Further things which may be of interest to members

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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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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
 
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.
 
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
 
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.
 
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
 
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
 
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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