The recovery of a marketable nickel sulfate has been outlined. Arsenic is also removed in the stripping of the copper prior to crystallizing nickel sulfate but, unfortunately, this copper-arsenic sludge is too contaminated with other elements—principally antimony and bismuth—to permit the sludge to be utilized as a source for a marketable copper-arsenic alloy. Hence, this sludge must be circulated back and the arsenic and other impurities eventually find outlet in the smelter circuit. While antimony, always, and bismuth, sometimes, dissolve sufficiently in the electrolyte to be a nuisance and frequently a headache, these two elements, together with some arsenic, report in the slimes and are removed along with lead and wind up eventually in a lead smelter or refinery. As illustrated in Fig. 12, the step by which these elements are removed is called cupellation and is performed on the slimes after they have been first leached to remove the substantial amount of copper always found in the slimes. After copper removal, the slimes melt rather readily especially if enough soda is added to combine with the sulfate radical, thus freeing the lead and making it possible to readily oxidize the lead arsenic, antimony, and bismuth into a low-melting oxide slag readily treated in a lead smelter for the recovery of these elements and the inevitable contaminating noble metals.

(It should be noted that lead, arsenic, antimony, and bismuth have dual outlets from the copper circuit; one from the smelter and the other from the refinery. While the quantity of these impurities that must be removed from the refinery circuit could not possibly justify the expense of their recovery, still the slag must be treated in a lead circuit and the elements recovered because of the contamination of silver and gold from the slimes from which they are removed.)

To resume with the treatment of the slimes, after the “litharge slag” is removed by cupellation it is usual to further oxidize the now rich silver alloy with caustic and niter to remove selenium and tellurium as sodium selenite and sodium tellurate. These soda slags are readily water soluble and solutions prepared from these slags can be purified and the tellurium separated from the selenium and eventually commercial selenium and tellurium prepared for the market. These steps involve complex chemistry and chemical engineering processes that could well occupy a complete lecture such as this.

After the removal of the selenium and tellurium from the silver-rich alloy (dore) it is cast into small anodes and is electrolyzed to separate the silver from the more noble elements—gold and the platinum metals. The cathodically deposited crystalline silver, after suitable washing, is readily melted and cast into pure silver bars for the market or Mint.

The slimes from the silver electrolysis or parting plant are again melted and cast into still smaller and very valuable anodes for another electrolytic step whereby the gold is recovered as the cathode; and, from the solution and the slimes, the minute amounts of platinum metals are recovered. The amount of platinum metals in most concentrates is so small that its detection by conventional analytical or instrumental means is impossible. Only by subjecting selected, large samples to steps equivalent to the entire smelting and refining process can a quantitative determination of these elements be made. This “analytical” procedure is not usually performed for it would cost more in most cases than the total value of the platinum metals recovered.

Copper concentrates containing recoverable amounts of cobalt are not smelted by the conventional matte process but are roasted for complete sulfur removal and then directly smelted to metal. This metal is separated into two layers: the top “white metal” layer, rich in cobalt, is treated primarily for cobalt recovery and usually is added to a circuit where other cobalt concentrates are treated. The bottom copper layer, after oxidizing to strip the last of the cobalt into a return slag, is essentially blister copper and is refined by the usual procedures already outlined.

Another modification of the copper smelting circuit is practiced by the International Nickel Co. of Canada where the concentrates are particularly rich in nickel. In this process the sulfide concentrates are melted to a liquid matte, separating the gangue in a slag, and the matte is very slowly cooled permitting separate crystallization of the nickel and copper sulfides. This matte is then crushed, ground and subjected to a closed circuit flotation separating a rather pure nickel sulfide from a less pure copper sulfide which, however, can be treated by conventional smelting and refining procedures. It is believed that this process is unique, in that a mechanical method of separating elements is used after the smelting procedure has started. In effect it converts badly dispersed mineral constituents into a form that permits satisfactory separation early in the process avoiding much more complicated chemical procedures.

We have now traced the elements through the copper circuit and, in addition to copper, have found it necessary to produce marketable products of sulfur, nickel, arsenic, selenium, tellurium, silver, gold, and the platinum metals. We have diverted from the copper circuit for further treatment lead, antimony, bismuth, cobalt, zinc, cadmium, thallium and, incidentally, more silver than we like. In addition, one producer of copper treats his reverberatory furnace slag to recover zinc oxide fume which is treated in a zinc plant.


This illustrates the simple steps which would be used to produce lead from a pure lead sulfide concentrate associated only with gangue materials that would enter the slag. However, in actual practice the process is not quite so simple. In sintering, the flue dust is a simple circulating load and the gases may be treated for recovery of sulfuric acid before venting; but the fume may be either circulated back to the sintering machine in which case the elements which concentrate in the fume will build up until the quantity which exits in the sinter will equal the intake, minus the handling losses. In some cases, particularly where there is a high concentration in the fume of one or more elements either objectionable in the lead circuit or valuable in some other circuit, such as selenium, indium, and thallium, the fume is reconcentrated and shipped to another circuit.

In smelting the sinter, the major separation of gangue material is made. The fume from the lead blast furnace concentrates zinc, cadmium, and thallium and, upon circulation, the cadmium, if present in appreciable amounts on the charge, will concentrate sufficiently to justify a separate roasting step which will result in a fume sufficiently rich to feed directly to a cadmium recovery plant. This roasting step will result in the retention of most of the zinc, lead, and antimony in the residue which circulates back to the lead charge while the thallium and part of the indium and selenium concentrate with the cadmium and must be separated in the cadmium plant. Chlorine and fluorine also concentrate in this fume and further complicate its treatment. The slag from the blast furnace consists of oxides of iron, calcium, silicon, aluminum, magnesium, and manganese but contains, in addition, significant concentrations of the zinc, chlorine, fluorine, germanium and arsenic and important amounts of lead, antimony, cadmium, copper and silver. When the slag is charged to a fuming furnace, Fig. 15, most of the zinc, germanium, lead, cadmium, chlorine and fluorine report in the fume. A deleading operation on the fume returns the major amounts of all recovered elements, except the zinc, in a lead-rich fume to the lead sintering machines. In one plant, at least, a matte is also separated from the slag after the fuming operation and this step recovers substantial amounts of copper and silver from the dezinced slag. The recovery of part of the chlorine and fluorine into the fume is quite unwelcome and may necessitate a leaching operation in the lead circuit to prevent these objectionable elements from concentrating.

We have now dealt with the separation of sixteen elements and have not even followed the main flow of the lead. This liquid metal, as it flows from the blast furnace, Fig. 16, may contain: copper, silver, gold, arsenic, antimony, bismuth, indium, nickel, iron, tellurium, cadmium, zinc, sulfur and possibly other elements. While these elements may all be dissolved in the lead at the high temperature at which the lead, of necessity, is tapped from the blast furnace, upon cooling many of them separate from the lead in the form of a dross. It is true that if the charge to the blast furnace contains excessive amounts of arsenic instead of obtaining just lead and slag from the furnace, an additional liquid may be tapped called “blast furnace speiss”. Also, if the charge is high in sulfur there is some question whether the heavy layer tapped should be called matte or lead and if both arsenic and sulfur are high, a heterogeneous mixture of lead, speiss, and matte may be obtained which creates additional byproducts demanding special treatment. However, such practice is carefully avoided in most smelters and no effort will be made to follow the course of byproduct elements through such diversified and usually unnecessary circuits.

To resume the study of the material flow in the liquid lead: upon cooling to the melting point, the dross removed, concentrates the copper, sulfur, arsenic, and indium but, unfortunately, also contains antimony, zinc, cadmium, lead, selenium, silver, and gold. It is usually necessary to add sulfur to the lead toward the end of the drossing operation to secure the maximum removal of copper into the dross for it is through this dross that the copper is diverted from the lead to the copper circuit.

This dross, may be treated in either one of two processes: the silica-matte reverberatory smelting process or the soda-matte process. It is the author’s contention that if the arsenic is high in the dross or if the freight on the copper by-product is high, then only the soda matte process should be used. Since the writer’s knowledge of the silica-matte process is limited, particularly with regard to the flow of by-product elements, only the soda-matte process will be followed in this paper.

In the soda-matte process, the dross, with the addition of only a small amount of soda ash and still smaller amount of reducer to balance oxidation, is charged into a deep bath reverberatory furnace where a steep thermal gradient permits the separation of three layers: a bottom lead layer which is circulated back to the drossing kettle; an intermediate speiss layer which permits a very favorable distribution coefficient between lead and copper; and a top matte layer which serves as an outlet for the sulfur content of the dross and which also has a very favorable lead to copper distribution because of the intervening speiss layer. Both the speiss and the matte layer may be shipped to a copper smelter and carry from the lead circuit most of the copper, arsenic, sulfur, selenium, tellurium, nickel and indium. In addition, unfortunately, they carry antimony, silver, gold, bismuth, zinc, and lead.

It should be noted that the lead in the dross is very largely liquated from the higher melting elements and compounds in the dross reverberatory furnace and goes back to the drossing kettles for retreatment. In the case of a high-copper charge on the blast furnace because of this circulation it is possible to have the dross reverberatory furnace smelt more tons of dross than there are tons of lead plus copper on the charge and clearly shows one reason why a lead blast furnace is an undesirable smelting circuit for copper.

The lead bullion, after decopperizing in the drossing kettle, is transferred to the lead refinery circuit, as shown in Fig. 18, all of which (as in the case of copper) is made necessary by the presence of impurities (valuable or not) in the original smelter charge. The first step in lead refining (if the copper has been satisfactorily removed in the drossing circuit) is softening. This is accomplished by oxidation of the arsenic, antimony, and tin into a litharge slag in a reverb furnace or, as in the Harris Process, in a caustic soda slag in heated, steel kettles. In some cases, the softening can be done step-wise to effect a concentration of the arsenic, antimony and tin into separate slags. These slags are, after suitable treatment, smelted to antimony-lead or antimony-tin-lead alloys or are separately smelted to antimony metal and tin metal. In some cases the tin may be marketed as SnOz or SnCl4 and a very considerable tonnage of antimony is marketed asSb203. In all cases there are several complex refining steps which the material must go through to effect a suitably pure, refined metal, alloy or compound for the market. After the bullion is softened it is desilverized by adding zinc to the molten metal and separating a dross which contains all the noble metals as well as any copper or nickel which is present. This dross is then retorted to recover the zinc content for reuse. The residual metal is a lead alloy very rich in silver and gold which is cupelled as in the copper slimes treatment circuit which it frequently joins at this point. In the cupel the lead is oxidized to litharge in which is dissolved all the other less noble impurities and, unfortunately, distressing amounts of silver and gold requiring very careful handling and smelting of this circulating load. The treatment of the noble metals is the same as in the copper circuit.

After the lead bullion has been desilverized, residual zinc must be removed and recovered by one of three processes: by vacuum dezincing, which permits its reuse in desilverizing; by chlorine dezincing, resulting in a marketable ZnC2; or by oxidation, diverting the zinc back to a zinc smelter. Bullion, high in bismuth, must be debismuthized before the lead can be marketed. This is done either by the addition of calcium and magnesium or potassium which results in the formation of a dross in which the bismuth is concentrated and from which pure bismuth is eventually made after many concentrating and purifying steps. The residual calcium and magnesium or potassium are recovered from the lead as a slag which is returned to the process of producing the reagent alloys. The lead is now refined and, after a final clean-up step with a molten caustic slag, is cast into marketable form. It is true that a considerable tonnage of bullion is processed by the Betts’ electrolytic process, and results in a simpler flowsheet for the main body of lead. However, the impurities in the bullion concentrate in the slimes and must be separated and recovered by procedures that are essentially the same as those employed in the “kettle” refining process and involve an equal number of processing steps.

We have now followed the elements through the lead smelting and refining circuits and find that to recover pure lead we have dealt with nineteen elements and produced from four to eight marketable products without mentioning alloys. We have diverted speiss and matte to the copper smelting circuit; have joined with a copper refinery circuit for the most economical recovery of noble metals; and have diverted zinc, cadmium, indium, thallium, germanium, selenium, and tellurium into suitable circuits for their ultimate recovery.