The series arrangement has the advantage of requiring electrical connections to be made at the first and last plates only, whereas the parallel system requires a connection at every plate; but in the series system the leakage of current due to the short-circuiting action of the sediment and sides of the tank is from 10 to 20 per cent, so that the parallel system is more generally used. The connections between the various plates and the circuit in the parallel systems are made by copper rods, which are run at two different levels along the edges of the tanks, one bar for each set of plates. In some instances these rods are of the inverted V shape, so that the edges will cut through any corrosion which may happen to form at the points of contact. The vats are arranged, with respect to each other, so that each is accessible from all sides and free circulation of the elec trolyte is possible. This circulation is some times obtained by blowing a stream of air through the electrolyte, but more frequently by arranging the vats in steps, and piping so that the electrolyte may pass from the top of one vat to the bottom of the next, by the action of gravity. This maintains a uniform density of electrolyte, which is necessary for the proper formation of the deposit. The electromotive force required is from 0.2 to 0.4 volt per tank, with a current density of 15 to 20 amperes per square foot of cathode plate surface. The individual vats are connected in series so that the total voltage may be approximately the same as that which the generator furnishes, being usually 110 volts. One ampere of current de posits on the cathode only about one ounce of refined copper in 24 hours, and the current density must be kept below 40 amperes per square foot to avoid, mushrooming and conse short-circuiting. In practice from 400 to ampere-hours are required per pound of copper deposited, the theoretical amount ac cording to Faraday's law being only 386.2 am pere-hours. The loss varies from 4 to 20 per cent according to the system employed.
The main product of refining is commercial cathodes, which are sometimes shipped to con sumers, but more frequently cast into wire-bars, ingots, cakes or slabs of standard dimensions and weight. They usually assay from 99.86 to 99.94 per cent pure copper. The yield in com mercial cathodes is from 97 to 99 per cent of the anodes treated, excluding the anode scrap which varies in weight from 7 to 15 per cent of the original anode in parallel plants, but this scrap is not a loss as it is collected and recast into anode plates. Besides electrolytic copper most plants secure gold, silver, platinum and palladium from the slimes, and sometimes selenium, tellurium and other rarer metals. Nickel salts are usually recovered from the solutions.
There are in the United States 10 electrolytic copper refineries with a total capacity of 2,780,000,000 pounds per year; one refinery in Canada with a capacity of 14,000,000 pounds per year. The actual production in 1917 was about 2,300,000,000 pounds, representing approxi mately 74 per cent of the entire world's produc tion of copper for the year. Or, deducting from the total production the amount that does not require refining, about 275,0001300 pounds from Michigan, the United States proddction amounts to over 81 per cent of the total pro duction of refined copper. The other 19 per cent is produced in a number of plants of com paratively small capacity in England, Wales and Continental Europe.
Aluminum.— Practically the whole output of this metal for the entire world is now pro duced electrolytically. The only process used on a large scale is that invented independently in 1886 by Charles M. Hall in the United States
and by Paul L. T. Heroult in France. This process consists in electrolyzing alumina dis solved in a fused bath of cryolite. The alum ina is obtained from the mineral bauxite which occurs abundantly in Arkansas, Georgia, Ala bama and Tennessee. The natural material, being a hydrated alumina containing silica, iron and titanium, must be treated in order to drive off the water and eliminate the impurities. This is accomplished by a chemical process. In practice it requires about two pounds of alum ina for each pound of aluminum produced. The flux or bath in which the alumina is dissolved consists of cryolite, a natural double fluoride of aluminum and sodium (A1Fa.6NaF) found in Greenland. This is melted in a large carbon lined, sheet-iron tank which constitutes the negative electrode, a group of suspended carbon rods forming the positive electrode. A current of several thousand amperes at six to seven volts is used. Only a portion of this voltage is required to decompose the alumina, the bal ance amounting to about four to five volts rep resents the heat required to keep the bath melted. The passage of the current causes the aluminum to deposit on the bottom of the tank as a fused metal, whence it is drawn off period ically. The oxygen set free combines with the carbon of the positive electrodes and passes off as carbonic oxide. The reaction is 2A1-1-3CO. About one pound of carbon is con sumed for one pound of aluminum produced. An excess of alumina is kept floating on the bath so that it is saturated at all times. Ac cording to Faraday's law the weight of alumi num deposited by 1,000 amperes is 0.743 pound per hour. The actual yield of metal by the Hall process is about 85 per cent of this theo retical amount. The metal when drawn from the tanks is cast into rough ingots which are afterward remelted and converted into com mercial shapes, such as sheets, rods, wires, etc. The United States in 1917 produced about 180,000,000 pounds of aluminum, which was about two-thirds the total production of the 4orld. Before the European War the share of the United States in the total production was tinder 50 per cent.
Zinc.— The very high temperature (1300° C. or 2370° F.) necessary to reduce zinc from its common ores, and its generation as a vapor, due to its boiling point being at 930° C. (1700° F.) present difficulties which offer an unusually open field for success to an electrolytic method of reduction. Several processes. are in use. They all provide for the preliminary roasting of the ore— which is essentially lead sulphide, zinc sulphide and gangue—at a low red heat, so as to convert the sulphides into oxides and sulphates. The roasted ore is then treated with dilute sulphuric acid, the zinc being dissolved as sulphate, leaving the lead sulphate as an in soluble residue to be smzIted by the usual dry methods. Most of the silver present remains with the lead, a small portion passing into solu tion with the zinc. It is in fact the recovered silver that sometimes makes the process profit-, able. It is necessary to free the zinc solution from iron, copper and other foreign metals—a matter of considerable difficulty. When suf ficiently purified, the zinc sulphate is electro lyzed, the anodes being of lead, and the cathodes thin sheets of zinc. The operation is 'in reality a reduction of the sulphate, in no sense a refining process. As the reduction proceeds the electrolyte becomes more and more acid, and when hydrogen in quantity is evolved at the cathodes, the electrolyte is run off, and used again to leach roasted ore.