The zinc circuit is in some respects simpler than the lead and copper circuits since only ten recoverable elements are encountered (zinc, cadmium, lead, copper, silver, gold, germanium, indium, mercury, and sulfur) and, in all but one process (the blast furnace) the nonrecoverable elements are left in a heterogeneous residue that is either discarded or shipped to a lead or copper smelter for recovery of values and the elimination of unwanted elements in a waste slag.
In addition, much zinc of marketable grade is produced directly from the extraction process without resorting to additional refining steps after the primary separation of zinc from the gangue. Even when subsequent refining steps are used they are always carried out in the original smelting plant and the “refinery” is not divorced from the metal smelting process in the way that is most usual with copper and lead.
However, the metallurgy of zinc itself, in many respects, is more difficult than the metallurgy of either copper or lead and has resulted in five different rival processes for recovering zinc, all of which are operating with economic justification at the present time.
Fig. 4 illustrates the basic retort process which involves roasting to remove the sulfur, retorting to reduce and vaporize the zinc, and condensation to recover the zinc as a liquid which can be cast into a marketable shape. The transfer of the zinc by vaporization from the charge to the condenser results in a separation of zinc from impurities which are associated with its concentrates, which is quite remarkable, and the condensed metal is usually seriously contaminated with lead and cadmium only and even these can be controlled in the roasting process within suitable limits for most applications without resorting to refining steps on the condensed metal. Where Special High-Grade zinc is required, retort zinc can be fractionally distilled to separate lead, indium, cadmium and iron to produce a zinc of practically any desired degree of purity.
There are four different retort processes which start with a roasted and sintered concentrate, Fig. 20. In the roasting operation (which is frequently combined with sintering), in addition to eliminating sulfur as the dioxide, lead can be lowered enough to permit adequate control in the final zinc for all but the highest grades. In addition, fluorine, chlorine, and mercury can, happily, be eliminated.
In the sintering process, in addition to removing the last of the sulfur, additional lead, and any fluorine, chlorine, mercury, and so forth, which escaped roasting or were circulated back, a substantial elimination of the cadmium can be effected and the fume can be concentrated sufficiently in this element to be fed directly into a cadmium recovery plant.
Of the four retorting processes, two (the horizontal and vertical retort process) produce zinc vapor by heating the sinter plus carbon in a closed container. In both processes a relatively tight, but thin-walled, container must be used for while the heat must be driven through the wall of the retort, the zinc vapor must not be contaminated with the gaseous products of combustion of the heating fuel. Thus, the zinc vapor is diluted essentially only by the carbon monoxide resulting from the carbon reducing the zinc oxide. In this respect the electro-thermic process is quite similar, for here the heat is generated by the electric resistance of the charge, thus avoiding a thin-walled retort, but nevertheless, creating a zinc vapor mixed only with the carbon monoxide generated by the reduction of the zinc oxide with carbon. In the horizontal retort process the zinc vapor is generated in many small retorts and is condensed in an equal number of small refractory condensers from which it is laboriously tapped by hand. A recent modification of the process has successfully employed a single, large condenser for a bank of retorts. In this modified horizontal retort process, in the vertical retort process, and in the electrothermal process each furnace produces a large, continuous supply of vapor necessitating the development of quite large condensers for continuous operation.
In the three retort processes so far discussed the residue after the zinc is distilled is withdrawn as a solid and may contain, in addition to gangue, zinc, copper, lead, silver, gold, and germanium and is usually shipped to a lead plant for recovery of these elements.
The fourth distillation process for zinc, is the relatively new blast-furnace process, developed in England since World War II. In this process the sinter is smelted in a special hot top, blast furnace and the gases, although relatively poor in zinc (as compared to other reduction and volatilization processes) are successfully condensed in a large circulating bath of lead which is sprayed vigorously into the gas stream. The circulating lead is subsequently continuously cooled and the zinc continuously liquated from the bath in a nice thermally balanced operation. The waste gases from the condenser are used as a fuel to preheat the charge, including the air blown into the blast furnace. The elements charged to the blast furnace, other than the zinc, cadmium, and part of the lead, are discharged from the blast furnace as a liquid which is separated into a metal or matte and a slag. The slag is usually low enough in values to be sent to waste. The metal or matte is most advantageously sent to a lead plant for the recovery of lead, silver, gold, copper, and, possibly, additional elements. Thus in spite of the basic differences in gangue behavior between the blast furnace and the retort processes the separation step of both processes involves the volatilization of the zinc and cadmium from the other elements and, hence, the recovery of the residual elements is essentially the same.
In all four retorting processes if Special High-Grade zinc is desired it is usually necessary to redistill and fractionate the zinc to obtain the desired purity. From this process a lead alloy containing indium is obtained by liquation from the lead-rich fraction. From the cadmium fraction commercial cadmium is readily concentrated.
The electrolytic process for the recovery of zinc, being a hydrometallurgical process as contrasted to a pyrometallurgical process, results in a basically different separation of zinc from the elements associated with it in zinc concentrates. The sulfur is separated from the zinc in a roasting operation and recovered, usually, as acid, just as in the pyrometallurgical processes. In the next step (a sulfuric acid leach) the zinc is almost completely separated from: the associated lead and all the gangue material, including most of the iron which remains as an insoluble residue and which, quite like the retort residue, is usually shipped to a lead plant for the recovery of lead and other values - principally silver. The impurities which dissolve with zinc must be practically completely removed from the solution before it can be subjected to electrolysis for the production of commercially pure zinc cathodes. Impurities in the electrolyte, in amounts that completely defied detection a few years ago, are sufficient to prevent the efficient electrolysis of the zinc and, in certain cases, may even completely prevent the extraction of any zinc by electrolysis of the solution. In the electrolytic process for zinc we are compelled to remove impurities to the ultimate degree without regard for the economic values of the elements removed. From the solution, copper, nickel, cobalt, arsenic, antimony, germanium, andiron are removed as precipitates that are shipped to either copper or lead smelters for the recovery of at least the copper and germanium and occasionally other elements. Cadmium is usually recovered as a separate precipitate which is treated in an electrolytic cadmium recovery plant that is usually operated as part of the electrolytic zinc plant. The elimination of the chlorine and fluorine must be carefully controlled prior to the leaching process or uneconomic discard of electrolyte will be required to keep these elements at acceptable levels. The level of sodium salts in the electrolyte must be regulated and some concentrates cannot be treated due to excessive magnesium content which can be controlled in the electrolyte only at a relatively high cost. Calcium is also an impurity which must be managed expertly if expensive cleaning and maintenance of equipment is to be avoided. Manganese is perhaps the only welcome impurity in the electrolytic zinc process, for if the concentrates do not contain enough of this element it is actually purchased and added to the leach process.
While the recovery of contaminating elements in the electrolytic zinc process is, as in the other zinc processes, less important economically than in the metallurgy of lead and copper, the control of impurities in the electrolyte is of greatest importance and requires the greatest skills of the chemical engineer to effect the refining steps on bulk solutions at a rate and cost which can be tolerated. The analytical methods to monitor these solutions have only recently caught up with the requirements of the operators who, all too frequently, had to “operate in the dark” and, frequently, in the past did not know just which of several suspect elements was causing low current efficiency in the electrolysis plant.
To summarize: In recovering zinc from its concentrates one has to deal with twenty-two elements (excluding carbon and oxygen) of which at least three (zinc, cadmium and sulfur) are recovered in marketable form, at least six are shunted to a lead smelter for either direct recovery or further shunting to a copper smelter for eventual recovery via a copper refinery. All of the by-products usually contain enough zinc and, frequently, cadmium and, on occasion, germanium to justify circulating them to a lead plant where zinc, and other elements recovered from the slag, are circulated back to the zinc plant.
The copper, lead, and zinc circuits are combined to show only the principal points of exit of the various recovered elements. There is no recirculation of recovered elements indicated, no indication of the flow of unrecovered elements, nor even an outline of the steps needed to recover the “impurities”—silver, gold, platinum, palladium, sulfur, selenium, tellurium, nickel, cobalt, arsenic, antimony, bismuth, tin, cadmium, indium, thallium, germanium, and mercury after they have been separated from the main flow of copper, lead and zinc. The author contends that this is, indeed—“The World’s Most Complex Metallurgy”.