> Deano


That's because you live in Australia. It's a well known fact that it's upside down. After all, it's called Down under. :mrgreen: 

Göran


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## 4metals

Nice design,

A few questions.

The filter paper is sealed by sandwiching it between two O rings rather than having suction pull it to the plate? If that is the case it allows for more filtration area because the smooth area on a Buchner funnel is where the paper seals but makes the effective diameter less. Your holes for filtration extend all the way to the edge of the inside diameter.

One thing I like about this design is the filter paper is actually clamped in place which prevents solids from sneaking under the paper to porcelain seal on a typical Buchner, which happens all too often. The down side is it takes longer to change a paper because it needs to be disassembled and re-assembled. A small price to pay for only having to filter once!

Where is the outlet for suction, I don't see it on the pictures? Do you actually draw from the bottom of the slope so all the filtered liquid is sucked into the flask? Or is the flask a trap to prevent liquid from entering the vacuum pump? 

I would think off the shelf PVC flanges like these,  https://www.usplastic.com/catalog/item.aspx?itemid=26315
 would also work. They have an O ring groove as well and with PVC construction they would be good to filter acids. A tabletop Buchner is pricey, even the plastic ones, this seems an affordable alternate. 

Thanks for posting this.


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## Shark

I was thinking the same as 4metals, before I read his post. That funnel is genius and I want one. I have some 4 inch pvc and some 10 inch, just nothing in between. I will be shopping tomorrow though. I would appreciate all the info I can get on it.


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## Deano

There is only one o ring which is permanently set in the groove in the base.

The filter paper is clamped between the bottom of the top section steel edge and the o ring in the base.

The clamping is purely by gravity so the disassembly is quick and easy.

You would need a larger piece of plastic pipe on the top section to provide the weight for clamping.

The weight of the top section is 5 kg.

Keep in mind that you will need a support for the filter paper, my rationale for using stainless steel throughout was that if I was using a piece of punched ss sheet for the paper support I might as well use ss throughout.

I have pushed a lot of samples through the unit and there is very little evidence of corrosion anywhere.

Filter paper changes take around 5 seconds, certainly no longer than changing a Buchner funnel.

The suction outlet is at the bottom of the slope, the flask is to prevent liquids entering the vacuum pump.

I use 250mm papers so the o ring diameter is 210mm, this allows around 15mm of paper outside the outer edge of the o ring for easy gripping for removal.

If you want I will detail the dimensions, this is the last of a line of prototypes.

Deano


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## 4metals

I have a very old tabletop Buchner funnel a client gave me years ago. It is heavy porcelain with a 32 cm plate and the mass of the thick porcelain ring seals the paper the same way. It has proven useful over the years for small leach test lots because the top of the old beast holds about 5 gallons of liquid. The guy's logic for giving it away was he didn't want to buy 32 cm papers just for that funnel. Worked for me!



> The suction outlet is at the bottom of the slope, the flask is to prevent liquids entering the vacuum pump.



So all of the liquid exits the funnel and is captured in the flask, making the flask both the receiver and a trap before the vacuum pump.

If you take apart one of the larger Bel Art tabletop Buchners, the big plate is supported by a bunch of short lengths of plastic pipe, the ends are cut so the liquid flows above and beneath them trapping as little of the filtered liquid as possible, and they fill the entire void to support the plate. A 50 cm funnel probably has 50 or 60 of these things filling the bottom. The bottom is flat though so they are all the same length. Every once in a while you have to remove the plate and rinse it out or just use a lot of rinse water without the paper in place. The point is the supports do hold up the plate which is more flexible than Deano's stainless plate so a larger diameter is do-able.

One consideration for members designing their own funnel is to check availability of the filter papers you prefer and make the ID of the top smaller than the paper and the base big enough so the paper fits inside the bolts. Then you will have an easy time changing paper. Nobody likes having papers too big for a funnel and having to break out the scissors! How do I know that???


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## Quwannar

Dear Mr Deano,
Before İ would like to ask a question about its reaction vessel at the neutral pH of saline /hypochlorite leaching. İ inform you that İ have been carefully reading all your posts for a long time. İf there is a god of leaching, you, Mr. Deano, must be prophet of leaching. Great Respects. 
At neutral pH of saline/ hypochlorite leaching at atmospheric/ambient pressure conditions,
Does hypochlorous acid (HCOl) release the oxigen atom into solution at atmospheric conditions? İf yes, Must a reaction vessel be 
something like a pressurized tank? 
İ got a flotation concentrate of cu pb and zn which contains 65 ppm gold. 
Best regards
Erdem


----------



## Deano

There are two things which are important regarding neutral pH saline/ hypochlorite leaching processes.

Firstly the pH must be kept in the range 6 to 8 to ensure that the active oxidant group is the HOCl radical. 

If the pH is greater than 8 the majority of the oxidant is present as the OCl group , this has a much lower oxidation power than the HOCl group and will have difficulty keeping gold in solution.

If the pH is lower than 6 then the majority of the oxidant will be present as Cl gas, this group has high oxidation power but will want to evolve from the solution as chlorine gas. This means that you are going to be working with toxic fumes, not recommended. Also at the lower pH levels there will be large losses of the chlorine to the atmosphere, this will raise the cost of processing.

In the preferred pH range of 6 to 8 there is very little loss of chlorine gas and the smell test shows a similar chlorine level to that of a swimming pool. It does, however mean that good pH control is needed.

HOWEVER, if you are treating a float concentrate you will have two effects which are not good.

The high levels of metals present will consume a lot of the oxidant and will also try very hard to drop the solution pH below 6. This means that you will need to monitor the solution energy with a redox probe to ensure enough oxidant is present not just to leach all the base metals and gold but to keep the leached metals in solution.

If the redox or Eh level drops then gold will be lost from the leach solution.

This type of feed will very rapidly drop the solution pH so very close monitoring of pH is also required.

If I was processing this type of concentrates I would look at a cyanide leach at pH 11.

If permitting for such a leach type is not possible you have four alternatives.

Firstly sell the concentrates to a smelter or other type of commercial refiner or a mine which will toll treat them.

Secondly use a ferrocyanide leach at pH 11. If used under UV light it forms a cyanide leach but will get you around cyanide permitting regulations.

Thirdly carry out an acid leach of 10% HCl to remove the base metals before using any off the above leaches.

Keep in mind that a HCl leach will need to be carried out in all plastic equipment. The leach is slow, think weeks at ambient temperatures, intermittent agitation is needed. It can be sped up if heat is used, solar works well if the leach only contacts black poly solar tubing.

You will need to responsibly dispose off the metal laden solution, the easiest way is to use up the acid in leaching the metals, decant and/or filter the liquid off and then raise the pH of the liquor to 7. Let the metal hydroxides settle and decant and/or filter the liquid off again, let the solids dry and put them into appropriate storage.

Fourthly the gold grade of your concentrates will allow you to transport them to a location where cyaniding is permitted, this may be in a neighbouring country.

Deano


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## Quwannar

Dear Deano,
Thank you very much for your valuable information in detailed. At your previous posts in this thread you described that it could run the saline/hypochlorite leach in a open Vat thats is attached to carbon column. As a hard obligation, and as long as pH by adjusting HCL must be kept between 6 and 8.and kept ORP above 1000 mV. The most active ingredient of chlorine compounds will be HOCl So running under this conditions with a close monitoring Can mentioned leaching especially in open Vat be easily carried out? 
But in the most literature İ read, they say that HOCl in aqeous solution at atmospheric conditions (contact with air and sunshine) turn into HCL rapidly by loosing oxigen molecule. 
As to concentrates i have 70 tonnes Pb, Zn and Cu. Chalcopyrite contains just gold with 65 ppm İf İ separate chalcopyrite via re-flotation. 280 kg of chalcopyrite will contain about 4 kg gold.... 
İf you were me under this circumtances what would you do Sir?
Any comments would be appreciated. 
Erdem


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## Deano

Most papers are not written by people who have to apply a practical aspect to a process, they usually are for the maintaining of grants or CV padding. The old publish or perish effect.

That said, you are correct in that there will be some loss of HOCl but this is minor in relation to the amount used in dissolving metals.

It is not absolutely necessary to have the mv above 1000, gold will be solubilised in a chloride matrix at Eh generally above 700mv. The important part is that the gold will not remain solubilised at this Eh so you end up in a cycle of gold getting dissolved, being precipitated from the solution and then getting redissolved.

This means that the gold level in solution will not, under these lower Eh conditions, continue to rise with time so that it will finally represent all of the available gold.

However if you provide the dissolved gold with somewhere to preferentially adsorb onto so that this adsorbent acts as a gold sink, then gold will continue to dissolve to replace the gold lost from the solution to the gold sink.

In an oxidising chloride matrix the best and cheapest gold adsorbent is activated carbon.

Regarding your concentrates, I am presuming that the numbers you give me are from actual assays and not just theoretical projections.

Your treatment options really are governed by two factors.

What ore reserves do you have and what treatment rate can you sustain.

If you have several years worth of reserves at a treatment rate which gives you 70 tons of float cons per 1 to 3 months then the obvious approach is to sell the float cons to a smelter.

There are many smelters which will pay for all the metals involved in your cons provided that you can offer them regular supply of product.

You will not get full LME price for these metals but you also do not have the cost of refining the cons.

I do not know what refiners/smelters there are near you but it would pay to approach some of the South Korean companies if you cannot come to an arrangement locally.

When talking to any company regarding your cons it is always advisable to be flexible regarding the product/s you can offer them.

Some companies will apply small penalties for lead contamination so a refloat to get a cleaner copper product could be very viable, especially as on your numbers most if not all of the gold reports to the chalco fraction.

It is always advisable to talk to several companies to see who will offer you the type of deal that financially best suits you.

I have yet to see a case where it gives an operator a better return to treat the float cons than to have them treated by a specialist smelter.

The only exception is when the ore reserves are small or the treatment rate is low.

Deano


----------



## Deano

I was recently doing a drop of gold from a high grade liquor containing very high levels of copper along with high levels of iron and other base metals.

The aqua regia digest was carried out with a finish of boiling so that the fumes emitted were pure white HCl.

The Eh of the filtered solution was 680 mv before addition of metabisulfite and 390 mv when the reaction had been carried out to completion.

The accompanying photos demonstrate the effect of re-dissolution of gold by base metals in a high chloride solution where the Eh is not generally thought to be high enough to dissolve gold.

The rate and level of the re-dissolution are enhanced by the ultra fine sizing of the precipitated gold as well as the heat of the reduction reaction.

The photos start with the addition of metabisulfite sprinkled onto the surface of the liquor, a gold precipitate forms , this precipitate then re-dissolves.

I also added a solid lump of metabisulfite to the beaker, this lump sank to the bottom of the beaker and a bubble trail of precipitated gold can be seen rising upwards until it is redissolved. These photos are in the second batch.


----------



## Deano

Continuation of photos from above post.










Deano


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## samuel-a

Seems like Copper (1) is formed and redissolved rather than metallic gold.


----------



## anachronism

samuel-a said:


> Seems like Copper (1) is formed and redissolved rather than metallic gold.



If that's the case what is the copper redissolved by because it wouldn't be the HCl- and why would SMB drop copper in preference to gold? 

I've seen base metal heavy solutions redissolve gold and racked my brains as to why when I know for a fact there's no Nitric around. Hoke alludes to this process in her stock pot chapter too when she mentions coming back the next day and finding your gold redissolved.


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## Quwannar

Dear Deano; 
the accompanying photos above your post is a marvelous demonstrative examples of re-dissolution of gold by base metals in high chloride liqour. Even if Eh is below 400 mV in Aqua Regia you told us that this phonemenon can occur. what is electrochemically mechanism of that. is that replacement reactions depending on metal reactivity series? how can we hinder this phonemenon? 
Could you advice us the cementing procedures with copper bus bar in pregnant solution of saline/hypochlorite which contain high lead, copper and zinc sulfides?
İ am working on copper lead and zinc of bulk flotation concentrates bearing high gold content. My aim is to leach mentioned conc with saline/hypochlorite in a small vat to get the gold. İ observed this phonemenon on Aqua Regia of flotation conc. with SnCl2 test before. in fact, you warned me, due to high base metal content in Aqua Regia of flotation conc.'s., that re-dissolution can take place by other base metal. 
also how can İ set up vat leaching unite for saline/hypochlorite leaching? 
any valuable comments would be appreciated.
erdem


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## Deano

Many people look at a Pourbaix diagram and think that the lines thereon represent a hard cut line where a reaction will suddenly change from one form to an other depending on the change in conditions on each side of the line.

This approach is incorrect, the lines represent equilibrium phases where the reactions represented each side of the line are more likely to occur.

This does not mean that these are the sole reactions which can and will occur, just that these are the predominant reactions.

The further you move the conditions from the lines the greater the likelihood that the represented conditions will be the dominant ones but there will still be some of the alternative conditions present.

Think of it as at the lines the species on each side of the line are in equilibrium.

As you move away from the line so there will be more of the dominant species for that side of the line and less of the species from the other side of the line. You do not, however, get to a point where there will be only one species present to the total exclusion of the other side of the line species.

The other side of the line species may only be present in very minute amounts but it will still be present.

In the case of gold chloride the line for equilibrium reaction of solubilising of gold is 0.994 volts for gold 3 chloride.

Above 1.4 volts you need massive levels of agitation to prevent the formation of gold oxide layers which inhibit the formation of gold chloride.

As you move the voltage down from 0.994 volts towards zero volts so you will change the ratio of gold metal to gold chloride, this means that you will have less gold in solution and more gold in suspension.

It is only when you reach zero volts that all of the gold is in suspension with no gold in solution.

The question is how do you make the above work for you in a gold based process.

In the following it is assumed that all conditions off temperature, agitation, Eh and chloride levels remain unchanged.

If you can remove the dissolved gold from the liquor then some of the metallic gold will dissolve to maintain the ratio of dissolved to undissolved gold.

The best easy way to remove the dissolved gold is by addition of activated carbon, gold chloride is a very fast loader onto this material.

Note that a lot of the gold chloride will reduce back to gold metal on the carbon and will thus be available for re-leaching in the equilibrium reaction.

If you are using a canister of carbon as an adsorbent for gold chloride, it is necessary to assay both the inflow and outflow line gold levels to see when the adsorption rate is matched by the desorption rate. This is the stage where you send the carbon for stripping and replace it with fresh carbon so that the process of gold recovery can continue.

If the dissolved gold tenor is high enough then the dissolved gold can be removed from the solution by electrowinning, usually enough over voltage is applied that the electrowon gold is cathodically protected from redissolving.

There are few gold processes now where the dissolved gold is not continuously removed from the leach liquor during the processing, zincing is the obvious candidate.

There are ways of increasing the ratio of dissolved gold in a process.

By raising the temperature you increase the energy available to dissolve the gold metal, a rule of thumb is that the reaction rate doubles for every 10C increase in temperature.

By increasing the chloride level in the solution you increase both the stability of the gold chloride complexes and the ease of forming them, both effects will increase the level of dissolved gold.

Gold chloride complexes are notorious for adsorbing onto ore particles, any increase in temperature of the pulp will desorb some of this adsorbed gold.

Having high levels of especially divalent metals in solution will increase the solution energy available to dissolve gold, these metals will be oxidised to the higher valence state by dissolved oxygen in the liquor and then reduced to the lower valence state when they react with particulate gold.

However by far the greatest effect of increasing the dissolved gold ratio is by having the gold in a form where the surface area is greatest.

This is best achieved by having the gold as a precipitate from a solution, dilute is best but even from a concentrated solution the gold particle size is still such that the surface area is high.

If you are leaching gold from a float concentrate in a chloride matrix you are making things really difficult for yourself.

The float con will have high levels of divalent base metals which will consume high levels of hypochlorite.

You will have the effect of contact passivation which will require all of the base metals contacting gold to be dissolved before the gold can dissolve.

It becomes a very expensive exercise but will certainly enrich the suppliers of hypochlorite.

My advice is to sell the cons to some one who will treat them in large scale, this may be a specialist treater or a larger mine who already has a treatment circuit in operation. You may a be able to find someone who is sending their cons to a smelter nearby or overseas and have your cons shipped with theirs.

The aim of treatment is to maximise profit, you will find that in general the cost of you leaching the cons is greater than than the costs of a third party treatment.

Deano


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## Quwannar

Deano said:


> A cyanide leach is the simplest and easiest to run in a vat format so if you are time poor you do this first.
> 
> The hypochlorite leach is more efficient than the cyanide leach at allowing the leach access to the gold through the coatings on the gold particles. It does , however, require more supervision than the cyanide leach as the hypochlorite will react with more metals than will a cyanide leach.
> 
> Usually there is a lot of iron in the slag from the electrowinning stage of processing. This iron will not interfere with a cyanide leach but will consume highish levels of hypochlorite.
> 
> By removing what gold grades you can by using a cyanide leach you minimise the contact time required for the hypochlorite leach and thus lessen the supervision time needed.
> 
> Assaying the residual slag is difficult, it is not fire assay's finest moment.
> 
> Aqua regia digest will often (read usually depending on the flux formulation) form a gel and be unreliable in the values.
> 
> The best method I have found is to cover the slag sample with 10% HCl and let it sit for several months. Keep topping the liquid level up with water.
> 
> After leaving the brew to stew for as long as your impatience will let you, you can filter off the residual acid solution and do any of the standard assays.
> 
> This generally shows a gold level in the residual slag of 100 - 400 ppm.
> 
> It all comes down top how much slag you can get hold of and how long you can leave it in the HCl soak as to what level of recovery you get.
> 
> You will certainly recover more gold than the plant operators who just throw the slag in the mill expecting total recovery.
> 
> Deano


Dear Sir Deano;
İn your method you developed "covering the slag 10 % HCl and let it sit for several months" can we get reaction time shorter for shaking table concentrate of slags by using ferric chloride.?
Due to highly strong metallic bond with iron you tought us that there will be especially difficulty in assaying and recovering the gold
from slag. 
Any comment would be appreciated.
Erdem


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## Deano

The only way to shorten the slag dissolution time in HCl is to apply heat to the process. Usually the cost of the heat is relatively large compared with the value of the time saved so heating is used only when the heat is free or near free.

Think heat sources where heat is vented to the atmosphere such as boilers etc.

Heating the process will require you to keep a closer watch on the liquid levels and will also require a more robust form of venting of the fumes from the process.

There are generally three sizes of gold particles contained in slag.

There are the coarse blobs and pellets which are released from the slag by fine jaw crushing and screening.

There are the fine gold particles, some of which are released from the slag by jaw crushing.

Both of these liberated fractions of gold will report to the concentrates when tabled carefully.

If the tabling is really done carefully or a re-run of the cons is performed, then the final con should be clean enough that a fluxless smelt can be performed.

If flux is needed to bring the gold particles together then the outside of the gold particles is dirty with base metals.

You will get gold losses into the slag from such a smelt, this can be avoided by performing a HCl leach on the gold particles before smelting.

The third size of gold particles in the slag are those which are not released from the slag by jaw crushing and remain scattered through the slag like plums in a pudding, albeit very small plums.

These gold pieces can only be recovered by performing a leach on them.

In order for the leach to be effective you need two things to happen.

Firstly you need the gold particles to be accessed by the leach liquor, this means dissolving the slag and contained metals.

Secondly you need the leaching conditions to be such that the dissolved slag does not precipitate out of solution and recoat the gold particles.

Always keep in mind that I am talking here about slag from a commercial gold mine containing a large range and quantity of metals and metallic complexes.

This type of slag will present you with many more headaches than a slag coming from a relatively clean source material such as electronic scrap.

The best method of dissolving both the slag and contained metals is by using a long term leach of HCl.

Other acids will dissolve only some of the slag components and contained metals such as iron.

Even aqua regia is not effective, you can prove this to yourself by running aqua regia digests on splits of milled slag. The recoveries will always be substantially less than the recoveries gotten by gravity systems.

The HCl leach is maintained until there is only fine sludge remaining.

You now have an acid chloride pulp which you wish to keep acid to prevent any of the dissolved parts precipitating from solution.

The obvious gold solvents for this situation are thiocyanate and chlorine, thiocyanate being preferred for ease of use. Chlorine will require the use of a sealed system to avoid health concerns.

An alternative method of recovering the gold from the undissolved slag is fine milling with gold leaching in the mill.

This method is not as efficient as the HCl digest method but is substantially more effective than just performing a leach on unmilled or undigested slag.

If the mill is fully plastic lined or constructed, then the most efficient leach system is neutral pH hypochlorite/saline.

The hypochlorite will digest the metals adhering onto the gold particles and then dissolve the gold particles themselves.

Unfortunately the slag, under these conditions, will go into a rolling circuit of dissolving and precipitating so that in order to keep the gold clean it is necessary to run an overload of ceramic balls in the mill.

If the mill is not fully plastic wetted then it is possible to use cyanide to dissolve the gold but a fully graded overload of steel balls is required.

In both of the above mill examples the pulp after leaching will need filtering before zincing or carbon contact for cyanide, or carbon contact or dechlorination and zincing for the hypochlorite.

Deano


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## Quwannar

Dear Sir Deano;
I have a 500 kg of ancient silver mine slag
sample in which densely contain gray color of metallic prills. Thats gray colured prills is so fragile that when I crush them in my impact crusher they can easily scatter into fine particle. 
when I screneed them below 2 mm in kitchen sieve and than I tabled 40 kg of sampşe at my shaking table in which I built. I got 750 grams of concentrates. They can be easily seen the semi-yellowish phase zone in my table while tabling is under way. Here is a scene of phase zone in which magnified 100 times. When I looked at them, all of particles in concentration zone was stanied with shining gold. They can easily be seen gold particles adhered to black color of fayalite parts. As you geniously proposed, I mean that yes Sir' slags greatly responds easy gravity separation. 
After the a100 mesh sieve of fine milling I will do 10% HCL leaching with heating on 750 grams of concentrates. What do you think about adding H2O2 to speed up reaction in order destroy fayelite parts. Can it directly be run the saline/hypochlorite leach at neutral conditions and or thiocyanate leach with sulfuric acid You proposed. 
Thank you very much indeed again. 
Erdem


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## Deano

A lot depends on your definition of "ancient" slag.

Old european silver slags typically have low or no borax levels so you need to see if the slag sample you have will dissolve in HCl.

The HCl dissolution of borax type slags is only successful if run as a long term leach, trying to speed up the process by adding a strong oxidant usually just starts the cycle of dissolution/precipitation which is what you do not want.

I suggest that you try a HCl leach at room temperature and just let it run for several weeks at room temperature to see if dissolution does occur.

If it does dissolve the slag then you will have to run a chlorine leach, around pH3 with around 20% sodium chloride added to solubilise the high silver level and allow the leach access to the gold.

Always keep in mind that much of the gold will be suspended in the slag dissolving leach liquor so you cannot filter and then come in with an alternate gold leach onto the solids, much of your gold will be lost in the filtrate.

Also remember that thiocyanate is not a leach for silver, the silver thiocyanate formed is insoluble and will block access to the gold.

If you try a neutral pH hypochlorite leach the dissolved slag will precipitate out and cause you much grief.

Deano


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## Quwannar

Dear Sir Deano;
I live in Blacksea Region of Anatolia. The other name is Minor Asia. There are many historic mining place in our mountains at the Black Sea region. there are lots of tonnes of slag here... Our mining history goes back Rome and Byzantium era and maybe early ages. Actually I dont think so that borax in smelting of ores was used as flux in slag in that historic period. 
Most of slags seems to be ancient.
Best regards
Erdem


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## Deano

If the slag is soluble in HCl and has no borax present then I suggest that you do a milling and gravity separation step followed by the HCl dissolution step on the table tails.

I would expect that the bulk of the silver/gold will be in the gravity cons but there will be substantial levels of precious metals in the table tails.

The cons are best treated with a cyanide leach but the tails must be treated with the HCl leach.

Deano


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## Deano

One area of recovery which is not widely recognised is that activated carbon will extract metals from fume cupboard or similar extraction air flows.

Any form of processing where metal containing solutions are heated will produce vapours containing those metals.

The temperatures do not have to be at, or even near, the boiling temperature of the liquid phase.

Apart from the obvious application of precious metals, this method will also recover base metals from heated solutions.

Particularly in the case of copper recovery, the copper can be recovered from the carbon by running a cold cyanide wash followed by electrowinning from the eluate.

Deano


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## Quwannar

Dear Sir Deano;
First of all I submit my great respect to you. I am on the right way on recoverying gold and silver from the old slags in which has no borax. I feel that you are pointing us a new perspective of a some kind of revolution on processing of slag in which has high gold content. 
An attached picture below is the wet assay results (Aqua Regia digest and AAS readings) of silver slags and tabling concentrate and tailings readings respectively without HCL leaching.
According your evaluations and your instructions, If a slag is run for without HCL leaching, There will always be inconsistent readings due to occurring the slica gel phase in aqua regia digesting and not accessing of aqua regia solution into slag metallic pool crust in which save gold. Gold value in every reading is fluctuating.
So I built a small leaching column with a small peristaltic pump that is enabling the mixing. perhaps it may be prototype of a hypochlrite and cyanide leaching column. Meanwhile, I am running for 10 % HCL solution of leaching with a heat source. Temperature is about 35 C. I will wait for several weeks until slags dissolve in sludge form. and then I am gonna send them in private laboratory for assaying. let's see that how will the assay reading change? 
Silver content is so high in readings. So Can I recover the silver and gold with a saline/hypochlorite leaching at pH 3.
Best regards
Erdem
View attachment 1


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## Deano

When you dissolve the table tailings slag in HCl there will be a lot of the precious metal values suspended in the leach liquor.

If you filter the sludge from the liquor after your HCl leach, the solids will be OK to run aqua regia digest on.

The filtrate will contain particulate precious metals in suspension, if you try to dry out this liquor you will recoat the metal particles with problem metals and complexes.

The filtrate must be leached untouched, no drying.

If the silver levels in the filtrate are low then a thiocyanate leach can be used, the silver thiocyanate complex is insoluble but the gold values will dissolve.

If the silver levels in the filtrate are high then adding salt and leaching with hypochlorite at pH 3 is preferred. Have good ventilation, low levels of free chlorine are not the same as no levels.

Always work with known weights of material so that you can calculate recoveries. You show results demonstrating concentrating of gold values in concentrates but without weights of the three component phases, heads, cons and tails, you do not know what actual grades referred to the head are.

Deano


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## Quwannar

Dear Deano;
Thank you very much again. HCl leaching is run for seven day at 37 C°.All silver as free fine particle in filtrate is floating inside liqour. When I stop the mixing, some fine particles is also floating at surface of liquor. HCl leaching is still under way by topping 10 % HCL solution to keep volumetric level. What is expreminental solid-acid liquid ratio in HCL leaching. If I take a sample from liquor and solids precipiant, Can AAS analysis be done for silver and gold either liqour without hypochlorite leaching or with running AR digesting of solids sludge. My aim is to determine the effectivetness of HCL leaching dissolution rate. 
I am trying to set up closed hypochlorite leaching vessel against chlorine gas escape made by UPVC pipe. To simultanously monitor both pH and Eh parameters in leach solution, I am planning to put a mini plastic valve under fume hood
to rubber pipe of peristaltic pump to take sample from leach solution. What should be the best design of a hypochlorite leaching vessel at pH of 3 to monitor reaction parameter.? When leaching goes on with a peristaltic pump, with the help of another low capacity of peristaltic pump at the same time, Can the gold and silver be loaded and or adsorbed onto active carbon surface in its column?
Any consideration would be appreciated. 
Best regards erdem.


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## Deano

HCl leaches are usually run at 10 -20% solids.

There will usually be enough base metals dissolved inn the leach to dissolve some of the precious metals, but the bulk of the precious metals will not be dissolved.

Usually I just carefully pan the slag each week or so and estimate the amount of precious metals in the pan concentrate, all parts of the panning method are returned to the leaching container for further leaching.

AAS requires a digest to be performed on the HCl leach tails, the real problem with doing this is that the sample is now not available for further HCl leaching.

Do your aqua regia digest after you have decided that the HCl leach has stopped showing any more precious metal values. This will then give you the level of precious metals available by using this method.

Always keep in mind that there will be large variations in precious metal values between slag samples, you may need to run several samples and average the values to get a reasonable average value for a large heap.

pH 3 leaching is usually done in a polypropylene vessel of square section. Agitation is by a down thrust polypropylene impellor belt driven from a remote motor. Sealed bearings are used throughout.

Dip samples can be taken for measurements or probes can be lowered into the pulp when required for readings.

The leach will generally ruin any probes left continuously in the pulp. 

Pulp density of the pH 3 leach varies from 20m - 50% solids.

The higher the % solids the faster the changes in pH/Eh of the leach and the more checks and adjustments required.

Gold and silver chlorides will load very quickly and completely onto carbon, unfortunately so will base metal chlorides.

A frequent check of precious metal tenors in the carbon column tailings stream is required to assure that the carbon has not been loaded with base metals.

At lab scale you cannot ash the carbon and then digest the ash residue as there will be heavy losses of the silver chloride and commitant losses of gold.

At full scale you would strip the precious metals from the carbon with a standard cyanide elution.

Deano


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## Quwannar

Dear Sir Deano;
Thank you very much indeed again. 
While 10 % HCL leaching on slags is going on, all fine gray colour metal part of slags is disolving day by day. When I panned them, lots of free metal particle can easily be seen at the concentration zone of pan. Color of filtrate in beaker are between pale grayish and greenish color. 
İt means that base and precious metal is transfering into filtrate liqour as an chloride salt form of metals. At this moment, Can I send a sample of filtrate liqour for AAS reading? Otherwise Should I wait for dissolving all metal in remaining sludge?
Moreover, you advised us that corbon loading of hypochlorite leach solution have disadvantage due to gold and silver loss in ashing stage of process in lab and full scale. 
Well Could you please suggest us how we can precipitates the gold and silver together from hypochlorite pH 3 leach solution by means of zincing? 
What is reaction conditions of zinc precipitation of hypochlorite ph 3 leaching. 
Best regards
Erdem.


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## Deano

Always keep in mind that every mining operation will have different values and conditions.

If you are looking to treat slag from a range of mines then the only constant will be that you have stopped the HCl leach because no further dissolution of the slag is occurring.

It is only when you have finished the HCl leach stage that you start doing assays, doing these assays earlier will give you only partial values.

When you are sure that there is no more slag being dissolved you filter the pulp.

The filtrate has full strength aqua regia added 1:1 by volume and is then simmered until the fumes change from orange to white.

It is preferred that the then cooled liquor has the precious metals extracted with DIBK/1% Aliquat336 organic and run in the AAS against standards of the same matrix.

If you have a very large copper content in the slag then you may need to perform a second extraction as the high copper level can crowd out the precious metal values in the first extraction.

Alternatively you can just run the cooled solution through AAS with background correction on, this is not favoured due to enhanced precious metal readings caused by the presence of base metals, especially iron.

If you are feeling prosperous then you can get ICP analysis done on the liquor, usually very expensive for single samples.

The solids on the filter paper are weighed and then separately digested in either aqua regia or cyanide, preferably cyanide with aqua regia on the washed residue.

This tells you how much of the precious metals are present as fully liberated particles in the solids.

The values registered will tell you which parts of the processing are viable in larger scale and which leach products should be discarded without further processing.

You need to get to this stage before you start worrying how you are going to leach the precious metals as you do not yet know which of the leach products you are going to treat.

Important

Always weigh your slag sample before the HCl leach so that you can do calculations on the leach products.

DEano


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## Quwannar

Dear Sir Deano;
First of all I would like to thank you very much again. I wish you had happy new year Sir with your family and your friends. 
İn my country, I learned that both commercial and universities laboratories have no knowledge 
and skills about running of DIBK/Alıquat 336 organic exraction techniques to specify the precious metal content. If I can purchase DIBK and Aliqıat 336 , I can easily do either Aqua regia or organic extraction of filtrate and participitate with mentioned chemicals. So I will be prepared them for AAS reading of precious metal in a commercial lab. It will also cost to me cheaper. 
Could you please teach us the organic extraction techniques and its standart and/or practical protocols with DIBK/Aliquat 336 in the high metallic salt solution (filtrate) and solid parts of sample. 
How much DIBK and Aliquat 336 in wolume will We use for per volume Aqua Regia solution? 
I am sory for my poor english I have difficulty in expressing and conveying my toughts to the forum. 
Best regards 
PS; I celebrate new years of all friends in our GFR community.


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## Deano

DIBK ALIQUAT 336 SOLVENT EXTRACTION METHOD FOR GOLD Atomic Absorption Spectroscopy
Digest sample in a 500 – 600 ml beaker containing around 250ml of 1 part nitric acid to 4 parts hydrochloric acid.
The aqua regia must be freshly made, do not bottle for later use.
The beaker is heated to a simmer on a hot plate in a fume cupboard, orange brown fumes will evolve.
Continue the heating until the fumes change colour to white, give it a few more minutes and remove from hot plate to cool.
When cool the solution is filtered, usually in a Whatman No1 paper and the paper is rinsed with water from a wash bottle to get all traces of gold into the filtrate.
After filtering has finished the filtrate is made up to usually 250ml with water and poured into a clean beaker to evenly distribute the gold values in the solution.

In a clean 50ml glass extraction tube add 50ml of the final filtrate and then add 5 ml of DIBK Aliquat 336 (1%) ( straight hypochlorite solution is ok, but Aqua Regia must have Nitrates expelled as above). This is a 10:1 concentration ratio.
Shake thoroughly 100 times. Allow to settle. If there are drops of solvent adhering to the walls of the extrtaction tube the tube is sharply but gently finger flicked to dislodge those drops and have them join the organic phase.
If any haze is seen in the solvent section, or there is concern of excessive iron in the sample, add drop by drop some concentrated hydrochloric acid in a very slow and controlled manner dropwise with a 2 ml or similar dispenser. The now clear DIBK solution at the top of the extraction tube is ready to scan directly on AA.
Special note: The standards used on DIBK extraction must be made of DIBK Aliquat 1%, eg. The 0 ppm reference is straight DIBK Aliquat 1%, and the 1ppm and 5ppm need to be made up from straight DIBK Aliquat 1% solution and addition of Gold chloride standard.
Flame adjustment is only needed if maximum sensitivity is required , the flame can be of lesser intensity for DIBK than the flame used for direct aspiration of aqueous liquors.
Usually the small improvement in sensitivity when using DIBK is not worth the fuss of flame adjustment and the flame is set so that aqueous samples may be direct aspirated with out flame out and this setting is used for all liquids.
NOTE THAT THE DIBK STANDARD READING WILL DIFFER FROM A GOLD CHLORIDE OR CYANIDE READING DUE TO THE DIFFERENCE IN UPTAKE RATES FOR THE SAME STANDARDS.
**This method removes all Iron and interfering impurities from the leach to negligent levels even in saturated Iron solutions. DIBK solvent cannot be used on it’s own but needs an addition of 1% of Aliquat 336 (quartenary ammonium salts) to make it effective and have a relatively high viscosity to enable it’s use, this is called DIBK Aliquat 336 (1%) solution. It is made by having 10ml of Aliquat 336 placed in a 1lt volumetric flask and bringing the volume to exactly 1000ml with DIBK solvent. This is then the stock solution to make standards and direct use.
Oxidizing or non oxidizing solutions to be tested can be used with this method without any prior treatment but strong oxidisers eg. Straight Aqua Regia or any other nitrate containing oxidiser must have the nitrates expelled. Strong oxidizing solutions are more harsh on components in the nebuliser of the Atomic Absorbtion Spectrometer as well as being promotors of side reactions in the extractant components which can lead to false metal readings.

If the 10:1 concentration ratio gives an absorbance value which is higher than the value for a 5 ppm gold standard then a lower ratio is used.

A 1:1 ratio will just clean out interferents without enhancing the reading range.

If the solution you are testing has such a high precious metal level that a 1:1 ratio will still land you outside the 5 ppm limit, you can reverse the ratios such that the DIBK : liquor ratio is, say, 10:1 rather than the standard ratio of 1:10.

The reason you try to keep the metal readings in the DIBK below the 5ppm standard reading is that from 0 to 5 ppm the readings are straight line. This means that the absorbance reading for a 5 ppm standard will be 5 times as high as the absorbance reading for a 1ppm standard.

However, if you plot the readings above 5 ppm you will get what is called a calibration curve.

Most AAS machines have calculation circuitry in the electronics which will give you the actual value of the true absorbance reading for these higher values.

All of the above works brilliantly in a lab where clean synthetic solutions are being used, it tends to be more variable when real world gunged up solutions are being run.

The above represents the difference between an operator's idea of a bare bones AAS unit and the developer's idea. The developer's idea is that a simple unit only has relatively small amount of circuitry dedicated to the above type of calculations, the operator wants this type of circuitry kept to an absolute minimum to avoid processing errors.

The real difference is in the interpretation of the word "relatively", to the developer it seems to mean that he has restrained himself from installing as much software as he could possibly fit into the unit.

You can avoid most of the analytical problems just by keeping the DIBK/liquor ratios such that the absorbance values are maxed out around the 5 ppm standard values.

DIBK usually comes in 20 litre drums, Aliquat 336 comes in various 50 to 1000 ml containers.

The mixed DIBK/Aliquat solution is usually dispensed in 5 ml pump volumes from a dispenser mounted on a Winchester bottle.

Deano


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## Quwannar

Dear Deano,
You are a great man. Great and golden compiled knowledges for me. Thank you very much for your valuable sharings. I think that it has come the time to own an AAS equipment. Because it would be hard to tell or persuade the laboratories for a proper and a right analysis method required.
Thank you sir again.
Best Regards
Erdem


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## Deano

Many people have difficulty with the idea that any change in leach conditions will involve a trade-off.

The simplest gold leachant is cyanide using oxygen as the oxidising agent.

If you want to speed up the leach dissolution rate with physical inputs you can add heat or increase the amount of agitation applied.

Both of the above have limits above which the leach rate will not increase and both have input costs.

Basically you speed up the leach rate but it will cost you more to do so.

You are trading off time against money.

Similarly you can increase the leaching rate by adding different and stronger oxidisers rather than using the free oxygen from the air as dissolved oxygen in the leach solution.

These stronger oxidisers are usually peroxides of some form but exotic di and trivalent metals have been used.

The use of these oxidisers will increase processing costs directly but will also give problems down the line if you are zincing for gold recovery.

It will cost you more to remove the oxidiser from the leach solution to minimise redissolution of the zinced gold compared with using just dissolved oxygen from the air which is usually removed by vacuuming the liquor after filtration.

Here you are having an extra processing cost being traded off against time.

There are many cyanide based formulations available commercially. Most of these contains eclectic mixes of chelates and alternative gold solvents.

Apart from the extra cost relative to a straight cyanide leach you will have the problem of many other metals being leached by these added components.

In a zinc recovery from these leach types you will have to zinc out many of these extra dissolved metals in order to recover your gold.

In carbon processing you will have great difficulty in stripping the gold from the carbon, you are forced into ashing the carbon in order to recover the gold.

This is in addition to the loss of gold adsorption capacity on the carbon due to the loading of these extra metals.

In leaching some ores such as copper types you can minimise cyanide consumption by dropping the cyanide in leach tenor.

This makes the leach more specific for gold over copper but will slow the gold leach rate to the point where it can only be run as a heap or vat leach.

Here you are trading off cyanide cost against time and the inconvenience of having high copper levels on the carbon and thus in the carbon strip liquors even though you can minimise the copper levels in the strip electrowin liquor by performing a cold cyanide pre-strip on the carbon to remove most of the copper before attempting to strip the gold with a hot solution.

Usually when you are working with a high cyanide soluble copper ore you are trading off the lower cyanide consumption for longer leach times and more complex processing methods.

Pretty much all variations of all leaches involve tradeoffs of some form or another, the trick is in identifying the tradeoff and deciding whether you wish to go down this path.

Deano


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## Quwannar

Dear Deano;
Analysis result of 10 % HCL leach of silver slag is given below. 
Head slag of sludge of table tailing below minus 150 mesh parrticle size. 
Au 3,97 ppm
Ag 17.04 ppm
Filtrate of this table tailings 
Au 31.8 mg/lt
Ag 370 mg/lt. 
Pb 2715 mg/lt
Zn 817 mg/lt
Cu 94 mg/lt
As seen from results, the filtrate is more valuable part than solid sludge. 
Can I recover the both Au and Ag together from liquor with saline/hypochlorite leaching method at pH 3 ?
Then can I load the gold and silver complex onto activated carbon?
İt is possible to sell the activated carbon loaded by mentioned metals. 
Many thanks in advance
Best regards 
Erdem


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## Deano

You need to have weights and volumes in order to tell what values are where.

What is the dry weight of the sludge and what volume of leach liquor as the filtrate.

You give the metals grade of the filtrate in mg/litre, check with your analyst to see if these values were milligrams/litre or micrograms /litre. It is kind of important as there is a thousand fold difference between the two.

Your recovery method depends entirely on which of the above grades is present, let me know the above values and I can advise you further.

Deano


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## Quwannar

Dear Deano;
I dont actually know the weights of sludge and the volumes of filtrate used by laboratory. They said that they could be capable of doing a standart DIBK/alliquat 336 extraction test after Aqua Regia digest. I sent a 75 ml of filtrate and 90 gram of sludge from which I run for a 200 gram slag sample insid 1.000 ml of 10 % HCl. 
Units in report is ppm and mg/l. Milligram per liter
They did not state the weights and volumes. 
Best regards
Erdem


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## Deano

What we need to know is the final volume of liquor from your leach and the final weight of solids after the leach.

If you started with 1 litre of solution, how much solution was left at the end of the leach.

Similarly how much solids was left after leaching.

I presume that the lab used aqua regia digest techniques rather than fire assay on both samples, if not please report which and why.

I am not worried how much sample the lab used.

Deano


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## Quwannar

Dear Deano;
I put the 200 gram of re-milled slag minus 100 mesh in 100 ml of 10 % HCl in one liter water. I completed the volume level with 10% HCL to maintain the loss due to evaporation. İn fact. I did not measure the volume loss and weight the sludge remained. And then, I picked up the sample mentioned amounts in my previous reply and I sent to laboratory. 
So You can not do a math to help me about the yield and the chemical comsumption in the further leaching stage. I am so sorry for deadly mistake. 
I am right now repeating the same expriment in specified wolume and the weights.
Expertise in lab said that you would do aqua regia digestion for especially in sludge. But you dont know what to do with the filtrate. I remembered your instructions to them about the filtrate. Lab guys prepared 1;1 volume filtrate and aqua regia. And did the dibk and alliquat 336 extraction and AAS reading. 
Thank you very much for your valuable time and your efforts. 
Best regards
Erdem


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## Deano

You are trying to find out several things from the HCl leach.

1. What % of the slag is dissolved in the leach.

2. What % of the precious metals reports to the leach residue and the leach liquor.

3. What is the grade of the precious metals in the residue and liquor.

The above findings will show you which products of the leach process are viable to attempt to recover the metals from and the likely return from each product.

These results may affect the leach type to be employed in commercial processing and so are very important.

Usually once you have the indicative numbers from the leach test, if they appear viable, you would do multiple repeat tests in order to get an average grade of possible recovery for the whole deposit.

Always keep in mind that the numbers will vary greatly from sample to sample depending on the grade and processing of the original smelting.

Your first numbers, even though not related to known volumes and weights, appear to show the slag contains the high precious metal levels which are to be expected in the slag.

Usually three things will happen in relation to the processing of this type of material.

1. People will take the head slag and apply a standard assay analytical technique without first doing the HCl leach.

The results will be low and disappointing and you will be discouraged.

2. People will look at the results from the HCl leach products and decide that the numbers are too high and that you have a flaw in your test regime.

Unfortunately these people are often big wheels in processing and their pronouncements will carry a lot of weight and once again you will get discouraged.

3. Someone with a vested interested will propose an alternative method of processing and when it fails will claim that the grades were not there in the first place.

The best way to negate all of the above is what you are doing, namely simple repeat testing with independent analyses.

It is important that all of your testing is recorded in detail so that you have all of your numbers available for review if needed.

Finally, if you carry out the HCl leach on just the table cons by themselves, they will clean up to the level where they can be dried and smelted to give you a silver bar with low but very important gold levels.

It is important that the smelt is done without flux in a clay crucible otherwise you are going to restart the entire metals in slag process.

Usually you put the clay crucible inside a graphite crucible as a safety precaution so that when you are reusing the clay crucible you do not get too worried about any breakage of the eroded clay crucible.

Keep in mind that the metals from the table con will usually be the majority of the metals recovered so it is worth doing the processing correctly.

Deano


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## Quwannar

Dear Deano 
Our Master;
I do understand the sipirit of your statements well about the proof of precious metals entities in slag by instrumental analysis. I could not still persuade lab guys working in highly prominent labs. Everyone does not want to give up their the standart comforts. Thats their culpa maxima. 
As you brillantly propose, the coming back from the outcome has always been useful tool in the science and art of gold and its engineering.Because man do want to see the gold. 
The method you developed is a some kind of a revolution in slag processing. And your posts in GFR must also be lectured in faculties of engineering. 
Meanwhile, I smelted 40 gram of dirty concentrate containing some fayelite with some dry borax. Smelting conditions was not suitable for full needs. I saw real gold disseminated at the slag surface. 

Knowledge and wisdom be upon you sir.
Best regards
Erdem.


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## Deano

In analysing digest solutions for precious metals by AAS using DIBK / Aliquat336 as an extractant there is often a gel formed in the organic phase which prevents aspiration.

Most users of this technique have their own preferred method of breaking this gel phase so that a liquid sample is available for aspiration.

These methods usually employ an acid of some form, phosphoric and hydrochloric are among the most common.

The gel is usually formed by the presence of micro particles of silica from the digest and the best way to break the gel is by the addition of Hydrofluoric acid to the DIBK after phase separation.

I do not know of anyone who is really comfortable with using hydrofluoric acid as such, sooner or later there will be spillage or droplets onto the skin or gloves.

A much safer alternative is to add gently some ammonium bi fluoride powder to the gel, the dissolution of the silica is rapid and it is not necessary for all of the gel to be broken, only enough for aspiration to occur.

The AAS reading for gold is totally unaffected both up and down by the addition of up to 1 gram of the powder to the gel.

Deano


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## Deano

When I started in the gold game last century the recycling industry had not really started in regard to e waste.

It was a case of learn as you go with large quantities of scrap.

The first thing I learned was to use only cyanide as the gold solvent, any acid based methods were either too expensive or metallurgically tricky.

If you were using alkaline cyanide solution it made sense to remove any coating metals with alkaline solution so that you did not have to go through a neutralisation step which was costly when running large quantities of material.

I found that tin, zinc and lead could be cheaply, easily and completely solubilised using sodium hydroxide solution.

You had to play a game of time and temperature versus removal efficiency of the metals.

At 1% caustic solution you needed to have the temperature up around 80C for good metal removal.

At 20% caustic you could get good removal at room temperature.

Temperatures and caustic strength were pretty much straight line between the two extremes.

I generally went for the 20% option as there were no heating costs.

The thickness of some of the metal coatings meant that some agitation was needed, I just used a peristaltic pump to move the liquor through the material, this meant that any reactions with pump components were eliminated, you just needed an alkali compatible hose.

As there was no sorting of components in those days the time of the alkali leach was controlled by the leach time of the thickest coating, could be up to 24 hours.

All dissolved metals were removed from the leach liquor by running the liquor through a column of activated carbon and the caustic level adjusted before recycling the liquor.

If you are looking to use acid based methods for the gold dissolution step it is not hard to do an alkaline leach, rinse and then come in with the acid leach.

Deano


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## anachronism

Dean

Did you put the liquor at 20% Caustic through the carbon or did I miss your point? If I recall correctly from what you've told me an alkaline solution loads to a lower degree on carbon than an acidic one- could you please clarify? 

Jon


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## Deano

Good to see that your memory is as good as always.

Yes an alkaline solution will generally load to a lower degree than an acid equivalent solution will, but both will load to a level which meets the requirement of removing the metals from solution.

Consider the two solutions to be the equivalents of a Model T Ford and a Ferrari, both will get you to your destination without all of that tiresome walking but the Ferrari will get you there faster.

With base metals it is best if any heated solutions are cooled before treating with carbon.

Deano


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## METLMASHER

:shock: Great reading, and my heartfelt thanks to Deano and the GRF


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## Deano

Heap leaching

One of the many myths about heap leaching is that any recovery below the theoretical amount is due to the mineralisation being locked up in matrix.

Occasional mention is made of possible channelling of the liquor flow but this is usually relegated to a minor secondary position on the “why did my recovery not reach near optimal” scenario.

If the leached heap is pulled apart so that the internals are fully exposed, a totally different pattern emerges.

Large areas of the heap, average working volume 10% of the heap, are totally unleached.

In a copper heap leach the difference between the whitish leached areas and the green unleached areas is stark while the sheer volume of unleached material is stunning.

The unleached material forms large zones, many house sized, in a variety of sizes and shapes.

It is obvious that the leach liquor has not performed any leaching in these zones.

As well as these unleached zones inside the heap there are other unleached zones around the perimeter of the heap.

One of the other myths regarding heap leaching is that the leach liquor will travel substantial distances horizontally in the heap.

This is incorrect and the liquor flow appears to be controlled more by gravity than any other factors in a heap leach.

This means that there is an unleached zone of ore which starts approximately two metres in from the outer edge of the top of the heap and will include all ore from the line starting two metres in from the top edge to the outer edge of the heap.

The two obvious questions in relation to the unleached zones are;

Why are they there

And is there a way to leach them without having to remodel the heap entirely.

Unleached zones are purely an effect of using sprinklers to distribute the leach liquor onto the heap.

Apart from the sprinkler side effect of depositing gypsum on the top of the heap due to excessive evaporation and thus forming an impermeable cap which prevents liquor penetration occurring evenly across the top of the heap, the liquor flow from a sprinklered heap tends to run through the heap in defined channels.

This is the cause of the unleached zones in the heap.

The size and number of these dead zones is also dependent on the ore type, milling and placement protocols.

There is also no sideways distribution of the leach liquor in the heap so the leaching start line is vertical from the inside edge of the top bund wall.

The unleached zones within a heap may be leached by changing the liquor introduction method to the heap from sprinklers to flooded paddocks.

This alters the liquor flow patterns inside the heap so that the flow now goes through the previously dead zones.

In the paddock method the bund walls are made around 1m high by 2m wide, no compaction is employed.

This allows the liquor flow to move sideways through these walls and thus gain the two metres horizontal distance before distributing vertically.

Before the bund walls are formed the top 0.5 metre is removed from the top of the heap. This removes the gypsum layer from the sprinklers and allows better distribution of the liquor through the heap.

Depending on the flow rates through the heap and how the treating circuit is designed and sized, a paddock size is often around 100 x 50 metres.

The surface is levelled before the bunds are constructed.

A vital part of the surface conditioning is the compaction of the surface so that a 300 to 400 mm depth of water in the paddock will take around 8 hours to drain dry.

After the liquor has drained from the paddock the paddock is lain idle for an hour or so to allow air to be drawn down by the liquor to access the heap’s interior.

The paddock is then refilled and the sequence repeated.

Fill time is usually around 1 to 2 hours, scouring of the surface is prevented by introducing the pumped liquor into the paddock via a tangential pipe entrance into an old truck tyre with the liquor spilling out via the centre of the tyre.

A further major advantage of the paddock method is that the alternate cycles of liquor flow through the heap followed by the air access to the ore will greatly speed up the degradation of sulphides present in the ore particles, this applies for both acid and alkaline leaches.

This sulphide degradation will allow the leach solution access to more metal values in the ore particles.

Deano


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## snoman701

Hey deano....

When running a low concentration cyanide leach how do you monitor cyanide levels in the leach? Let’s say you have a target of maintaining 10 ppm free cyanide. 

Are you just doing a particular titration, or are the meters good to that level when calibrated against good standards?

Does the use of a ferrocyanide leach complicate monitoring since free cyanide could be low when total cyanide (free cn+ferro-cn+aucn) is otherwise appropriate? 






Sent from my iPhone using Tapatalk


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## Deano

The handheld meters have pretty well relegated titration to being used just as a regular check on the calibration of the meters.

The meters are good to extremely low levels of cyanide, the usual problems are lack of care with the probe, especially leaving the probe in solutions of free cyanide, this will affect the zero point and also allowing the probe to dry out.

All that you are interested in is the free cyanide level and that is what the meters will register.

If you are using the ferrocyanide leach in sunlight then pretty well all of the conversion to free cyanide will have occurred.

You are only interested in the free cyanide level when trying low cyanide gold preferential leaching, especially with copper present.

Deano


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## snoman701

Deano said:


> The handheld meters have pretty well relegated titration to being used just as a regular check on the calibration of the meters.
> 
> The meters are good to extremely low levels of cyanide, the usual problems are lack of care with the probe, especially leaving the probe in solutions of free cyanide, this will affect the zero point and also allowing the probe to dry out.



Is the titration against silver nitrate with potassium iodide the best method to check calibration? Or is there a better version? 

What brand meters (probe) hold up well? 


Sent from my iPhone using Tapatalk


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## Deano

Titration testing is done to check for gross errors in calibration of the meters or if it is suspected that the cyanide standard used has degraded.

You usually use the same cyanide source in calibration as used in the plant, If that source has degraded then a new check source is used.

Pretty well all sources of cyanide are reliable in regard to freshness of supply, this was the most common problem with cyanide stored in mine sites.

Most probes come out of China and are surprisingly reliable if looked after.

Buy by price, most of the meters also come from China and are similarly reliable.

Deano


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## Greenhorn

Deano, in your cyanide heap leach, you said you set up three activated charcoal filters to catch the gold in solution, how do you test to see when the first filter is saturated and know when to pull it for refining? I'm assuming there is a tool for this, or is it a simple test for the presence of gold in solution? (stannous paper?)

Also, the forum talks about testing concentration of cyanide in order to maintain safe and effect recovery, is there a tool to do this as well?

Thanks,
Greg


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## Deano

If you are running a heap leach the liquors are usually tested by AAS to see when the carbons need stripping, after a few cycles you will learn the approximate time it takes to load the first column so you really only need to start testing the liquor shortly before this time.

Usually you would also check the loadings on the second column at the same time, after a few cycles the gold missed by the first column can start to load up appreciably on the second column.

The AAS readings are usually done by direct aspiration of the liquor, no organic extraction is necessary. This testing can be done quickly and simply by any one with an AAS, a local lab should, depending on instrument availability, be able to do the test in a few minutes.

For safety you keep the pH at around 11, use a small handheld meter for spot checks.

For efficient leaching you are interested in free cyanide levels, these are checked using a cyanide meter with cyanide probe.

Both the above meters are purchased online, most are of similar quality so buy on price.

Deano


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## Greenhorn

The probe, I did find one listed as a photometer, readings in mg/l (ppm) what range of meter do I need to get, or in other words, what ppm should this solution be.

In your photo's of the heap leach, you show a guy walking under the bird netting to set/adjust the water spray hoses, these open heap leachs, if the ph is kept around 11, is there almost no cyanide released through evaporation, mostly a safety question because caution should always be kept, but on a hot day with no wind, is there a potential of injury to anyone/animal walking near these vats?

I know you mentioned that one of a couple problems with gravity separation is loosing gold to what is referred to as the slims as they are a whole different animal. For a small operator, will these slims cause problems in these smaller vats? And is there any justifiable reason to wash these slims out of the crushed ore and treat them separately (I understand it's a cost/time thing) but can it have a benefit. I ask because you had mentioned that you can have areas in the heap that will develop water returns to bottom that may bypass some parts of the ore and that maintaining the whole head ore is better then trying to concentrate it.

I want to express my own thanks to you for sharing this information and also answering questions about it. The previous owner of the mine I have turned 90 last year, he said he would love to be on site and help teach what he knows, but he's since moved to warmer/drier climate and his wife fears that travel may end his days and I'd rather that not be the case. I've reached out to my local GPAA, but they all seem to enjoy panning in the rivers, getting a couple of grams of gold over the weekend. I was told specifically by one person that if he can hear birds singing above him then he feels he's in the wrong place. I'm sure in my ventures I may find someone as helpful as you, your skills are very valuable to those like myself.

Thanks,
Greg


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## Deano

The standard cyanide level for heap leaching ranges from 50 to 500 ppm, the higher level is used for a fairly clean ore and the lower level is used where there is a high cyanide soluble copper level in the ore. So you are wanting a meter reading up to 500 ppm cyanide for direct readings but one with a lower maximum can be used if the cyanide solution is diluted with clean cyanide free water to get the readings into the meter's operating range.

Note that the meter will read free cyanide in solution, this is approximately half the ppm level of sodium cyanide itself. So if you add 1 gram of sodium cyanide to one litre of water you have made a 1,000 ppm solution of sodium cyanide but the free cyanide level will be approximately 500 ppm.

The main reason for HCN off gassing from a heap is when you use sprinklers to spray the cyanide solution over the heap, the surface area of exposed liquid surface is dramatically increased as is the resulting HCN off gassing.

If the solution is applied as a flow into a paddock there is virtually no surface area increase and so the HCN off gassing is minimised.

The above is similar to a CIP tank where the solution is pumped into the tank rather than sprayed, you do not see people on the walkways above the tanks suffering from HCN poisoning. The above is an OHS detail in which a lot of care is taken to keep safe, apart from not wanting to cause injury to operators you would void insurance cover if not operating in a safe manner.

If running a vat leach the milling is done to a coarser size than for tank leaching such as CIP. Each ore will have a minimum milling size which, if over milled, the higher production of fines will plug the heap either entirely or in zones.

Tank leaching is done to a sizing where gold leaching is finished within 24 hours whereas vat leaching can run for months.

The rule of thumb when milling for vat leaching is that if a damp sample of the milled ore is formed into a hard squeezed sausage around 30mm diameter and held in a hand with around 50mm of the sausage protruding from the thumb/1st finger circle, if the sausage slumps like you would expect from clay then the milling is too fine and the grind is coarsened until the sausage will break off rather than slumping.

A coarse screened hammer mill or rolls mill will usually provide a milling profile which is suitable for a vat leach, it is very dependent on the ore type and the milling profile of that ore as to what milling regime is used.

If there is a high level of sulfides present in the ore it may be better to run a series of jaw crushers to do the milling, this will minimise fines production as sulfides tend to slime in high impact type milling.

If you wash out the fines you will need to treat these fines in a tank leach so you will end up with two leach circuits and the associated extra costs.

If the fines have a high sulphide level then it adds extra costs to breakdown the sulfides to allow the leach to contact the gold locked inside the sulfides. It is much simpler and cheaper to mill in a manner which minimises slimes formation and use the extended vat or heap leaching time to breakdown the sulfides naturally in the vat or heap.

Deano


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## Greenhorn

Thanks again Deano, but got a couple questions for clarification.
You talk about heap leach, and I know that that is, but you also had been mentioning vat leach as well, I've been assuming your using vat as it's a smaller version of acre sized heap leaching, but your last comment made me wonder if the vat is somewhere between open heap leach and a tank.
In regards to the tank, I thought you were more for heap leaching for effectiveness, I like the idea of only leaching for 24 hours, but didn't think from what I've read on the forum that this wasn't even achievable. 
I live in Idaho, and am surrounded by dairy farms and grain silo's. So you when your talking above, I was visioning a large round container, 3+ meters in diameter, and how ever much high as needed, with a steel punch plate at the bottom covered in a filter/mess to prevent the ore from going through and only fluids can drain below that into a cone that then feeds into a recycling system like you outlined with the heap leach in an open field. 
Would this be the tank system your referring to, or is more of a vat?
And your definition of coarse material for this system, are you talking something like 6-8 mm or more like something in the 20 mesh, not sure where to start with the sagging vs breaking of the 30 mm cylinder of ore.
Is the use of ordinary steel work with cyanide? Does it react with it? Or does it need to be layered with a plastic film like you talked about in the heap leach?

The ore (that I still need to get permitting to open the shaft to even get a sample) as I understand it is gold running with iron pyrite, he gave me a 25-30 mm diameter piece of high grade that has both gold grains and the pyrite cube structures. He told me that he was only chasing the gold and discarded any other materials found. So I'm not sure of any copper or silver contents until I have an assay done. When he was operating it back in the 80's, they removed the vein material, crushed to around a 200 mesh, ran on a shaker table to remove freemill gold and then sent the rest off for processing. At the time I didn't know about running head ore vs just concentrates. I was also under the impression that they took the ore offsite to process, so only thing remaining next to the shaft is the quartz monzonite pile of waste. I really haven't look through it much, as the forest has slowly reclaiming where it's been dumped.


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## snoman701

Greenhorn...you should download a copy of Rose's Chemistry of Gold. I think it will help fill in some gaps between what you already know, and what Deano is posting. Some things have changed a lot in 100+ years, but some haven't changed much.


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## Deano

CIP tanks are made of ordinary mild steel, usually with cathodic protection to minimise rusting. Because cyanide leaching is run at an alkaline pH there is no attack on the steel by the pulp components.

If the previous operator was separating the sulfides from the rest of the ore then I would recommend that you look at doing the same.

Some sulfides will cyanide well but many will be refractory and need some other approach, there is usually a good reason why a particular option is used.

If you have a mix of free gold and gold in sulfides then you have three options for treatment.

The first is to mill and vat or heap leach, if you are in a farming area then you are not likely to get permitting for cyanide use so look at the next options.

The second option is to sell any mined ore to a nearby operator and let them have the worry of treating it. The gold grade of the ore will determine the distance it is profitable to haul the ore to an operator.

The third option is to separate the sulfides and free gold from the ore and sell these concentrates to a smelter or toll treater.

Realistically there are two methods of separation which will give you all of your gold values in the one concentrate, these are flotation and gravity concentration.

Flotation is not a process you learn from books, it is best learned hands on in an operating plant with a good operator to teach you. The upside is that you are not looking to be an all round flotation expert but to just be a competent operator on a single ore type.

The simplest gravity system is a table but you would need to run at least two size fractions of your ore in separate streams in order to get reasonable recovery of the gold values.

Sulfides are notorious for sliming when milled so you are looking for the milling method which minimises sliming.

In small scale mining you can run a series of jaw crushers with intermediate screens, this will get you down to 300 microns particle size maximum but the capacity is limited. having said that, there are some very cheap and efficient jaw crushers coming out of China which will let you set up an efficient milling circuit at a reasonable cost.

It is probably worth talking to the previous owner to see who he sent the ore to, was he sending the entire ore less any free gold or was he sending just concentrates less the free gold. this may give you a better idea of what processing is possible financially.

Deano


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## Deano

Occasionally there is a need for a bottle to be cut so that the bottom section can be used as a beaker or similar.

There are many various methods used to achieve this but the only one I know of which allows you to get a sloping cut is the oil method.

The bottle is filled with oil, usually used engine oil, to the level where the cut is to be made.

The bottle can be leaned over at any angle which does not have the oil come out the spout.

If you are wanting a square cut you just place the bottle on its base, but angled cuts are done with the bottle lying on a sloping bed of sand.

A piece of rebar which will fit through the spout of the bottle is heated to red hot either with an oxy set or in a really hot fire.

The rebar is then inserted through the bottle spout and into the oil, keep it there until the glass cuts.

The glass will cut cleanly along the oil level, if it does not cut you did not have the rebar hot enough.

Deano


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## SamW

Deano said:


> Table optimising
> 
> Many years ago I was involved with a project to maximise gold recovery from a milled ore by using a small (3') wilfley table to clean up concentrates and middlings from a bank of larger tables.
> 
> Due to environmental constraints chemical processing was not allowed on site.
> 
> A short period of operation revealed that running the table in its standard format was leading to losses of fine gold where the particle size was less than 50 micron screen size.
> 
> The gold particles were flattened from the milling and appreciable quantities were also lost from the plus 100 micron screen sizes.
> 
> Despite doing all of the standard optimisings of adjusting feed size screening, feed rate onto the table, table side tilt and water flow I could not get a major recovery improvement.
> 
> I went into the literature and found that the largest number and most useful papers were those from the British tin industry.
> 
> Basically they said that keeping a tight sizing on the feed was vital, I was already doing that so OK there.
> 
> Making sure that the table was level on the longitudinal axis was a fundamental which I was also doing.
> 
> Feed rate was best when a loose bed was set up along the table, don't put too little feed on the table nor try to put too much feed on the table.
> 
> Keeping the feed rate constant was also very important, I was feeding from a wet sump with a screw feeder so OK there also.
> 
> Side tilt was to be such that clay fraction particles were washed over the side of the table but the tilt was to be little enough that a middlings product could be readily separated at the end of the table.
> 
> Even when I had the table set up to cover all of the above parameters I was still losing fine gold.
> 
> The only area where I did not have full control was the side water coming onto the table, no matter how much I tried I could not get a perfectly even flow across the table.
> 
> I decided that I needed to improve the delivery of water to the table.
> 
> I did so by running a length of 3/4" copper pipe suspended about 4" above the side of the table where the water exited the original flow boxes.
> 
> The pipe was blanked off at the table exit end and was connected to a hose and ball valve at the feed supply end.
> 
> Ever 1" along the pipe was drilled a 1/8" hole.
> 
> When the water was turned on a curtain of water sprayed down onto the top edge of the table and delivered an even flow of water which could be easily adjusted for flow rate with the ball valve.
> 
> Even this change only partly improved the recovery, it was evident that there needed to be a difference in the water flow rate supplied to various parts of the table.
> 
> After a lot of testwork I settled on having the water delivery pipe as two pipes.
> 
> The holes remained the same size and spacing but the delivery was split into two parts.
> 
> The first part was as above but only extended 2/3 of the table.
> 
> The second part covered the last 1/3 of the table, each part was as a separate length of pipe so that adjustments could be made to either part without affecting the other part's flow rate.
> 
> In order to keep the water holes at the 1" spacing the pipe for the last 1/3 of the table was fed from the bottom of the table and the pipe ends almost touched.
> 
> Each pipe length had its own ball valve for separate flow adjustment.
> 
> The side wash water pipes were fed from an overflow overhead tank so that a constant head was maintained.
> 
> This was important on a mine site where valves were being opened and closed in other parts of the circuit, this would affect the pressure to the wash pipes.
> 
> This setup allowed recovery of free gold down to 25 microns, the disappointing part was the low weight of the 25 to 50 micron gold recovered, it looked a lot as a sheet like paint on the table but weighed far less.
> 
> On the plus side there was a substantial improvement in the plus 50 micron gold which did weigh well.
> 
> If run from a municipal water supply the overhead header tank may not be necessary depending on the vagaries of the particular supply.
> 
> 
> Deano



Deano! what a wealth of information in this thread!

Can I ask what sort of overall gold recovery rate you were achieving with the large tables? I'm currently trialing a full size Wilfley on my gold ore and yet to find someone with this specific experience. Would be great to know what sort of recovery I should expect before spending too much time on something that may not be achievable.


Cheers
Sam


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## Deano

Gold recovery on tables is dependent on the size of the gold put onto the table, the shape of the gold particles, where in the ore the gold is located and how the table is operated.
Many ores have some or all of the gold associated with sulfides, you are then carrying out a sulphide recovery rather than a gold recovery. The milling required for gravity recovery of sulfides is always aimed at a coarser grind, but some sulfides are very fine grained and need a finer grind for liberation, this makes the recovery of the sulfides by gravity a much more difficult process. Usually at this point the sulfides are recovered by flotation rather than by gravity. Many flotation reagents will give recovery of free gold as well as sulfides.

If you carry out fraction assays and find that the sulfides are not viable to float then you are looking at gravity separation. Tests must be run to find the minimum grind size which liberates the gold. Milling finer will flatten the gold particles and make gravity separation more difficult.

The maximum gold recovery gotten by gravity separation is dependent on a mix of all of the above factors, there is no all encompassing number for % recovery unless the system is optimised for the above factors.

The project I wrote about was run by a large mining company, so all of the test work above was done and the best size for milling was established.

Table recovery of total gold was around 80%, it was not viable to chase the remainder which was associated with sulfides and other lower grade ore particles.

To get this recovery % we had to fit the water spray bars on the primary tables in a similar configuration to the ones on the clean up table. This increased the recovery from 50% to 80%, no viable grades of free gold were found in the table tailings for either the full size primary or smaller secondary tables.

Deano


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## jobinyt

I suggest that knowing what someone else got, someplace else, is of little value. However, you asked the right question "What should you expect?" What you are doing is density separation. For that you need well classified material - sized just above and just below the material you want to recover. Take a large sample - screened to size of what you plan to run. Weigh and volume measure a sample to be run; watch closely, and see what you get. Then split the run material in half - bigger and smaller - and run each again. Then, take 2 more samples from the original lot: one the next mesh size up, the other the next mesh size down and run those across your table. Again, watch closely, and see what you get. Sometimes your target will vary in size enough to make more than 1 classification and run necessary. At any rate, you should now have a good idea of what you can do with your material. The key is classification - tight size ranges allow separating particles of about the same physical size but very different densities. Only through experiment will you learn what you can recover from your materiel.


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## SamW

Thanks Deano and jobinyt. That gives me some good information to work with. I plan on sharing my progress in a new thread soon!


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## jobinyt

I look forward to it.


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## goldshark

Just a quick note to say that surface tension of the water is a big part of fine gold recovery. Using a recirculated water source with Jet Dry type additives, helps to relieve the problem. the water can get pretty heavy with fine silts, so a rather large settling pond helps.


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## Yggdrasil

I came over a text regarding one of these "green" gold leachants.
In the text there was a recommendation on mixing the leachant with HypoChlorite leach.
Since the original "green" solution is made up of SodiumCyanate, I wondered how these two play together.
You can push the HypoChlorite leach, to say, pH 9 and then pull the Cyanate leach down to the same.
But one end up with two leaches in not ideal zones and in the end even may destroy each others.

Are there any credit to this kind of process?
Aren't it better to alternate them?

Any thoughts from our experts here?


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## Deano

In any pH form of hypochlorite leaching there is always some free chlorine in the leach solution, you can smell the chlorine evolved from the leach. The greater the agitation applied to the leach solution the greater the rate of evolution of the free chlorine.
Free chlorine in solution likes nothing better than to oxidise soluble organic complexes such as cyanide and its derivatives.
If you run a mix of cyanide and hypochlorite in a leach solution you will end up with a solution containing chloride complexes from the hypochlorite and oxidation products such as carbon dioxide from the cyanide.
Not the cheapest or most sensible form of leaching solution.
I have always had high regard for people who try new things but a little research would say that this mixture is not going to be a winner either as a gold leach or as a cheap gold leach.
Deano


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## Yggdrasil

Deano said:


> In any pH form of hypochlorite leaching there is always some free chlorine in the leach solution, you can smell the chlorine evolved from the leach. The greater the agitation applied to the leach solution the greater the rate of evolution of the free chlorine.
> Free chlorine in solution likes nothing better than to oxidise soluble organic complexes such as cyanide and its derivatives.
> If you run a mix of cyanide and hypochlorite in a leach solution you will end up with a solution containing chloride complexes from the hypochlorite and oxidation products such as carbon dioxide from the cyanide.
> Not the cheapest or most sensible form of leaching solution.
> I have always had high regard for people who try new things but a little research would say that this mixture is not going to be a winner either as a gold leach or as a cheap gold leach.
> Deano


Excellent, you confirmed my suspicion.
I found this recommendation odd and wanted a second opinion.

Thanks Deano


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