Dissolving large lead bars to recover gold and silver

[Geraldo]

Here are some ways of recovering gold and silver from lead:
(1) Dissolving the lead in hot weak nitric acid. The silver will also dissolve, but it could be cemented with copper. This will generate lots of waste solution and the lead could be difficult to recover.
(2) Cupel the lead as suggested by 4metals.
(3) Melt and use the Parkes process - collect the Au/Ag with zinc. This was for silver but I assume the gold would also be collected. A similar process published by the US Bureau of Mines used aluminum instead of zinc. The BoM process was mainly for removing the PMs and copper from contaminated Pb/Sn wave solder used in the printed circuit industry.

The Parkes process will indeed recover gold as well, although not as quantitatively as silver. It requires increased zinc and residence time. Most lead refineries actually use the Parkes process for gold and silver removal from lead. I will warn you that it is a relatively ugly process to run. Also, like all the pyromet. lead refining processes, it becomes more difficult at the very small scale - skimming becomes harder to do selectively (all skims are lead contaminated, as the amount of skim decreases, the relative amount of lead in the skim goes up), heat losses are large and maintaining temperatures is difficult with a small charge, maintaining correct atmospheres is also more difficult. I know this might sound counter-intuitive, but we found it quite difficult to get very good results with a very-well designed lab-scale continuous drossing furnace, while the full scale version provided easier operation/control (and better results, barring some lethal tin accretion difficulties).

(4) Melt, cast bars, and dissolve and plate out the lead. The Au and Ag would be caught in the anode bag.

Hence my suggestion of the Betts process. You will find that there was a great deal of development in that process (and still is - Teck's research department has tested and continues to test lead electro-refining technology to this day - over 90 years now).

Electrolytic systems are my forte' and I do love the idea of electrorefining PM bearing lead. I downloaded the Betts book and will look it over when I get the chance. With the right electrolyte composition and, if the solution contamination could be kept to a minimum (when the contaminants reach a certain detrimental level, however, most can be removed by selective treatments) , the solution could be used over and over. This would reduce waste solutions, considerably, and would allow the lead to be accumulated, melted, and sold.

Then you will enjoy reading about Betts. There are also still a lot of papers published about the process. Most of the work for the last 15 years has revolved around slimes stability - keeping the PM-bearing slimes on the anode, since nobody bags their anodes in the large refineries. The negative parts about the Betts process - it is relatively slow, it is capital-intensive for a large operation, and it is rare (I think only 3 lead refineries run the process world-wide - everyone else uses a complete pyro-refining circuit).

Being an old plater, I read Gerald's post with interest. Lead is deposited (mainly) from solutions containing either fluoboric, fluosilicic, sulfamic, or (maybe) acetic acids. For the electrorefining of lead containing silver, however, since the silver will co-deposit from the fluoborate system, the fluosilicate or sulfamate systems would be the best electrolyte choices. In these, the Au and Ag would end up in the bagged anode sludge.

The fluosilicic acid (same as the acid used in the Betts system as discussed by Gerald - this acid has many different names) would work great, but it seems to be more dangerous than Gerald thinks it is. However, many things we use are dangerous. With the proper knowledge, setup, and controls, the dangers of most any system can be reduced or nearly eliminated. The best info I could find on the dangers and properties of this acid is in the following link. Although the article is concerned mainly with the use of this acid to produce fluoridation of drinking water, it provides much general information. I would suggest that anyone thinking about using this acid read this article. Please note that, unlike sulfamic acid, fluosilicic acid will etch glass similarly to HF.

http://ntp.niehs.nih.gov/ntp/htdocs/Chem_Background/ExSumPDF/Fluorosilicates.pdf

Here's a general search for the Betts process. It seems that it is the standard method for working lead from sulfide deposits.

https://encrypted.google.com/search?hl=&q=betts+process+lead&sourceid=navclient-ff&rlz=1B3MOZA_enUS408US409&ie=UTF-8&aq=1&oq=Betts+process&lr=all

Unfortunately, it ISN'T the standard method for primary lead recovery - it is actually quite rare. Most primary lead is produced in a pyro-refinery circuit. Generally, these are rows of large pots undergoing various skimming processes (often skimmed with a backhoe), and nearly all batch. I think Trail may be the only refinery to use a continuous copper drossing process - despite the advantages of a continuous process, using a "wheel" of circulating molten lead through copper drossing and softening, most refineries stick to kettle-based approaches.

Were I to experiment with this, especially for use on a small scale, I would definitely attempt to use the sulfamate system, instead of the the fluosilicate. Sulfamic acid is cheap. I recently bought a 50# bag from Univar for $.58/pound and it's available on eBay for about $4/#. Also, sulfamic acid is much safer than the fluosilicic acid. In both systems, if tin is present in the lead, it will co-deposit with the lead.

In order to produce a non-spongy, sound cathode deposit, whether using the fluosilicate or the sulfamate system, a certain amount of lead must be dissolved in the solution before any electrolysis takes place. If you don't start with any lead in the solution, you will spend a lot of time screwing around with the crappy cathode deposit. In this link, notice that the starting solution in the Betts process contains 70g/l of lead, as lead fluosilicate. Also, in the literature, there are additives that will smooth out the deposit even more.

http://www.metsoc.org/virtualtour/processes/zinc-lead/lead.asp

According to the literature (the definitive book, Modern Electroplating, by Lowenheim), the usual operating conditions for a lead sulfamate plating solution are:

Lead - 110 to 165 g/l (as lead sulfamate)
Free sulfamic acid - enough to lower the pH to 1.5
Temp - 24 to 50C (75 to 122F)
Cathode current density - 0.5 to 4 A/dm² (4.6 to 37 A/ft²)

If the pH is too low or the temp is too high, the solution can decompose. The pH of 1.5 (using glass electrodes) is maintained by sulfamic acid (to lower) or ammonia (to raise). In this solution, both the anode and cathode efficiencies are 100%.

The main problem is how to obtain the lead sulfamate (or lead fluosilicate) in order to make up the solution. You can purchase lead sulfamate (or, fluosilicate) but, it may be expensive - for a large scale, though, this would be the best way to go. A cheaper way would be to dissolve either lead hydroxide, lead carbonate, or lead oxide (litharge) in a solution of fairly hot, fairly saturated, sulfamic acid or fluosilicic acid. To make these lead compounds, I would think that one could use sodium hydroxide or sodium carbonate to precipitate these (or a combination of these compounds) from a solution of pure lead dissolved in nitric acid. I might mention that strong nitric won't work well for dissolving lead, as lead nitrate crystals will soon coat the lead and prevent further dissolution. In my experience, a hot solution of 7 parts distilled water and 1 part nitric will dissolve the lead and prevent crystallization.

For economic purposes, the use of acetic acid (vinegar) might be worth a shot also, using similar logic as in the above discussion. That would certainly be the cheapest, but I have no idea as to what the parameters would be. I did find some discussion of this on this sciencemadness thread.

http://www.sciencemadness.org/talk/viewthread.php?tid=10814#pid130783

I am having trouble multi-quoting, so I will try to add some comments via the inferior "re:" method...

Re: fluosilicic acid safety - nothing is "safe." But at <100 g/L acid, the electrolyte is pretty tame. I was scared of the "HF" part at first, having had a very bad experience with HF many years ago (burning a hole right into the bone in my index finger, due to a pinprick hole in my glove). But the silica component basically tames this beast, and it is so dilute as I suggested. I wouldn't recommend bathing in it, but on bare skin it stings less than vinegar. As with all electrochemical processes, you have to avoid making acid mist. However, the voltage of electrorefining is so low, there is no gas evolution, so mist is very limited. If you were to enter a Betts lead electro-refinery, you would find it smelled funny, but that is due to the organic additive. Still, caution is always smart!

Re: acetic acid - I am not aware of any successful lead refining processes using acetic acid. I can't say why this was rejected for the Betts process.

Re: sulfamic acid - there are experimental papers on lead refining using sulfamic, but as I recall there are stability problems and, in particular, plating quality problems for thick deposits. For thin, very adherent deposits, lead electroplating solutions used to contain sulfamic acid. Nowadays there isn't a huge demand for lead electroplate (kinda like cadmium electroplate..."mysteriously" became unpopular *cough* toxic *cough* carcinogenic *cough* bioaccumulates). I would have to look it up though.

Re: initial lead levels - this makes for smooth plating in a continuous process. In the small scale, you can start your solution at 0 lead, run the cell as a "starter" at half-amperage, make crappy cathode for a while until you get a decent solution built up, then pull the cathode and start with a new one. Obviously in the large continuous process, fresh acid is added to make up for losses, but there is rarely any lead added to the solution "on purpose."

Re: contaminant removal - there are troublesome contaminants that could build up, but it will depend entirely on the feedstock. Copper is a problem, as is arsenic and thallium. If it gets bad, you can always run a stripper cell - run some refined lead as both anode and cathode, plate out the impurities on the cathode, then recycle the acid to the main process. While large refineries have to do this occassionally, I never had the slightest trouble with impurities in the "Rubbermaid" cells. A more common contaminant issue is degraded organics, which can be dealt with (at high cost) by filtering electrolyte through activated carbon.

Re: fluosilicic acid supply - in the Trail refinery, it came in via train tanker cars, so I don't know where small amounts are to be had. The large chemical supply houses sell reagent grade, for big dollars of course. Some googling might be in order. Remember that iin electro-refining, the electrolyte, if maintained, doesn't "wear out" - this isn't like leaching with AR or something. Technically these electrolytes last for years and years under continuous use. The organic additives degrade, and sometimes the electrolyte will be contaminated and require purification, but once you have the electrolyte, you can use it "forever."

This is a fun discussion! I still have to read the rest of the thread and the other links.

Best Regards, Gerald

 

[Geraldo]

I would use some old Lead-Acid Battery plates to make up the solution. It's cheap and can be found most anywhere. The Lead Oxide in the plates would dissolve quickly.

Lead acid battery plates often contain antimony, arsenic, tin, copper and/or thallium. The tin, copper and thallium would all cause difficulties (contaminating your cathode and potentially needing to be stripped out if you got them at high enough levels). Also, the "oxide" on the plates is usually contaminated with sulfate, as is the paste. Sulfate would be a significant problem for the Betts process, I believe.

It isn't really necessary to "bump" the electrolyte anyway - it will attain sufficient lead levels in due course.

Best Regards, Gerald

 

[Geraldo]

Excellent idea, especially for a small hobbyist setup! However, with a serious setup, I most always suggest going the full route and using the proper chemicals. This eliminates any unforeseen problems and gets you off to a good start.

In my last post, I kind of ran things together and didn't really separate the systems that I would recommend for the 2 different types of users. To summarize.

For a serious, larger scale, continuous operation. For a trouble-free operation, use the Betts fluosilicate system, exactly as the literature dictates, using lead fluosilicate and fluosilicic acid to make up the solution. This has been the standard system for about 100 years, all the bugs have been worked out, and there's tons of info available. Learn all you can about the safety involved when working with these chemicals and install the proper equipment. Don't cut corners. I would suggest getting some analytical equipment to run frequent analyses of the lead content of the solution and make the needed adjustments. Although the stated efficiencies are 100%, that will not usually be the case. For example, the undissolved slimes on the anode can reduce the anode efficiency. Since the deposit comes from the solution and the lead in the solution comes from the anodes dissolving, any reduction in the anode efficiency will eventually deplete the solution of lead.

For the hobbyist or small scale operators, I would probably play around with the other solutions I mentioned. I think most any of them would work, once you find the right parameters. To put these together and to get the right mix, some math ability would certainly help. And, they are safer than the fluosilicates. To experiment, first use about 1 liter beaker quantities and never change more than one variable at a time.

I will propose that I still love the simplicity and ease of operation of even very small Betts circuits. Of course, I am a little biased, as I haven't tried any other lead electro-refining systems.

Re: bumping lead levels in electrolyte - it isn't necessary. Yes, the very first batch will start out rough. It will improve though as the circuit gets up to steady state. Remember that the lead levels in electrolyte aren't "depleted" by the circuit, like they are in electrowinning. They aren't "established" by adding lead fluosilicate or any other lead source. They are determined by the anode:cathode surface area ratio, and cell flaws (short circuits, non-uniform current densities, crooked/bent anodes/cathodes etc.). And the lead levels in electrolyte of course also reduce the electrical resisitivity of the electrolyte - so for fresh electrolyte, at the very start, plating is slow and current losses are higher. That only persists for a few hours, and is a one-time hit. For my 16 sq ft (roughly) anode rubbermaid cell, it probably cost me an extra 10 cents worth of power.

Re: undissolved stuff on the anode - in the trade, we call them "slimes". It actually takes quite a bit of work to make them stable enough to cling uniformly to the anode, rather than falling to the bottom of the cell. That is why for hobbyists/small refiners I would suggest bagging the anode, so as not to lose any precious gold and silver. Also, avoid stirring of the solution in the cell - fine gold and silver particles can migrate to the cathode and get co-plated with the lead, thereby being lost to the refiner!

Even if the slimes all stay on your anode, the reduction in anode efficiency isn't significant, nor troublesome. Current efficiency goes down slightly, but for the small refiner, this doesn't matter. For the large refiner, current losses are mainly due to shorts, contact losses (contact between the anodes, cathodes and their respective busbars) and of course transformer losses.

Gerald

 

[Geraldo]

Parkes process as noted by GSP is the way to go for speed and simplicity, imho. 8)

The lead stays as metal, and the good stuff sticks to the zinc. It is a good idea to bring the temperature of the lead to some 1200 C, and then let it cool slowly to 1000 C. Copper and other base metals magically float to the surface (no cupelling) and can be skimmed as a dross. Then you add the zinc for the parkes process to collect the goodies.

Here's a photo of the Parkes process in operation. Doesn't look like much fun to me. I wonder what the average life-span of the guys doing the skimming is?

http://www.daviddarling.info/encyclopedia/S/silver.html

Other things I don't like about it:
(1) 1200C lead!!! Are you sure that's right? If so, lots of bad fumes that must be contained.
(2) The skimmings always contain some lead. This is most always cupelled.
(3) After cupelling the skimmings, the zinc is distilled off. Bad fumes. More containment equipment. You might be able to dissolve the zinc with something like weak H2SO4 without dissolving any silver but, maybe not, since the silver and zinc are alloyed.

All in all, a very dangerous system that requires too many separate operations with much safety equipment required. It's not as simple as you say and I think it would be much more expensive to properly set up.

Give me an electrolytic system any day. Much, much safer, cleaner, and, to me, much, much simpler. The lead melts at about 328C - maybe a little higher (or, it might be lower) with the contamination in it. With the fluosilicates, some fume control would be needed.

You should take a trip to a standard pyro lead refinery sometime. Once you get over how hideous smelting actually is, you will be probably be appalled to see how hideous the entire lead refining circuit is. At the Herculaneum Smelter, despite lots of ventilation, they were still forced to put everyone in full face supplied air respirators in order to keep blood lead levels below the (very high) government-mandated maximums. By the way - our modern concept of "hell", fire and brimstone, comes from Babylonian lead smelting. You had a life expectancy of a couple weeks in those ancient lead smelters, and suffered terribly every single second.

Yes, the Parkes process is ugly. The skim is ugly. Sticky, difficult to skim. The distillation is typically done in sealed refractory retorts, so it isn't so bad. You just have a LOT of them. At some refineries you can walk past ridiculously long rows of retorts, all being heated at one end. By ridiculous, I mean hundreds of them. You have to hope they are set properly (sealed) so they don't crack and spew fume in all directions. That is a mess.

You will also need de-zincing after the Parkes process, since you need to add excess zinc to precipitate silver and gold quantitatively. There is vacuum dezincing, chlorine gas injection, and a caustic process (yes, molten caustic at 1100C).

As I suggested, nearly all the pyromet lead refining steps are fairly ugly.

Copper drossing (add molten elemental sulfur, stir) isn't that bad, and copper dross can be dried with sawdust to minimize entrained lead. If you don't control the temperature, it gets crusty, and then you have a small nightmare. And if you have too much tin and insufficient reducing conditions, you will get tin spinel accretions that will freeze your pot or furnace completely shut. Reducing conditions are typically maintained with either a reducing flame (sometimes not enough - and that is a whole bunch of CO, and requires an afterburner), or addition of coke or tar. Also, if you have a failure in the reducing conditions, you will of course evolve SO2.

Softening (removing antimony and arsenic) is done with oxygen injection and skimming. Not bad, except that you have arsenic trioxide dust pouring out the top of that pot (better hope your vent hood is working well, and of course you have a baghouse for your vent, right?). Oh, and don't ever take your respirator off. At least the dross isn't too bad (but still 40% lead for a large furnace or pot, and typically 99% lead for a 5 liter crucible level melt).

Thallium removal uses zinc chloride to make a PbCl2/ZnCl/TlCl3 dross. That dross is actually quite liquid, so easily skimmed. But again, do you enjoy sucking on thallium dust? Lead chloride is very bioavailable too, unlike metallic lead, so hygiene is even a bigger issue.

I don't want to get into detinning, dead softening, or the rest...processes where they inject chlorine gas...and it gets worse...

Bottom line: pyro lead refining on the small scale requires a really good method of temperature control (induction furnace, excellent thermocouples), a great ventilation system with an afterburner and a baghouse AND emergency gas absorption system (for accidental SO2 or Cl2 releases), plus personal protective equipment (at least a good half-face respirator, preferably full-face and supplied air), and a good set of fire-resistant overalls, leather aprons, welder's gloves etc. Even then, you would be amazed at how well a sticky dross (particularly the chloride-based drosses) burn right through thick welder's gloves. It is only 900C but it cuts like it was a plasma torch or something. Molten lead isn't that bad, but the hot drosses...

Best Regards, Gerald

 

[Geraldo]

Over the last several years, while preparing for retirement and stashing anything I could come across that contained gold, I stumbled upon a lead bar covered in gold. It is thin in places but thick (up to 1/8" ) on one end. A stripping cell comes to mind as one way to recover the goodies. Does anyone have any other options? How much larger than the bar should the stripping cell be?

Are you saying that the bar is 1/8" thick in places or the gold is 1/8" thick in places and thin in others? If the latter, how do you know that? Also, if the latter, that doesn't make much sense at all. If plated, the gold would be thin and fairly uniform over the entire bar. I may be wrong, but I see no way the gold (high melting point) could be cast around the lead (very low melting point). At the high melting point of gold, at 1/8" thick gold, the metals would surely alloy and the bar would look terrible. Also, for what practical reason would the maker of the bar want to apply gold in such an extreme variation of thickness? None, that I can think of. It's not impossible but very unlikely, unless, for some extremely weird reason, the guy maybe placed only the end of the bar in the solution and plated it steady for two days or more - that's about how long it would take to plate 1/8" thick gold (3125 microns). What is considered to be heavy gold eletroplate is only about 2.5 microns thick.

A number of years ago, there were many gold plated lead bars sold as paperweights. The markings on them were made to resemble old bars from the gold rush days. I saw many of them. The plating on them was very thin and the value probably didn't exceed $.25 worth of gold per square inch, at today's market.

If you want to remove the gold, a concentrated sulfuric stripping cell would work best, since it wouldn't appreciably attack the lead. I wouldn't expect there to be much gold, but I've been wrong before.

+1 on Conc H2SO4 stripping cell.

I would go with a few cents of electricity. It really is your friend. Which is funny, because for large scale, electrowinning, refining, dissolution is extraordinarily capital-intensive, and I always steer such projects toward furnaces and fire...

Regardless, I have a tough time seeing how it is economically feasible to recover gold plate from a few bars like that. But I am often wrong...

Best Regards, Gerald

 

[HAuCl4]

Have you actually used the Parkes system, from start to finish?

Of course yes. With modifications. A little nastier than Miller, but not by much.

Look at the problem again: Are you really willing to run/refine 1,000 Kgs of lead through a cell to recover 300 grams of gold and silver per each ton of lead?. How long is it going to take and how big is your cell going to be?. Not viable imho. Worse off than a copper refinery it seems, which is simpler. Maybe I'm just biased, and indeed an skilled electrolytic technician can do it. :?:

Notice that in Parkes, the amount of zinc (or aluminium?. need to investigate this. Do you have the BoM paper?) involved is proportional to the amount of silver and gold to be extracted from the lead, and not proportional to the lead amount. Think solvent extraction at pyrolytic temperatures, that's an easy way to picture it in your mind. 8)

The, now relatively small, amount of silver/gold laden zinc can go directly through a relatively small acid process (inquartation) and get it all done, silver and gold, in 1 day or less. Forget about distilling the zinc from the gold/silver, although it is doable with the proper retort (Ammen has a good sketch I think), it's too much work to recover the few kilos of inexpensive zinc.

 

[HAuCl4]

Just for you, only for you GSP, do not tell anybody else because the setup is "proprietary", ok?. :lol:

1-You need 2 silicon carbide crucibles. One that will be filled almost to the brim with lead (90-95% is ok). Sounds weird but bear with me. The other crucible is 5 times bigger, we'll call it "the blender crucible".
2-You put the feed lead into the small crucible and feed it metal heating slowly with lead barely melt till it is almost full. Then you put 2 pieces of wood charcoal in, and a lid. Heat up in furnace to 1200 C, then let it cool to 700 C-800 C inside furnace, then take the lid out, punch a hole in the top crust beside the pouring lip, and pour lead metal into the big room temperature "blender crucible". Basically all lead pours out ( I have done it with hand held crucibles and with tilt-pour furnaces, and the top copper dross stays in the small crucible). After the small crucible cools, you hammer the dross out by hand.
3-You add zinc to the "blender crucible". 4-5 times the expected silver/gold is fine. Add 2 pieces of wood charcoal. Then put special lid with a hole and a "blender blade" attached to an electric motor in a fixed stand. Heat in furnace to 600-700 C, and turn on the blender for 30 minutes. Smoke processor overhead sucks few fumes that come out. Turn off blender, wait 10 minutes, and let it cool to 500 C and pour the lot into a big conic preheated mold. This is the moment of maximum fumes, if you have charcoal plenty, it isn't too bad.

The ingot when cool is in 2 parts, the bottom is lead and the top is the zinc with the goodies. A hammer blow breaks it cleanly in two pieces. The zinc piece goes to the acid process. The lead can be recast into ingots and sold after you verify with an assay that it contains no PMs.

The bonus is that you will feel like a caveman after you are finished, with a great sense of accomplishment!. 8)

Of course there are details, but not many. You can fill in the holes or ask questions. Simple really. :lol:

 

[Geraldo]

Just for you, only for you GSP, do not tell anybody else because the setup is "proprietary", ok?. :lol:

1-You need 2 silicon carbide crucibles. One that will be filled almost to the brim with lead (90-95% is ok). Sounds weird but bear with me. The other crucible is 5 times bigger, we'll call it "the blender crucible".
2-You put the feed lead into the small crucible and feed it metal heating slowly with lead barely melt till it is almost full. Then you put 2 pieces of wood charcoal in, and a lid. Heat up in furnace to 1200 C, then let it cool to 700 C-800 C inside furnace, then take the lid out, punch a hole in the top crust beside the pouring lip, and pour lead metal into the big room temperature "blender crucible". Basically all lead pours out ( I have done it with hand held crucibles and with tilt-pour furnaces, and the top copper dross stays in the small crucible). After the small crucible cools, you hammer the dross out by hand.

Isn't that a bit hot (800 C) for a kettle-type basic copper drossing? Most kettle-drossing is done just above the lead melting point (usually around 330 C), as it relies primarily on the reduced solubility of copper as the temp decreases.

If you wanted to quantitatively remove copper from the lead, a positive source of sulfur would be preferable (like elemental sulfur or raw tar) - then high temps above 900C are used to ensure fast copper matte production and to avoid the sticky matte from being entrained in the molten lead. Or are you counting on the wood charcoal having sufficient sulfur?

3-You add zinc to the "blender crucible". 4-5 times the expected silver/gold is fine. Add 2 pieces of wood charcoal. Then put special lid with a hole and a "blender blade" attached to an electric motor in a fixed stand. Heat in furnace to 600-700 C, and turn on the blender for 30 minutes. Smoke processor overhead sucks few fumes that come out. Turn off blender, wait 10 minutes, and let it cool to 500 C and pour the lot into a big conic preheated mold. This is the moment of maximum fumes, if you have charcoal plenty, it isn't too bad.

The ingot when cool is in 2 parts, the bottom is lead and the top is the zinc with the goodies. A hammer blow breaks it cleanly in two pieces. The zinc piece goes to the acid process.

Good method of separating Parkes dross from the lead! I believe this used to be done by a number of lead refineries some decades ago. I recall a key was to keep the molds heated or insulated to cool the lead/dross mix as slowly as possible. Not a good method for a larger refinery, due to the time, handling and real-estate requirements (plus re-melting for further purification stages), but should be easier on the small refiner than trying to skim the wretched Parkes dross.

The lead can be recast into ingots and sold after you verify with an assay that it contains no PMs.

The bonus is that you will feel like a caveman after you are finished, with a great sense of accomplishment!. 8)

Of course there are details, but not many. You can fill in the holes or ask questions. Simple really. :lol:

I am assuming you are disposing of these lead ingots as scrap that will be fed to some primary lead refinery somewhere that can deal with the zinc contamination? Or are you suggesting drossing out the zinc from the lead with caustic or chlorine?

Best Regards, Gerald

 

[mikesmadness]

[Upon thought, I totally agree with your views on temperature related layers. I might have to re-visit my beliefs in this piece. Thank you to all in this forum!!![/i]

 

[HAuCl4]

Geraldo: I'm assuming the OP just wants the gold and silver out at minimal cost, and to sell the lead, copper dross, etc to a bigger processor, as would I. This is no attempt on my part to make a lead, copper or zinc refinery out of this contraption setup. The drossing, before the "blending" doesn't need to be perfect either, it is just to mostly clean the lead from base metals and make the zinc-gold-silver drossing afterwards much simpler.

edit: One important bit that I missed is that always check the bottom of the crucibles after they are cool, because if you have iridium or other heavy metals it may/will be there stuck to the crucible. It doesn't apply to 99.99% of the cases though but it is easy to remove by throwing a few ounces of silver to collect it and melting again and extracting the cool button if you are in the 0.01% of cases. :lol:

 

[goldsilverpro]

Gerald,

I must admit that I have no hands-on experience in the electrorefining of lead and I bow to your obvious superior knowledge on the subject. My posts were based on my knowledge of plating and the large plating library that I own. I did spend a week in the million gallon Anaconda cellhouse (electrorefining of copper) in Butte, MT, in the early 80s. And, I spent a couple of months developing a successful electrowinning setup for a guy that had accumulated a million gallons of contaminated nickel plating solution that averaged about 0.6 pounds of nickel/gallon. That's about it. I did spend about 10 years in the plating industry and, since then, I have a lot of experience using electrolytic systems in PM recovery. In general, the solutions using in plating are almost identical to those used in electrorefining. The differences are the end product and the contamination. With plating, the solutions are quite pure with no contamination and the quality of the cathode deposit is everything. With electrorefining, there is much contamination to deal with and this presents unique problems.

I did read (actually, heavily skimmed - it was quite long and there seemed to be much superfluous information in it - history, etc.) the Betts book you gave a link for, with great interest. To me, though, one of the most interesting portions was his discussion on the Moebius and Thum silver cells. He experimented with using small amounts of gelatin (I think about .1% = 1 g/l) in the standard silver nitrate cell solution and says that, instead of producing crystals, the deposit was smooth and adhered to the cathode, not unlike an electroformed plating deposit. With the horizontal Thum cell, this would not work well. With the vertical Moebius cell, however, this might work great on fairly thick passivated stainless cathode sheets. He says that the deposit is brittle, which would be good, since it could probably be broken loose in chunks from the stainless, similarly to the way the brittle silver deposit is removed, by tapping with a hammer, from the rotating stainless drums used to plate out the silver from photo fixer and bleach-fix solutions. This could eliminate the need for constantly dealing with the collection of the loose crystal in the cell. Here again, I only skimmed this information and will read it more carefully to make sure that what I am saying is correct. He also discussed a nasty sounding electrolyte made from concentrated sulfuric and methanol and I may be confusing the two. I have used gelatin as a grain refiner in several types of plating solutions. If I remember right, the gelatin was prepared by first swelling it in a little room temp water and then adding hot water to dissolve it. He also mentions using gelatin in the fluosilicate lead solution.

HAuCl4,

I searched again, with no luck, for the BoM RI involving the use of aluminum to remove Au/Ag/Cu from contaminated Pb/Sn solder. I know it exists because I once had a copy of it. We discussed this once before on the following thread. On the 2nd link below, the publication IC 9059 includes the article, "Precious metals recover from electronic scrap and solder used in electronics." They are selling this for $28. Yesterday, I found another place that sold it for $12, but am unable to find it today. However, this article may be different than the one I had. I seem to remember that it was an RI and this is an IC.

http://www.goldrefiningforum.com/phpBB3/viewtopic.php?f=49&t=8415
http://www.woodenski.com/2neat/usbm/usbminformationcirculars.html

When iridium is present in a silver assay bead, it doesn't alloy with the silver. Instead, it appears as black particles (actually, tiny beads) mechanically stuck to the bottom of the bead and you have to be careful not to dislodge them. However, this is probably due to the low temperature used in fire assaying. In the phase diagram, it seems that no alloying takes place below 1000C. Then it starts to alloy slowly, with an alloy of about 2% Ir being formed at about 1250C. To keep it in the silver, I would think you would have to solidify the alloy quickly. Otherwise, I would think the much heavier Ir would precipitate and settle to the bottom. This might be a good thing, though. If you could keep the silver molten for awhile, say in a pre-heated cone mold, you may only have to deal with the lower portion of the silver when separating out the iridium. Just thinking out loud.

 

[Geraldo]

Geraldo: I'm assuming the OP just wants the gold and silver out at minimal cost, and to sell the lead, copper dross, etc to a bigger processor, as would I. This is no attempt on my part to make a lead, copper or zinc refinery out of this contraption setup. The drossing, before the "blending" doesn't need to be perfect either, it is just to mostly clean the lead from base metals and make the zinc-gold-silver drossing afterwards much simpler.

edit: One important bit that I missed is that always check the bottom of the crucibles after they are cool, because if you have iridium or other heavy metals it may/will be there stuck to the crucible. It doesn't apply to 99.99% of the cases though but it is easy to remove by throwing a few ounces of silver to collect it and melting again and extracting the cool button if you are in the 0.01% of cases. :lol:

O.K., gotcha. I thought since the OP had 50 pounds a day or something that the lead would be a significant part of the economics. In reality it is about $50/day, which may or may not be significant in this context.

Interesting point about the iridium.

Geraldo

 

[Geraldo]

Gerald,

I must admit that I have no hands-on experience in the electrorefining of lead and I bow to your obvious superior knowledge on the subject. My posts were based on my knowledge of plating and the large plating library that I own. I did spend a week in the million gallon Anaconda cellhouse (electrorefining of copper) in Butte, MT, in the early 80s. And, I spent a couple of months developing a successful electrowinning setup for a guy that had accumulated a million gallons of contaminated nickel plating solution that averaged about 0.6 pounds of nickel/gallon. That's about it. I did spend about 10 years in the plating industry and, since then, I have a lot of experience using electrolytic systems in PM recovery. In general, the solutions using in plating are almost identical to those used in electrorefining. The differences are the end product and the contamination. With plating, the solutions are quite pure with no contamination and the quality of the cathode deposit is everything. With electrorefining, there is much contamination to deal with and this presents unique problems.

Sorry, I didn't mean to come off as some know-it-all expert. I was just excited to get the opportunity to yack about lead refining - it has been years! My friends and acquaintances who still work in the lead industry don't really want to talk process-techno-nerd about their work. And of course, since it has been some years since I worked in a drossing plant or a lead electro-refinery, I have forgotten all the most distasteful aspects. I also forgot some important stuff - in a post above I couldn't remember what the most common bismuth removal process was called - it is the Kroll-Betterton process, using Ca and Mg to cement out Bi - so I skipped it entirely! I used to have such a bullet-proof memory...

I did read (actually, heavily skimmed - it was quite long and there seemed to be much superfluous information in it - history, etc.) the Betts book you gave a link for, with great interest. To me, though, one of the most interesting portions was his discussion on the Moebius and Thum silver cells. He experimented with using small amounts of gelatin (I think about .1% = 1 g/l) in the standard silver nitrate cell solution and says that, instead of producing crystals, the deposit was smooth and adhered to the cathode, not unlike an electroformed plating deposit. With the horizontal Thum cell, this would not work well. With the vertical Moebius cell, however, this might work great on fairly thick passivated stainless cathode sheets. He says that the deposit is brittle, which would be good, since it could probably be broken loose in chunks from the stainless, similarly to the way the brittle silver deposit is removed, by tapping with a hammer, from the rotating stainless drums used to plate out the silver from photo fixer and bleach-fix solutions. This could eliminate the need for constantly dealing with the collection of the loose crystal in the cell. Here again, I only skimmed this information and will read it more carefully to make sure that what I am saying is correct. He also discussed a nasty sounding electrolyte made from concentrated sulfuric and methanol and I may be confusing the two. I have used gelatin as a grain refiner in several types of plating solutions. If I remember right, the gelatin was prepared by first swelling it in a little room temp water and then adding hot water to dissolve it. He also mentions using gelatin in the fluosilicate lead solution.

Yeah, Betts talks about a lot of stuff. Remember that this book was a kind of memoir for him as well - the only public face for a lifetime of electrochemistry work at Trail. If you want to know something really shameful - a lot of his research reports and technical memos that he wrote for CM&S (later Cominco, then Teck Cominco, now just Teck) were unceremoniously tossed in the trash some years ago to make room for something mundane like old furniture storage or similar (without scanning etc.). So much of his work was lost, ignored etc. even while working there, so he had to document something publicly.

Gelatin used to be a pretty common plating "smoothener". In zinc electrowinning, it was largely replaced with "glue" (don't know what offhand) which was more stable and reduced organic inclusions in the product. I believe gelatin is still used for electroless silver plating solutions. As you suggest, people don't use it in the Thum cell - AFAIK people avoid additives in Thum Balbach and live with silver dendrite production. I have never personally seen a Moebius cell nor researched one, so I can't comment on the commercial practice. Outokumpu seem to love Moebius cells.

In the modern lLEAD electro-refinery, the two common additives are lignin sulphonate ("goulac") and a series of aloes. No glue, gelatin or similar cathode polarizers.

---

It is an interesting thought to try to make a brittle, semi-adherent silver deposit that could be continuously removed, rather than dealing with scraping dendrites from the bottom of a cell. Let us know if you decide to take that somewhere!

As for hammering - the old metallurgists used to call the 'tapping brittle electro-deposits with a hammer' - - "ringing". In Trail they used to electrowin tellurium, and recovered it by hitting the stainless cathode with a hammer, causing the plate to "ring" and making the brittle tellurium fracture and "pop" off. It is still done in primary nickel electrowinning - first the cathodes go through a shaker/vibrator that wiggles the cathode, then a big guy with a sledgehammer starts beating that plate until the nickel fractures off. The only drawback that I see is that the little pieces often fly all over the place - something that I wouldn't want to see if it were silver or gold. Somehow, the ringing cathode needs to be very well contained so that all the little bits are recovered.

Best Regards, Gerald

 

[Geraldo]

I searched again, with no luck, for the BoM RI involving the use of aluminum to remove Au/Ag/Cu from contaminated Pb/Sn solder. I know it exists because I once had a copy of it. We discussed this once before on the following thread. On the 2nd link below, the publication IC 9059 includes the article, "Precious metals recover from electronic scrap and solder used in electronics." They are selling this for $28. Yesterday, I found another place that sold it for $12, but am unable to find it today. However, this article may be different than the one I had. I seem to remember that it was an RI and this is an IC.

http://www.goldrefiningforum.com/phpBB3/viewtopic.php?f=49&t=8415
http://www.woodenski.com/2neat/usbm/usbminformationcirculars.html

Hmm. The modern Parkes process actually uses a Zn-Al alloy (pretty small Al content however, a few % if memory serves). This apparently allows operation at higher temperatures, reduces zinc oxide formation, and increases cementation kinetics (probably some sort of surface chemistry/passivation thing). I haven't encountered aluminum powder for Ag/Au recovery, but it makes sense - it is a strong reducer, forms alloys with most lead impurities like Ag, Au (and unfortunately, Cu, Zn, etc) and doesn't alloy with Pb. However, it does form alloys with tin - wouldn't it pull all the tin out of the Pb-Sn along with gold and silver etc., making a cement containing "everything but the kitchen sink"? Obviously the BoM avoided that somehow...

Or are you talking about cementing from an aqueous solution using Al powder? Because some Googling just now reveals a summary from that USBoM circular as:

Known in the prior art is an ecologically clean method for hydrometallurgical processing of friable concentration products of scrap of electronic instruments containing precious and nonferrous metals (see Dunning B. N. "Precious metals recovery from electronic scrap and solder used in electronics manufacture", Inf. Circ. Bur. Mines U.S. Dep. Inter.", 1986, N 9059, p. 44-56). The method involves treating the concentration products with alkali solution (20% solution of sodium hydroxide), passing aluminium into the solution and obtaining a solid sediment which is then treated with a sulfuric acid solution in an autoclave at a surplus oxygen pressure, dissolving nonferrous metals into solution and obtaining enriched concentrate of precious metals in the solid sediment. Said solid sediment is leached by countercurrent under pressure in three stages for stage-by-stage extraction of lead and tin and fuller dissolution of copper.

In the first stage of this method, i.e. in the process of alkalinous treatment preceding acidic leaching it becomes possible to dissolve almost all aluminium from the source material, said aluminium precipitating into sediment in the acidic medium. Besides, acidic leaching in the autoclave under oxygen pressure has made it possible to intensify somewhat the dissolution of tin and lead and to facilitate the dissolution of copper. However, to improve dissolving of copper, the acidic leaching process in the known method is conducted in three stages which increases the duration of the process and, consequently, reduces its efficiency.

Besides, the known method is noted for a comparatively low extraction of silver (82%) into enriched concentrate.

which is from this patent that claims to improve that process:
http://www.patents.com/us-5190578.html

Regards, Gerald

 

[goldsilverpro]

Gerald,

I meant what I said and didn't think you had a know-it-all attitude. I was very interested in your presentation. It's always great to listen to someone that has actually been there, big time.

---

From the summary you found, the process in IC 9059 is totally different.

The one I had melted the Pb/Sn solder and stirred in aluminum. The Al picked up the Cu, Au, and Ag, which were essentially the only contaminants, and purified the solder. And, like I said, it was a BoM RI (Report of Investigation). I would guess it was published in the mid to late 80s. Over the years, I have spent several search sessions trying to find it, with no luck. I feel sure that the mining schools, such as Montana School of Mines or the Rolla campus of the University of Missouri, would have it in their libraries.

The contaminated 60/40, Sn/Pb solder came from wave soldering machines used to solder components to circuit boards. The leads from the components are put through the holes of the boards, trimmed, and then the bottom of the board rides over the crest of the solder wave. The Au and Ag plating on the leads and a little Cu from the leads dissolve into the solder. At some point (I'm thinking from 2%-4% total contaminants), cold solder joints occur and the solder must be replaced.

----

I'm thinking that the glue used as an additive in some plating processes is hide glue. When I used it, I think I bought it from the plating suppliers. Beta-napthol is another additive used.

---

The nickel electrowinning operation of waste nickel sulfate plating solutions I mentioned was interesting, at least to me. The first tank we set up was a rubber lined steel tank of about 1000 - 2000 gallons (because the guy happened to already have this tank). We used bagged carbon anodes. Since the anodes were inert, they split water and the H+ ion entering the solution continually lowered the pH. Ideally, the nickel deposited at a pH of about 4. As the pH went down, the cathode efficiency dropped and, at some point (pH of 1?), little or no nickel deposited. To continually raise the pH and to replenish the nickel in the solution without a build up of the other ingredients present in the solutions (boric acid, etc.), we hung nickel carbonate in anode bags throughout the tank - phase 2, which never happened, would have involved circulating the cell solution through a large filter type chamber filled with nickel carbonate - maybe a fiberglass swimming pool sand filter. We started the tank with a solution containing everything needed to get sound plating (NiSO4, boric acid, etc). The nickel carbonate was made separately by adding sodium carbonate to nickel solution. We had a neat conveyorized filter that continuously produced about a 3" x 18" cake of rinsed, semi-dry nickel carbonate.

The nickel was deposited on passivated stainless cathodes. To prevent the nickel from wrapping around the edges, 3M plater's tape was used on the edges. Since we used sodium saccharin as a grain refiner in the solution, the nickel sheet (about 1/8" thick) could easily be removed in a few large pieces (It was still a little brittle), rinsed, dried, and sold. There was a little sulfur in the nickel from the saccharin but it didn't affect the selling price much.

All went smoothly for about a month and the same original solution was still pumping out beautiful nickel, at a maxed rate. The guy had stored the solutions in about 2000 steel drums with plastic liners in them. They were stacked about 3 pallets high in a fenced, paved yard. There were a few leakers and a few (maybe six) very small (maybe 2' wide) emerald green puddles on the pavement. The EPA spotted these and found out that the guy had originally transported the drums improperly. They brought in a huge tandem Marine lab truck with about 20 people and confiscated everything, including the tank, while wearing full rubber suits and masks (showtime!), and sent the guy about a $500,000 bill. It was really a chicken sh*t deal. I wouldn't want to swim in it, but that nickel solution is relatively innocuous. I heard later that some EPA guys had come by earlier and he stupidly ran them off. Of course, this pi**ed them off and they wanted him bad. Oh well.

 

[kjavanb123]

Hi,

During depopulating 14.7 kg of small socket PC boards, I have collected 470 grams of mostly solders, and SMDs. Since majority of this lot is solders, I was thinking to melt the whole thing, and run it in a solder (tin) parting cell, and I came across this informative threat.
I wonder if the same lead parting cell can be applied to this solder ingot that I have? Thanks

KJ