Processing old MLCCs - disclosed process

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orvi

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For a long time, there were numerous threads and questions regarding processing MLCCs. So I decided to show my way on how I do them, share my experiences and suggestions, and whole process in general to the PGM dore. In the future, I will probably come to the disclosure of the purification of the multi-metal dore to the PM mixture (speaking about PtPdAg here). Since I do not have any interest in resolving the Pt and Pd one from another.

FEED: 192 + 102 g of old soviet MLCCs. 192 g of orange and yellow "KMs" and 102 g of green "KMs". These components are famous for their very nice PtPd content, some types climbing to nearly 70g/kg PGMs. It is obvious that bigger pieces has better yields, as ratio of the weight of ceramics to the "dead" weight of legs, resin and solder is much higher. So because of this, yield fluctuate from like 25-65g/kg depending on type, size and manufacture year. There are exceptions, but usually yield of mixed batch is somwhere in the middle of this. And with ratio of Pt-Pd fluctuating depending on what types are predominantely present (there are mainly Pd ones with trace Pt, and Pt ones with trace Pd).

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Stock photo of the issued material - "fat" orange/red/yellow and green ones were present in this batch. There is numerous more types with varying PGM content.

COMPOSITION: orange KMs are pieces of ceramic with soldered legs on the sides of the ceramic rectangle, dipped in thick resin filled with silica dust. Material of the ceramic is predominantely BaTiO3, but strangely - ceramic also contain bismuth and some % of Si and Al. Embedded in ceramic, there are PGMs foils with small % of silver. On the sides of the ceramic "sandwich", ends of PGM foils are soldered to the copper legs. Solder is typically of composition of PbSnCd. I do also measure some silver. Mainly because cadmium in the solder, this material is cumbersome to process in "contained" way, because cadmium will burn happily when heated upper than 400°C or so. Also, there is toxic barium in the ceramics, which can be easily leached out by acids, producing very toxic solutions, which should be dealt with responsibly.

TYPICAL WAYS OF PROCESSING: On you tube, Owltech has great series of videos issuing processing this refractory and hard material. I recommend watching them and learn from his experience, successes, mistakes and advices. Generally, first step is to remove the resin coating off the caps. There are two main ways of doing this. First easy looking, but far harder to do safely - incineration. And second more tedious, but more contained - solvent disintegration or digestion.

With incineration, material is burnt - by doing this, cadmium in the solder also burn and evaporate as CdO yellow-brown cobwebs and smoke. Very hazardous operation, if not done in good fumehood with scrubber. Solder typically contain up to 10% Cd in mixed lots (some caps do not have Cd in solder, some do), and solder comprise of roughly 10-15% weight of the average lot, so not negligible ammount.

With solvent disintegration or digestion, caps are soaked in organic solvent (usually DMSO, but DMF also works), which swell the resin coating and enable to peel it off. This process is tedious as you need to do it one by one by hand. But it is very "clean" and all Cd stays in place. Digestion is decomposition of the resn in strong NaOH boiling solution. This operation is very messy - as honey-like goo is mainly produced from orange/yellow types of caps, which is then PITA to strip off the ceramics. Only convenient way I know is to decant the liquid, wash the goo with water, drip out as much water as possible and then cover with DMSO and take it near to the boiling point of DMSO (to at least 150°C) - this will slowly liquify the goo and it can be poured off the ceramics. On many occasions, whole boards with components are dipped in clear varnish - and this is resistent to NaOH. So in this case, you need to prior remove the varnish and then proceed with NaOH disintegration.

Overall, with ceramics stripped from resin, you heat the material in mesh basket with torch or more conveniently heatgun, and by shaking motion shake off the legs and good majority of the solder out. By doing this, you "enrich" the metal fraction in PGMs, removing the stubborn Sn, Pb and Cd - and practically all copper. Little PGMs follow the solder and are contained in it, but if the ceramics are not too much beaten or cracked, it is low number.

With bare ceramics in hand, there are generally two ways of approach. Pyro and wet route. Pyrometallurgic one is based most of the times on pulverizing the ceramics in the grinder, then mixing them with fluxes and collector metals and smelting them to obtain metallic dore, which is then cupelled free from lead (or scorifyed) and processed with nitric acid.
Wet route rely on disintegration of the ceramics in boiling HCL leaving PGM foils and some undissolved ceramic stuff, which is very crumbly and is washed from foils easily. Recovery rates are varying depending on the flux mixtures, temperature of the smelt etc... But generally, if not appropriate flux is used and temperature is lower than 1200°C or so, yields decrease. Wet route is time consuming, producing quite a bit of toxic liquid waste - but can be done in near perfect recovery. So can be pyro route, but conditions for achieving full recovery by pyro means are much more harsh, and propane furnance rarely gonna make it :)
 
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MY WAY OF PROCESSING:
I developed this route by several years of trying and alteration, and with great help of other refiners here, or in the forum. Espetially I am very greatful to Owltech and his educating videos, which also helped me a lot to develop process, which produce minimal waste and minimal trouble.
I usually go with slow incineration route in contained vessel with applied afterburner and fume scrubber to deal with Cd, but unfortunately I do not have means to do this right now from various reasons. So instead I used solvent approach now, followed by de-soldering the legs and direct smelting of ceramics in induction furnance.

RESIN STRIPPING FROM ORANGE CAPS:
I typically obtain the varnish-coated material, so first step is varnish removal. As I have good access to used laboratory solvents (considered as liquid waste), I generally use dichloromethane (DCM) for this, as DCM is very quick in this job, swelling and flaking off the resin in matter of dozens of minutes. And also fairly cheap and accessible. I use to soak for few hours tho - for one very good reason: DCM is very volatile (and also not very healthy) solvent, with boiling point of 40°C. It will soak into the resin very well and stay there for a while even if the heat is applied. Then it is slowly released.

With frequent mechanical stirring with stick or glass rod, varnish flake off from majority of the surface. Then I decant the used solvent and wash the caps in water. I prepare like 20% NaOH solution, cover the soaked caps with it in the beaker and start to simmer.

This is where the magic happen :) Immediately as hot NaOH solution touch the caps, they start effervescing the DCM, simultaneously the decomposition of the resin begin. By heating near the boiling point, escaping DCM solvent bubbles carry the resin goo to the surface of the liquid, where it is very convenient to scoop out with some old spoon. I do not need to imply that this operation must be done in good fumehood as DCM will evaporate and it is quite nasty to the human health if inhaled. It also does not smell nice and make terrible headaches and nausea after exposure. On larger scale, vapors can be distilled and collected solvent reused.

Of course, there is residual goo on the remains of caps which does not go to the surface, but you can get out good 2/3 of it by doing this.
Then, you decant off the liquid, wash the gooish ceramics with water and cover them with DMSO. Heat it strongly up to 150°C and above to liquify the goo and make it decantable. It takes few dozen of minutes to fully dissolve, so be patient. After decanting the DMSO goo, I wash now bare ceramics with water, and they are good to go to next stage.

RESIN STRIPPING OF GREEN CAPS:
With these, soaking them in DCM or DMF/DMSO is fully enough to flake off the resin. No need for NaOH boiling.

DE-SOLDERING:
With resin-stripped material, I move to solder-legs removal. I purchased small stainless steel basket (kitchen utensils) for this. With holes large enough for legs to fell through, but small enough to keep the smallest caps in. With heatgun, I heat and shake the material, which cause solder to melt and fell through the holes to the underlying stainless tray or pan. You will obtain bare ceramic rectangles with minimal solder on them ready to be smelted.

SMELTING:
This is the most critical operation, and it should be done right in order to get good recovery. Remember that BaTiO3 is fairly refractory and you need to apply high temperatures to properly liquify it in order to liberate PGMs from it. From my experience, no collector metal is needed, as there is still some few % Ag and lead/tin residues from the solder. And also bismuth in the ceramics, which will be reduced upon contact with graphite crucible. Typically, 50% PGM dore is obtained, which is enough dilution for PGMs to stay liquid at temperature of smelting.

I advise to use induction furnance,
as it is very convenient, no issues with attaining high temperature, and efficient in terms of required energy vs. actual used energy. Also, I must point out that even if majority of solder was removed, some Cd will still remain in the traces of solder, and this will burn out completely to CdO during the smelting stage - so anyway reducing the ammount of Cd present, some is still there and it must be considered before any heating or smelting. Good fume exhaust with scrubber is strongly advised for both sake you and the enviroment. I do not want to overload and fill up the scrubber quickly, so that is why I insist on the best sanely possible way to lower the content of Pb and Cd in the material.

Generally, there are two simple methods to fully decompose the ceramic - use of silica as flux, and boric acid as flux. By using 20wt% SiO2 to the whole de-soldered rectangles, you achieve eutectic of BaTiO3-SiO2 with melting point about 1260°C (+-, it depends batch to batch). There are actually two eutectics, so little less or little more (few%) will still do fairly well. Anyway, if you add significantly less or more, mp will raise sharply and you will need to go to very high temperatures to do this properly. Nice thing about BaTiO3-SiO2 melt is that it is very fluid and liquidy if mixed properly. So there is no issue with viscosity. It is worth to note that initializing the fluxing process require higher temperature than 1260°C - as sand and ceramics must effectively contact one another :) so if you struggle to attain like 1400°C from the start, it is wise to cake some of the charge with oxy/propane torch from the start to ease the onset of the fusion process.

On the other hand, boric acid upon dehydration converts to boron oxide - B2O3. This has very low mp, and eutectic of BaTiO3-B2O3 system goes as low as 1020 °C. By using stoichiometric boron oxide to the titanate, you obtain composition with melting point of roughly 1100°C. This is done by adding 1 part of boric acid to 1,9 parts persumed BaTiO3. Lowest melting point is at 30% less B2O3 than stoichiometry, so I advise to add 1 part of boric acid to roughly 2,5 parts of ceramics and observe the behaviour. And trust me, less is sometimes better than overshooting. Because if you add overstoichiometric boric acid, resulting slag would be viscous like honey :) even at very high temperatures.

All in all, BaTiO3-B2O3 system is more viscous than silica one, with advantage to be only possibility to run it at low temperatures. Without need of previously cranking up the temperature to kickstart the fusion - as B2O3 has melting point of less than 500°C. But it should be considered that at 1100°C, there maybe needed some collector addition for PGMs to keep them molten properly at these lower temperatures - which arise further problems during refining :) And don´t forget, the weight of the de-soldered ceramics is not the real weight of just ceramics, but also residual solder and PGMs.

PHOTOS OF THE LAST RECOVERY:
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De-varnishing and soaking the orange KMs in DCM. As you observe them, you will notice how they swell and enlarge :)

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Decomposition of the resin in NaOH. Notice the honey like goo floating on the top of the liquid, with partially stripped ceramics on the bottom. This is half-way done. Bubbling is caused by slow liberation of soaked DCM in hot solution, which instantaineously boil as it escape the resin.
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De-soldered ceramic rectangles in the stainless basket.
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Onset of the heating and smelting in induction furnance, using silica as flux. Mold is prepared, then heated on the top of the crucible, and slightly tilted to better coalesce the PGM dore on the bottom. It is wise to just pour little bit of the slag to the mold firstly, wait like 15 seconds to let the fine coating of slag on the bottom to solidify, and then pour the rest. This will prevent the issue of sticking the metal to the iron mold.
IMG_20221116_161947.jpg IMG_20221116_161956.jpg
Resulting PGM button on the left and very nice, homogenous glassy fragile slag on the right :) No sticking to the mold, very brittle, easy to break - and very very heavy slag due to barium.

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XRF of the dore, cracked in half to reveal the fresh metal surface. With these alloys of PGMs and base metals, many times even this small beads aren´t homogenous and best way how to relatively reliably measure their composition is to crack it in half in vice with hammer and shot the shown inside.

Weight of the button is 28,55 grams. That said, average PGM content in the batch was 52g/kg - which is very good result since it was mixed batch :)
I like to separate at least partially the Pt and Pd caps, as too much Pt will render nitric dore processing almost impossible, and as AR processing is extremely slow due to Pb and Ag content, I like the nitric treatment more. If you have just Pd caps, it would be much easier to process than mixed PdPt batches.

So, that´s it :) I hope it will be useful for someone struggling.
orvi
 

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All my resin-coated MLCCs are made in America, Japan, Mexico, or Taiwan. I do have a bunch those mid-size 'fat' ones with flat stacked foil sheets, also made in the same countries. I will test a few to see if those have any silver, but I think they're just Al-coated polymer sheets from reading up on the later-era resin-coated foil capacitors just before MLCCs became dominant.

Easy enough to crush a few and toss them into HCl. If the metal of the foil dissolves, it's not silver.
 
I was expecting a higher Au content.... That old soviet stuff must be pretty good.

Nice work...(y)
Gold isn´t usually found in MLCCs. PtPd were metals of choice for early designs, as they do not oxidize appreciably when ceramic-metal foil sandwich is sintered. Gold has not so good properties and low melting point. That is also reason why there practically isn´t such a thing as gold traces sintered in the ceramic (legends about CPUs and similar discussions). And why kovar is mainly used for leads, sintered into the ceramic matrix of the chip.

Anyway, I do not know from where the gold came from, but as it is a fairly complex alloy, it could be possibly "artificial" = XRF somehow assumed there is some gold, but it probably isn´t. Or some residual gold from the graphite crucible - which of course was used multiple times before this smelt :)
 
VERY nice tutorial orvi !!!!
Mold is prepared, then heated on the top of the crucible, and slightly tilted to better coalesce the PGM dore on the bottom.

one suggestion - if you are going to do any real amount of smelting I highly recommend investing in a cone mold like this

https://www.lmine.com/conical-slag-molds-c-1_67_115/conical-slag-mold8-p-3917.html
Edit to add; - this mold will take "up to" a 1.38 liter pour (so that is a maximum pour - smaller pours can certainly be done)

If something like that is to big for your smelt batch size you can custom make a sort of cone mold (to better accommodate you smelt size) by welding end caps on a piece of angle iron so the pour goes to the bottom of the V --- short & deep is better then long & shallow

Or - you can cut 4 iron triangles & weld them into a pyramid (again to make a mold customized to your smelt size)

pouring to a cone (or V) will give you much better separation of the metal from the slag - especially if the smelt/flux/slag runs nice & thin/fluid

if you properly carbonize the inside of your mold you should have little or NO problem with metal/slag sticking to the mold

Edit to add; - I LOVE my cone mold :cool::love:

Kurt
 
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VERY nice tutorial orvi !!!!


one suggestion - if you are going to do any real amount of smelting I highly recommend investing in a cone mold like this

https://www.lmine.com/conical-slag-molds-c-1_67_115/conical-slag-mold8-p-3917.html
Edit to add; - this mold will take "up to" a 1.38 liter pour (so that is a maximum pour - smaller pours can certainly be done)

If something like that is to big for your smelt batch size you can custom make a sort of cone mold (to better accommodate you smelt size) by welding end caps on a piece of angle iron so the pour goes to the bottom of the V --- short & deep is better then long & shallow

Or - you can cut 4 iron triangles & weld them into a pyramid (again to make a mold customized to your smelt size)

pouring to a cone (or V) will give you much better separation of the metal from the slag - especially if the smelt/flux/slag runs nice & thin/fluid

if you properly carbonize the inside of your mold you should have little or NO problem with metal/slag sticking to the mold

Edit to add; - I LOVE my cone mold :cool::love:

Kurt
Yeah, we do have various molds for smelting, altough we rarely do smelts bigger than 10kg of material. And these are mostly metal, so issue of separation isn't that relevant. But I have this idea that i let someone weld one for me in mind for a long time.. Because, no matter how many trades I do fairly well, welding and I - we aren't friends :D

This smelt was done bit under pressure, I needed to do it that day cos we are rebuilding the hood, equipping with stronger fan etc... Possible that no smelting would be possible for upcoming week or two. There was like 2 hour gap for me, so I needed to improvise a bit :)
 
Because, no matter how many trades I do fairly well, welding and I - we aren't friends :D
My brother is a licensed boiler maker. He can weld the "stack of dimes" upside down. He paid me a high compliment a while back. He told me I had graduated from bubble gum welds to gorilla welds. Still not that pretty, but at least they're strong. Practice, practice, practice.

Dave
 
Tested a couple of the small foil-layer capacitors of different colors, all about the diameter of a pea, and yup, just aluminum coating. The metal dissolved instantly in HCl, leaving just the thin polymer sheets behind.

If you have Western or Asian-made layered foil caps, they're just trash.

Just watch for the ones that are actually coated MLCCs. It can be tricky at that intermediate size.Coated MLCC.JPGCoated foil.JPGIntermediate caps.JPGMLCC ones tend to be very flat with slight bulges at the electrode ends, but not always! Some look just like the foil ones, only perhaps slightly smaller and flatter. Those big MLCCs are early generation, and have palladium electrodes. Don't want to toss them out with the aluminum!
 
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