1911 Encyclopædia Britannica/Copper

COPPER (symbol Cu, atomic weight 63.1, H = 1, or 63.6, O = 16), a metal which has been known to and used by the human race from the most remote periods. Its alloy with tin (bronze) was the first metallic compound in common use by mankind, and so extensive and characteristic was its employment in prehistoric times that the epoch is known as the Bronze Age. By the Greeks and Romans both the metal and its alloys were indifferently known as χαλκός and aes. As, according to Pliny, the Roman supply was chiefly drawn from Cyprus, it came to be termed aes cyprium, which was gradually shortened to cyprium, and corrupted into cuprum, whence comes the English word copper, the French cuivre, and the German Kupfer.

Copper is a brilliant metal of a peculiar red colour which assumes a pinkish or yellowish tinge on a freshly fractured surface of the pure metal, and is purplish when the metal contains cuprous oxide. Its specific gravity varies between 8.91 and 8.95, according to the treatment to which it may have been subjected; J. F. W. Hampe gives 8.945 (0°/4°) for perfectly pure and compact copper. Ordinary commercial copper is somewhat porous and has a specific gravity ranging from 8.2 to 8.5. It takes a brilliant polish, is in a high degree malleable and ductile, and in tenacity it only falls short of iron, exceeding in that quality both silver and gold. By different authorities its melting-point is stated at from 1000° to 1200° C.; C. T. Heycock and F. H. Neville give 1080°.5; P. Dejean gives 1085° as the freezing-point. The molten metal is sea-green in colour, and at higher temperatures (in the electric arc) it vaporizes and burns with a green flame. G. W. A. Kahlbaum succeeded in subliming the metal in a vacuum, and H. Moissan (Compt. rend., 1905, 141, p. 853) distilled it in the electric furnace. Molten copper absorbs carbon monoxide, hydrogen and sulphur dioxide; it also appears to decompose hydrocarbons (methane, ethane), absorbing the hydrogen and the carbon separating out. These occluded gases are all liberated when the copper cools, and so give rise to porous castings, unless special precautions are taken. The gases are also expelled from the molten metal by lead, carbon dioxide, or water vapour. Its specific heat is 0.0899 at 0° C. and 0.0942 at 100°; the coefficient of linear expansion per 1° C. is 0.001869. In electric conductivity it stands next to silver; the conducting power of silver being equal to 100, that of perfectly pure copper is given by A. Matthiessen as 96.4 at 13° C.

Copper is not affected by exposure in dry air, but in a moist atmosphere, containing carbonic acid, it becomes coated with a green basic carbonate. When heated or rubbed it emits a peculiar disagreeable odour. Sulphuric and hydrochloric acids have little or no action upon it at ordinary temperatures, even when in a fine state of division; but on heating, copper sulphate and sulphur dioxide are formed in the first case, and cuprous chloride and hydrogen in the second. Concentrated nitric acid has also very little action, but with the dilute acid a vigorous action ensues. The first products of this reaction are copper nitrate and nitric oxide, but, as the concentration of the copper nitrate increases, nitrous oxide and, eventually, free nitrogen are liberated.

Many colloidal solutions of copper have been obtained. A reddish-brown solution is obtained from solutions of copper chloride, stannous chloride and an alkaline tartrate (Lottermoser, Anorganische Colloïde, 1901).

Occurrence.—Copper is widely distributed in nature, occurring in most soils, ferruginous mineral waters, and ores. It has been discovered in seaweed; in the blood of certain Cephalopoda and Ascidia as haemocyanin, a substance resembling the ferruginous haemoglobin, and of a species of Limulus; in straw, hay, eggs, cheese, meat, and other food-stuffs; in the liver and kidneys, and, in traces, in the blood of man and other animals (as an entirely adventitious constituent, however); it has also been shown by A. H. Church to exist to the extent of 5.9% in turacin, the colouring-matter of the wing-feathers of the Turaco.

Native copper, sometimes termed by miners malleable or virgin copper, occurs as a mineral having all the properties of the smelted metal. It crystallizes in the cubic system, but the crystals are often flattened, elongated, rounded or otherwise distorted. Twins are common. Usually the metal is arborescent, dendritic, filiform, moss-like or laminar. Native copper is found in most copper-mines, usually in the upper workings, where the deposit has been exposed to atmospheric influences. The metal seems to have been reduced from solutions of its salts, and deposits may be formed around mine-timber or on iron objects. It often fills cracks and fissures in the rock. It is not infrequently found in serpentine, and in basic eruptive rocks, where it occurs as veins and in amygdales. The largest known deposits are those in the Lake Superior region, near Keweenaw Point, Michigan, where masses upwards of 400 tons in weight have been discovered. The metal was formerly worked by the Indians for implements and ornaments. It occurs in a series of amygdaloidal dolerites or diabases, and in the associated sandstones and conglomerates. Native silver occurs with the copper, in some cases embedded in it, like crystals in a porphyry. The copper is also accompanied by epidote, calcite, prehnite, analcite and other zeolitic minerals. Pseudomorphs after calcite are known; and it is notable that native copper occurs pseudomorphous after aragonite at Corocoro, in Bolivia, where the copper is disseminated through sandstone.

Ores.—The principal ores of copper are the oxides cuprite and melaconite, the carbonates malachite and chessylite, the basic chloride atacamite, the silicate chrysocolla, the sulphides chalcocite, chalcopyrite, erubescite and tetrahedrite. Cuprite (q.v.) occurs in most cupriferous mines, but never by itself in large quantities. Melaconite (q.v.) was formerly largely worked in the Lake Superior region, and is abundant in some of the mines of Tennessee and the Mississippi valley. Malachite is a valuable ore containing about 56% of the metal; it is obtained in very large quantities from South Australia, Siberia and other localities. Frequently intermixed with the green malachite is the blue carbonate chessylite or azurite (q.v.), an ore containing when pure 55.16% of the metal. Atacamite (q.v.) occurs chiefly in Chile and Peru. Chrysocolla (q.v.) contains in the pure state 30% of the metal; it is an abundant ore in Chile, Wisconsin and Missouri. The sulphur compounds of copper are, however, the most valuable from the economic point of view. Chalcocite, redruthite, copper-glance (q.v.) or vitreous copper (Cu₂S) contains about 80% of copper. Copper pyrites, or chalcopyrite, contains 34.6% of copper when pure; but many of the ores, such as those worked specially by wet processes on account of the presence of a large proportion of iron sulphide, contain less than 5% of copper. Cornish ores are almost entirely pyritic; and indeed it is from such ores that by far the largest proportion of copper is extracted throughout the world. In Cornwall copper lodes usually run east and west. They occur both in the “killas” or clay-slate, and in the “growan” or granite. Erubescite (q.v.), bornite, or horseflesh ore is much richer in copper than the ordinary pyrites, and contains 56 or 57% of copper. Tetrahedrite (q.v.), fahlerz, or grey copper, contains from 30 to 48% of copper, with arsenic, antimony, iron and sometimes zinc, silver or mercury. Other copper minerals are percylite (PbCuCl₂(OH)₂), boleite (3PbCuCl₂(OH)₂, AgCl), stromeyerite {(Cu, Ag)₂S}, cubanite (CuS, Fe₂S₃), stannite (Cu₂S, FeSnS₃), tennantite (3Cu₂S, As₂S₃), emplectite (Cu₂S, Bi₂S₃), wolfsbergite (Cu₂S, Sb₂S₃), famatinite (3Cu₂S, Sb₂S₅) and enargite (3Cu₂S, As₂S₅). For other minerals, see Compounds of Copper below.

Metallurgy.—Copper is obtained from its ores by three principal methods, which may be denominated—(1) the pyro-metallurgical or dry method, (2) the hydro-metallurgical or wet method, and (3) the electro-metallurgical method.

The methods of working vary according to the nature of the ores treated and local circumstances. The dry method, or ordinary smelting, cannot be profitably practised with ores containing less than 4% of copper, for which and for still poorer ores the wet process is preferred.

Copper Smelting.—We shall first give the general principles which underlie the methods for the dry extraction of copper, and then proceed to a more detailed discussion of the plant used. Since all sulphuretted copper ores (and these are of the most economic importance) are invariably contaminated with arsenic and antimony, it is necessary to eliminate these impurities, as far as possible, at a very early stage. This is effected by calcination or roasting. The roasted ore is then smelted to a mixture of copper and iron sulphides, known as copper “matte” or “coarse-metal,” which contains little or no arsenic, antimony or silica. The coarse-metal is now smelted, with coke and siliceous fluxes (in order to slag off the iron), and the product, consisting of an impure copper sulphide, is variously known as “blue-metal,” when more or less iron is still present, “pimple-metal,” when free copper and more or less copper oxide is present, or “fine” or “white-metal,” which is a fairly pure copper sulphide, containing about 75% of the metal. This product is re-smelted to form “coarse-copper,” containing about 95% of the metal, which is then refined. Roasted ores may be smelted in reverberatory furnaces (English process), or in blast-furnaces (German or Swedish process). The matte is treated either in reverberatory furnaces (English process), in blast furnaces (German process), or in converters (Bessemer process). The “American process” or “Pyritic smelting” consists in the direct smelting of raw ores to matte in blast furnaces. The plant in which the operations are conducted varies in different countries. But though this or that process takes its name from the country in which it has been mainly developed, this does not mean that only that process is there followed.

The “English process” is made up of the following operations: (1) calcination; (2) smelting in reverberatory furnaces to form the matte; (3) roasting the matte; and (4) subsequent smelting in reverberatory furnaces to fine- or white-metal; (5) treating the fine-metal in reverberatory furnaces to coarse- or blister-copper, either with or without previous calcination; (6) refining of the coarse-copper. A shorter process (the so-called “direct process”) converts the fine-metal into refined copper directly. The “Welsh process” closely resembles the English method; the main difference consists in the enrichment of the matte by smelting with the rich copper-bearing slags obtained in subsequent operations. The “German or Swedish process” is characterized by the introduction of blast-furnaces. It is made up of the following operations: (1) calcination, (2) smelting in blast-furnaces to form the matte, (3) roasting the matte, (4) smelting in blast-furnaces with coke and fluxes to “black-” or “coarse-metal,” (5) refining the coarse-metal. The “Anglo-German Process” is a combination of the two preceding, and consists in smelting the calcined ores in shaft furnaces, concentrating the matte in reverberatory furnaces, and smelting to coarse-metal in either.

The impurities contained in coarse-copper are mainly iron, lead, zinc, cobalt, nickel, bismuth, arsenic, antimony, sulphur, selenium and tellurium. These can be eliminated by an oxidizing fusion, and slagging or volatilizing the products resulting from this operation, or by electrolysis (see below). In the process of oxidation, a certain amount of cuprous oxide is always formed, which melts in with the copper and diminishes its softness and tenacity. It is, therefore, necessary to reconvert the oxide into the metal. This is effected by stirring the molten metal with a pole of green wood (“poling”); the products which arise from the combustion and distillation of the wood reduce the oxide to metal, and if the operation be properly conducted “tough-pitch” copper, soft, malleable and exhibiting a lustrous silky fracture, is obtained. The surface of the molten metal is protected from oxidation by a layer of anthracite or charcoal. “Bean-shot” copper is obtained by throwing the molten metal into hot water; if cold water be used, “feathered-shot” copper is formed. “Rosette” copper is obtained as thin plates of a characteristic dark-red colour, by pouring water upon the surface of the molten metal, and removing the crust formed. “Japan” copper is purple-red in colour, and is formed by casting into ingots, weighing from six ounces to a pound, and rapidly cooling by immersion in water. The colour of these two varieties is due to a layer of oxide. “Tile” copper is an impure copper, and is obtained by refining the first tappings. “Best-selected” copper is a purer variety.

Calcination or Roasting and Calcining Furnaces.—The roasting should be conducted so as to eliminate as much of the arsenic and antimony as possible, and to leave just enough sulphur as is necessary to combine with all the copper present when the calcined ore is smelted. The process is effected either in heaps, stalls, shaft furnaces, reverberatory furnaces or muffle furnaces. Stall and heap roasting require considerable time, and can only be economically employed when the loss of the sulphur is of no consequence; they also occupy much space, but they have the advantage of requiring little fuel and handling. Shaft furnaces are in use for ores rich in sulphur, and where it is desirable to convert the waste gases into sulphuric acid. Reverberatory roasting does not admit of the utilization of the waste gases, and requires fine ores and much labour and fuel; it has, however, the advantage of being rapid. Muffle furnaces are suitable for fine ores which are liable to decrepitate or sinter. They involve high cost in fuel and labour, but permit the utilization of the waste gases.

Reverberatory furnaces of three types are employed in calcining copper ores: (1) fixed furnaces, with either hand or mechanical rabbling; (2) furnaces with movable beds; (3) furnaces with rotating working chambers. Hand rabbling in fixed furnaces has been largely superseded by mechanical rabbling. Of mechanically rabbling furnaces we may mention the O’Harra modified by Allen-Brown, the Hixon, the Keller-Gaylord-Cole, the Ropp, the Spence, the Wethey, the Parkes, Pearce’s “Turret” and Brown’s “Horseshoe” furnaces. Blake’s and Brunton’s furnaces are reverberatory furnaces with a movable bed. Furnaces with rotating working chambers admit of continuous working; the fuel and labour costs are both low.

In the White-Howell revolving furnace with lifters—a modification of the Oxland—the ore is fed and discharged in a continuous stream. The Brückner cylinder resembles the Elliot and Russell black ash furnace; its cylinder tapers slightly towards each end, and is generally 18 ft. long by 8 ft. 6 in. in its greatest diameter. Its charge of from 8 to 12 tons of ore or concentrates is slowly agitated at a rate of three revolutions a minute, and in from 24 to 36 hours it is reduced from say 40 or 35% to 7% of sulphur. The ore is under better control than is possible with the continuous feed and discharge, and when sufficiently roasted can be passed red-hot to the reverberatory furnace. These advantages compensate for the wear and tear and the cost of moving the heavy dead-weight.

Shaft calcining furnaces are available for fine ores and permit the recovery of the sulphur. They are square, oblong or circular in section, and the interior is fitted with horizontal or inclined plates or prisms, which regulate the fall of the ore. In the Gerstenhoffer and Hasenclever-Helbig furnaces the fall is retarded by prisms and inclined plates. In other furnaces the ore rests on a series of horizontal plates, and either remains on the same plate throughout the operation (Ollivier and Perret furnace), or is passed from plate to plate by hand (Malétra), or by mechanical means (Spence and M’Dougall).

The M’Dougall furnace is turret-shaped, and consists of a series of circular hearths, on which the ore is agitated by rakes attached to revolving arms and made to fall from hearth to hearth. It has been modified by Herreshoff, who uses a large hollow revolving central shaft cooled by a current of air. The shaft is provided with sockets, into which movable arms with their rakes are readily dropped. The Peter Spence type of calcining furnace has been followed in a large number of inventions. In some the rakes are attached to rigid frames, with a reciprocating motion, in others to cross-bars moved by revolving chains. Some of these furnaces are straight, others circular. Some have only one hearth, others three. This and the previous type of furnace, owing to their large capacity, are at present in greatest favour. The M’Dougall-Herreshoff, working on ores of over 30% of sulphur, requires no fuel; but in furnaces of the reverberatory type fuel must be used, as an excess of air enters through the slotted sides and the hinged doors which open and shut frequently to permit of the passage of the rakes. The consumption of fuel, however, does not exceed 1 of coal to 10 of ore. The quantity of ore which these large furnaces, with a hearth area as great as 2000 ft. and over, will roast varies from 40 to 60 tons a day. Shaft calcining furnaces like the Gerstenhoffer, Hasenclever, and others designed for burning pyrites fines have not found favour in modern copper works.

The Fusion of Ores in Reverberatory and Cupola Furnaces.—After the ore has been partially calcined, it is smelted to extract its earthy matter and to concentrate the copper with part of its iron and sulphur into a matte. In reverberatory furnaces it is smelted by fuel in a fireplace, separate from the ore, and in cupolas the fuel, generally coke, is in direct contact with the ore. When Swansea was the centre of the copper-smelting industry in Europe, many varieties of ores from different mines were smelted in the same furnaces, and the Welsh reverberatory furnaces were used. To-day more than eight-tenths of the copper ores of the world are reduced to impure copper bars or to fine copper at the mines; and where the character of the ore permits, the cupola furnace is found more economical in both fuel and labour than the reverberatory.

The Welsh method finds adherents only in Wales and Chile. In America the usual method is to roast ores or concentrates so that the matte yielded by either the reverberatory or cupola furnace will run from 45 to 50% in copper, and then to transfer to the Bessemer converter, which blows it up to 99%. In Butte, Montana, reverberatories have in the past been preferred to cupola furnaces, as the charge has consisted mainly of fine roasted concentrates; but the cupola is gaining ground there. At the Boston and Great Falls (Montana) works tilting reverberatories, modelled after open hearth steel furnaces, were first erected; but they were found to possess objectionable features. Now both these and the egg-shaped reverberatories are being abandoned for furnaces as long as 43 ft. 6 in. from bridge to bridge and of a width of 15 ft. 9 in. heated by gas, with regenerative checker work at each end, and fed with ore or concentrates, red-hot from the calciners, through a line of hoppers suspended above the roof. Furnaces of this size smelt 200 tons of charge a day. But even when the old type of reverberatory is preferred, as at the Argo works, at Denver, where rich gold- and silver-bearing copper matte is made, the growth of the furnace in size has been steady. Richard Pearce’s reverberatories in 1878 had an area of hearth of 15 ft. by 9 ft. 8 in., and smelted 12 tons of cold charge daily, with a consumption of 1 ton of coal to 2.4 tons of ore. In 1900 the furnaces were 35 ft. by 16 ft., and smelt 50 tons daily of hot ore, with the consumption of 1 ton of coal to 3.7 tons of ore.

The home of cupola smelting was Germany, where it has never ceased to make steady progress. In Mansfeld brick cupola furnaces are without a rival in size, equipment and performance. They are round stacks, designed on the model of iron blast furnaces, 29 ft. high, fed mechanically, and provided with stoves to heat the blast by the furnace gases. The low percentage of sulphur in the roasted ore is little more than enough to produce a matte of 40 to 45%, and therefore the escaping gases are better fitted than those of most copper cupola furnaces for burning in a stove. But as the slag carries on an average 46% of silica, it is only through the utmost skill that it can be made to run as low on an average as 0.3% in copper oxide. As the matte contains on an average 0.2% of silver, it is still treated by the Ziervogel wet method of extraction, the management dreading the loss which might occur in the Bessemer process of concentration, applied as preliminary to electrolytic separation. Blast furnaces of large size, built of brick, have been constructed for treating the richest and more silicious ores of Rio Tinto, and the Rio Tinto Company has introduced converters at the mine. This method of extraction contrasts favourably in time with the leaching process, which is so slow that over 10,000,000 tons of ore are always under treatment on the immense leaching floors of the company’s works in Spain. In the United States the cupola has undergone a radical modification in being built of water-jacketed sections. The first water-jacketed cupola which came into general use was a circular inverted cone, with a slight taper, of 36 inches diameter at the tuyeres, and composed of an outer and an inner metal shell, between which water circulated. As greater size has been demanded, oval and rectangular furnaces—as large as 180 in. by 56 in. at the tuyeres—have been built in sections of cast or sheet iron or steel. A single section can be removed and replaced without entirely emptying the stack, as a shell of congealed slag always coats the inner surface of the jacket. The largest furnaces are those of the Boston & Montana Company at Great Falls, Montana, which have put through 500 tons of charge daily, pouring their melted slag and matte into large wells of 10 ft. in diameter. A combined brick- and water-cooled furnace has been adopted by the Iron Mountain Company at Keswick, Cal., for matte concentration. In it the cooling is effected by water pipes, interposed horizontally between the layers of bricks. The Mt. Lyell smelting works in Tasmania, which are of special interest, will be referred to later. (See Pyritic Smelting below.)

Concentrating Matte to Copper in the Bessemer Converter.—As soon as the pneumatic method of decarburizing pig iron was accepted as practicable, experiments were made with a view to Bessemerizing copper ores and mattes. One of the earliest and most exhaustive series of experiments was made on Rio Tinto ores at the John Brown works by John Hollway, with the aim of both smelting the ore and concentrating the matte in the same furnace, by the heat evolved through the oxidation of their sulphur and iron. Experiments along the same lines were made by Francis Bawden at Rio Tinto and Claude Vautin in Australia. The difficulty of effecting this double object in one operation was so great that in subsequent experiments the aim was merely to concentrate the matte to metallic copper in converters of the Bessemer type. The concentration was effected without any embarrassment till metallic copper commenced to separate and chill in the bottom tuyeres. To meet this obstacle P. Manhès proposed elevated side tuyeres, which could be kept clear by punching through gates in a wind box. His invention was adopted by the Vivians, at the Eguilles works near Sargues, Vaucluse, France, and at Leghorn in Italy. But the greatest expansion of this method has been in the United States, where more than 400,000,000 ℔ of copper are annually made in Bessemer converters. Vessels of several designs are used—some modelled exactly after steel converters, others barrel-shaped, but all with side tuyeres elevated about 10 in. above the level of the bottom lining. Practice, however, in treating copper matte differs essentially from the treatment of pig iron, inasmuch as from 20 to 30% of iron must be eliminated as slag and an equivalent quantity of silica must be supplied. The only practical mode of doing this, as yet devised, is by lining the converter with a silicious mixture. This is so rapidly consumed that the converters must be cooled and partially relined after 3 to 6 charges, dependent on the iron contents of the matte. When available, a silicious rock containing copper or the precious metals is of course preferred to barren lining. The material for lining, and the frequent replacement thereof, constitute the principal expense of the method. The other items of cost are labour, the quantity of which depends on the mechanical appliances provided for handling the converter shells and inserting the lining; and the blast, which in barrel-shaped converters is low and in vertical converters is high, and which varies therefore from 3 to 15 ℔ to the square inch. The quantity of air consumed in a converter which will blow up about 35 tons of matte per day is about 3000 cub. ft. per minute. The operation of raising a charge of 50% matte to copper usually consists of two blows. The first blow occupies about 25 minutes, and oxidizes all but a small quantity of the iron and some of the sulphur, raising the product to white metal. The slag is then poured and skimmed, the blast turned on and converter retilted. During the second blow the sulphur is rapidly oxidized, and the charge reduced to metal of 99% in from 30 to 40 minutes. Little or no slag results from the second blow. That from the first blow contains between 1% and 2% of copper, and is usually poured from ladles operated by an electric crane into a reverberatory, or into the settling well of the cupola. The matte also, in all economically planned works, is conveyed, still molten, by electric cranes from the furnace to the converters. When lead or zinc is not present in notable quantity, the loss of the precious metals by volatilization is slight, but more than 5% of these metals in the matte is prohibitive. Under favourable conditions in the larger works of the United States the cost of converting a 50% matte to metallic copper is generally understood to be only about 5/10 to 6/10 of a cent per ℔. of refined copper.

Pyritic Smelting.—The heat generated by the oxidation of iron and sulphur has always been used to maintain combustion in the kilns or stalls for roasting pyrites. Pyritic smelting is a development of the Russian engineer Semenikov’s treatment (proposed in 1866) of copper matte in a Bessemer converter. Since John Hollway’s and other early experiments of Lawrence Austin and Robert Sticht, no serious attempts have been made to utilize the heat escaping from a converting vessel in smelting ore and matte either in the same apparatus or in a separate furnace. But considerable progress has been made in smelting highly sulphuretted ores by the heat of their own oxidizable constituents. At Tilt Cove, Newfoundland, the Cape Copper Company smelted copper ore, with just the proper proportion of sulphur, iron and silica, successfully without any fuel, when once the initial charge had been fused with coke. The furnaces used were of ordinary design and built of brick. Lump ore alone was fed, and the resulting matte showed a concentration of only 3 into 1. When, however, a hot blast is used on highly sulphuretted copper ores, a concentration of 8 of ore into 1 of matte is obtained, with a consumption of less than one-third the fuel which would be consumed in smelting the charge had the ore been previously calcined. A great impetus to pyritic smelting was given by the investigations of W. L. Austin, of Denver, Colorado, and both at Leadville and Silverton raw ores are successfully smelted with as low a fuel consumption as 3 of coke to 100 of charge.

Two types of pyritic smelting may be distinguished: one, in which the operation is solely sustained by the combustion of the sulphur in the ores, without the assistance of fuel or a hot blast; the other in which the operation is accelerated by fuel, or a hot blast, or both. The largest establishment in which advantage is taken of the self-contained fuel is at the smelting works of the Mt. Lyell Company, Tasmania. There the blast is raised from 600° to 700° F. in stoves heated by extraneous fuel, and the raw ore smelted with only 3% of coke. The ore is a compact iron pyrites containing copper 2.5%, silver 3.83 oz., gold 0.139 oz. It is smelted raw with hot blast in cupola furnaces, the largest being 210 in. by 40 in. The resulting matte runs 25%. This is reconcentrated raw in hot-blast cupolas to 55%, and blown directly into copper in converters. Thus these ores, as heavily charged with sulphur as those of the Rio Tinto, are speedily reduced by three operations and without roasting, with a saving of 97.6% of the copper, 93.2% of the silver and 93.6% of the gold.

Pyritic smelting has met with a varying economic success. According to Herbert Lang, its most prominent chance of success is in localities where fuel is dear, and the ores contain precious metals and sufficient sulphides and arsenides to render profitable dressing unnecessary.

The Nicholls and James Process.—Nicholls and James have applied, very ingeniously, well-known reactions to the refining of copper, raised to the grade of white metal. This process is practised by the Cape Copper and Elliot Metal Company. A portion of the white metal is calcined to such a degree of oxidation that when fused with the unroasted portion, the reaction between the oxygen in the roasted matte and the sulphur in the raw material liberates the metallic copper. The metal is so pure that it can be refined by a continuous operation in the same furnace.

Wet Methods for Copper Extraction.—Wet methods are only employed for low grade ores (under favourable circumstances ore containing from ¼ to 1% of copper has admitted of economic treatment), and for gold and silver bearing metallurgical products.

The fundamental principle consists in getting the ore into a solution, from which the metal can be precipitated. The ores of any economic importance contain the copper either as oxide, carbonate, sulphate or sulphide. These compounds are got into solution either as chlorides or sulphates, and from either of these salts the metal can be readily obtained. Ores in which the copper is present as oxide or carbonate are soluble in sulphuric or hydrochloric acids, ferrous chloride, ferric sulphate, ammoniacal compounds and sodium thiosulphate. Of these solvents, only the first three are of economic importance. The choice of sulphuric or hydrochloric acid depends mainly upon the cost, both acting with about the same rapidity; thus if a Leblanc soda factory is near at hand, then hydrochloric acid would most certainly be employed. Ferrous chloride is not much used; the Douglas-Hunt process uses a mixture of salt and ferrous sulphate which involves the formation of ferrous chloride, and the new Douglas-Hunt process employs sulphuric acid in which ferrous chloride is added after leaching.

Sulphuric acid may be applied as such on the ores placed in lead, brick, or stone chambers; or as a mixture of sulphur dioxide, nitrous fumes (generated from Chile saltpetre and sulphuric acid), and steam, which permeates the ore resting on the false bottom of a brick chamber. When most of the copper has been converted into the sulphate, the ore is lixiviated. Hydrochloric acid is applied in the same way as sulphuric acid; it has certain advantages of which the most important is that it does not admit the formation of basic salts; its chief disadvantage is that it dissolves the oxides of iron, and accordingly must not be used for highly ferriferous ores. The solubility of copper carbonate in ferrous chloride solution was pointed out by Max Schaffner in 1862, and the subsequent recognition of the solubility of the oxide in the same solvent by James Douglas and Sterry Hunt resulted in the “Douglas-Hunt” process for the wet extraction of copper. Ferrous chloride decomposes the copper oxide and carbonate with the formation of cuprous and cupric chlorides (which remain in solution), and the precipitation of ferrous oxide, carbon dioxide being simultaneously liberated from the carbonate. In the original form of the Douglas-Hunt process, ferrous chloride was formed by the interaction of sodium chloride (common salt) with ferrous sulphate (green vitriol), the sodium sulphate formed at the same time being removed by crystallization. The ground ore was stirred with this solution at 70° C. in wooden tubs until all the copper was dissolved. The liquor was then filtered from the iron oxides, and the filtrate treated with scrap iron, which precipitated the copper and reformed ferrous chloride, which could be used in the first stage of the process. The advantage of this method rests chiefly on the small amount of iron required; but its disadvantages are that any silver present in the ores goes into solution, the formation of basic salts, and the difficulty of filtering from the iron oxides. A modification of the method was designed to remedy these defects. The ore is first treated with dilute sulphuric acid, and then ferrous or calcium chloride added, thus forming copper chlorides. If calcium chloride be used the precipitated calcium sulphate must be removed by filtration. Sulphur dioxide is then blown in, and the precipitate is treated with iron, which produces metallic copper, or milk of lime, which produces cuprous oxide. Hot air is blown into the filtrate, which contains ferrous or calcium chlorides, to expel the excess of sulphur dioxide, and the liquid can then be used again. In this process (“new Douglas-Hunt”) there are no iron oxides formed, the silver is not dissolved, and the quantity of iron necessary is relatively small, since all the copper is in the cuprous condition. It is not used in the treatment of ores, but finds application in the case of calcined argentiferous lead and copper mattes.

The precipitation of the copper from the solution, in which it is present as sulphate, or as cuprous and cupric chlorides, is generally effected by metallic iron. Either wrought, pig, iron sponge or iron bars are employed, and it is important to notice that the form in which the copper is precipitated, and also the time taken for the separation, largely depend upon the condition in which the iron is applied. Spongy iron acts most rapidly, and after this follow iron turnings and then sheet clippings. Other precipitants such as sulphuretted hydrogen and solutions of sulphides, which precipitate the copper as sulphides, and milk of lime, which gives copper oxides, have not met with commercial success. When using iron as the precipitant, it is desirable that the solution should be as neutral as possible, and the quantity of ferric salts present should be reduced to a minimum; otherwise, a certain amount of iron would be used up by the free acid and in reducing the ferric salts. Ores in which the copper is present as sulphate are directly lixiviated and treated with iron. Mine waters generally contain the copper in this form, and it is extracted by conducting the waters along troughs fitted with iron gratings.

The wet extraction of metallic copper from ores in which it occurs as the sulphide, may be considered to involve the following operations: (1) conversion of the copper into a soluble form, (2) dissolving out the soluble copper salt, (3) the precipitation of the copper. Copper sulphide may be converted either into the sulphate, which is soluble in water; the oxide, soluble in sulphuric or hydrochloric acid; cupric chloride, soluble in water; or cuprous chloride, which is soluble in solutions of metallic chlorides.

The conversion into sulphate is generally effected by the oxidizing processes of weathering, calcination, heating with iron nitrate or ferric sulphate. It may also be accomplished by calcination with ferrous sulphate, or other easily decomposable sulphates, such as aluminium sulphate. Weathering is a very slow, and, therefore, an expensive process; moreover, the entire conversion is only accomplished after a number of years. Calcination is only advisable for ores which contain relatively much iron pyrites and little copper pyrites. Also, however slowly the calcination may be conducted, there is always more or less copper sulphide left unchanged, and some copper oxide formed. Calcination with ferrous sulphate converts all the copper sulphide into sulphate. Heap roasting has been successfully employed at Agordo, in the Venetian Alps, and at Majdanpek in Servia. Josef Perino’s process, which consists in heating the ore with iron nitrate to 50°–150° C., is said to possess several advantages, but it has not been applied commercially. Ferric sulphate is only used as an auxiliary to the weathering process and in an electrolytic process.

The conversion of the sulphide into oxide is adopted where the Douglas-Hunt process is employed, or where hydrochloric or sulphuric acids are cheap. The calcination is effected in reverberatory furnaces, or in muffle furnaces, if the sulphur is to be recovered. Heap, stall or shaft furnace roasting is not very satisfactory, as it is very difficult to transform all the sulphide into oxide.

The conversion of copper sulphide into the chlorides may be accomplished by calcining with common salt, or by treating the ores with ferrous chloride and hydrochloric acid or with ferric chloride. The dry way is best; the wet way is only employed when fuel is very dear, or when it is absolutely necessary that no noxious vapours should escape into the atmosphere. The dry method consists in an oxidizing roasting of the ores, and a subsequent chloridizing roasting with either common salt or Abraumsalz in reverberatory or muffle furnaces. The bulk of the copper is thus transformed into cupric chloride, little cuprous chloride being obtained. This method had been long proposed by William Longmaid, Max Schaffner, Becchi and Haupt, but was only introduced into England by the labours of William Henderson, J. A. Phillips and others. The wet method is employed at Rio Tinto, the particular variant being known as the “Dötsch” process. This consists in stacking the broken ore in heaps and adding a mixture of sodium sulphate and ferric chloride in the proportions necessary for the entire conversion of the iron into ferric sulphate. The heaps are moistened with ferric chloride solution, and the reaction is maintained by the liquid percolating through the heap. The liquid is run off at the base of the heaps into the precipitating tanks, where the copper is thrown down by means of metallic iron. The ferrous chloride formed at the same time is converted into ferric chloride which can be used to moisten the heaps. This conversion is effected by allowing the ferrous chloride liquors slowly to descend a tower, filled with pieces of wood, coke or quartz, where it meets an ascending current of chlorine.

The sulphate, oxide or chlorides, which are obtained from the sulphuretted ores, are lixiviated and the metal precipitated in the same manner as we have previously described.

The metal so obtained is known as “cement” copper. If it contains more than 55% of copper it is directly refined, while if it contains a lower percentage it is smelted with matte or calcined copper pyrites. The chief impurities are basic salts of iron, free iron, graphite, and sometimes silica, antimony and iron arsenates. Washing removes some of these impurities, but some copper always passes into the slimes. If much carbonaceous matter be present (and this is generally so when iron sponge is used as the precipitant) the crude product is heated to redness in the air; this burns out the carbon, and, at the same time, oxidizes a little of the copper, which must be subsequently reduced. A similar operation is conducted when arsenic is present; basic-lined reverberatory furnaces have been used for the same purpose.

Electrolytic Refining.—The principles have long been known on which is based the electrolytic separation of copper from the certain elements which generally accompany it, whether these, like silver and gold, are valuable, or, like arsenic, antimony, bismuth, selenium and tellurium, are merely impurities. But it was not until the dynamo was improved as a machine for generating large quantities of electricity at a very low cost that the electrolysis of copper could be practised on a commercial scale. To-day, by reason of other uses to which electricity is applied, electrically deposited copper of high conductivity is in ever-increasing demand, and commands a higher price than copper refined by fusion. This increase in value permits of copper with not over £2 or $10 worth of the precious metals being profitably subjected to electrolytic treatment. Thus many million ounces of silver and a great deal of gold are recovered which formerly were lost.

The earliest serious attempt to refine copper industrially was made by G. R. Elkington, whose first patent is dated 1865. He cast crude copper, as obtained from the ore, into plates which were used as anodes, sheets of electro-deposited copper forming the cathodes. Six anodes were suspended, alternately with four cathodes, in a saturated solution of copper sulphate in a cylindrical fire-clay trough, all the anodes being connected in one parallel group, and all the cathodes in another. A hundred or more jars were coupled in series, the cathodes of one to the anodes of the next, and were so arranged that with the aid of side-pipes with leaden connexions and india-rubber joints the electrolyte could, once daily, be made to circulate through them all from the top of one jar to the bottom of the next. The current from a Wilde’s dynamo was passed, apparently with a current density of 5 or 6 amperes per sq. ft., until the anodes were too crippled for further use. The cathodes, when thick enough, were either cast and rolled or sent into the market direct. Silver and other insoluble impurities collected at the bottom of the trough up to the level of the lower side-tube, and were then run off through a plug in the bottom into settling tanks, from which they were removed for metallurgical treatment. The electrolyte was used until the accumulation of iron in it was too great, but was mixed from time to time with a little water acidulated by sulphuric acid. This process is of historic interest, and in principle it is identical with that now used. The modifications introduced have been chiefly in details, in order to economize materials and labour, to ensure purity of product, and to increase the rate of deposition.

The chemistry of the process has been studied by Martin Kiliani (Berg- und Hüttenmännische Zeitung, 1885, p. 249), who found that, using the (low) current-density of 1.8 ampere per sq. ft. of cathode, and an electrolyte containing 1½ ℔ of copper sulphate and ½ ℔ of sulphuric acid per gallon, all the gold, platinum and silver present in the crude copper anode remain as metals, undissolved, in the anode slime or mud, and all the lead remains there as sulphate, formed by the action of the sulphuric acid (or SO₄ ions); he found also that arsenic forms arsenious oxide, which dissolves until the solution is saturated, and then remains in the slime, from which on long standing it gradually dissolves, after conversion by secondary reactions into arsenic oxide; antimony forms a basic sulphate which in part dissolves; bismuth partly dissolves and partly remains, but the dissolved portion tends slowly to separate out as a basic salt which becomes added to the slime; cuprous oxide, sulphide and selenides remain in the slime, and very slowly pass into solution by simple chemical action; tin partly dissolves (but in part separates again as basic salt) and partly remains as basic sulphate and stannic oxide; zinc, iron, nickel and cobalt pass into solution—more readily indeed than does the copper. Of the metals which dissolve, none (except bismuth, which is rarely present in any quantity) deposits at the anode so long as the solution retains its proper proportion of copper and acid, and the current-density is not too great. Neutral solutions are to be avoided because in them silver dissolves from the anode and, being more electro-negative than copper, is deposited at the cathode, while antimony and arsenic are also deposited, imparting a dark colour to the copper. Electrolytic copper should contain at least 99.92% of metallic copper, the balance consisting mainly of oxygen with not more than 0.01% in all of lead, arsenic, antimony, bismuth and silver. Such a degree of purity is, however, unattainable unless the conditions of electrolysis are rigidly adhered to. It should be observed that the free acid is gradually neutralized, partly by chemical action on certain constituents of the slime, partly by local action between different metals of the anode, both of which effect solution independently of the current, and partly by the peroxidation (or aëration) of ferrous sulphate formed from the iron in the anode. At the same time there is a gradual substitution of other metals for copper in the solution, because although copper plus other (more electro-positive) metals are constantly dissolving at the anode, only copper is deposited at the cathode. Hence the composition and acidity of the solution, on which so much depends, must be constantly watched.

The dependence of the mechanical qualities of the copper upon the current-density employed is well known. A very weak current gives a pale and brittle deposit, but as the current-density is increased up to a certain point, the properties of the metal improve; beyond this point they deteriorate, the colour becoming darker and the deposit less coherent, until at last it is dark brown and spongy or pulverulent. The presence of even a small proportion of hydrochloric acid imparts a brown tint to the deposit. Baron H. v. Hübl (Mittheil. des k. k. militär-geograph. Inst., 1886, vol. vi. p. 51) has found that with neutral solutions a 5% solution of copper sulphate gave no good result, while with a 20% solution the best deposit was obtained with a current-density of 28 amperes per sq. ft.; with solutions containing 2% of sulphuric acid, the 5% solution gave good deposits with current-densities of 4 to 7.5 amperes, and the 20% solution with 11.5 to 37 amperes, per sq. ft. The maximum current-densities for a pure acid solution at rest were: for 15% pure copper sulphate solutions 14 to 21 amperes, and for 20% solutions 18.5 to 28 amperes, per sq. ft.; but when the solutions were kept in gentle motion these maxima could be increased to 21-28 and 28-37 amperes per sq. ft. respectively. The necessity for adjusting the current-density to the composition and treatment of the electrolyte is thus apparent. The advantage of keeping the solution in motion is due partly to the renewal of solution thus effected in the neighbourhood of the electrodes, and partly to the neutralization of the tendency of liquids undergoing electrolysis to separate into layers, due to the different specific gravities of the solutions flowing from the opposing electrodes. Such an irregular distribution of the bath, with strong copper sulphate solution from the anode at the bottom and acid solution from the cathode at the top, not only alters the conductivity in different strata and so causes irregular current-distribution, but may lead to the current-density in the upper layers being too great for the proportion of copper there present. Irregular and defective deposits are therefore obtained. Provision for circulation of solution is made in the systems of copper-refining now in use. Henry Wilde, in 1875, in depositing copper on iron printing-rollers, recognized this principle and rotated the rollers during electrolysis, thereby renewing the surfaces of metal and liquid in mutual contact, and imparting sufficient motion to the solution to prevent stratification; as an alternative he imparted motion to the electrolyte by means of propeller blades. Other workers have followed more or less on the same lines; reference may be made to the patents of F. E. and A. S. Elmore, who sought to improve the character of the deposit by burnishing during electrolysis, of E. Dumoulin, and Sherard Cowper-Coles (Engineering Review, 1905, vol. xiii. p. 392), who prefers to rotate the cathode at a speed that maintains a peripheral velocity of at least 1000 ft. per minute. Certain other inventors have applied the same principle in a different way. H. Thofehrn in America and J. C. Graham in England have patented processes by which jets of the electrolyte are caused to impinge with considerable force upon the surface of the cathode, so that the renewal of the liquid at this point takes place very rapidly, and current-densities per sq. ft. of 50 to 100 amperes are recommended by the former, and of 300 amperes by the latter. Graham has described experiments in this direction, using a jet of electrolyte forced (beneath the surface of the bath) through a hole in the anode upon the surface of the cathode. Whilst the jet was playing, a good deposit was formed with so high a current-density as 280 amperes per sq. ft., but if the jet was checked, the deposit (now in a still liquid) was instantaneously ruined. When two or more jets were used side by side the deposit was good opposite the centre of each, but bad at the point where two currents met, because the rate of flow was reduced. By introducing perforated shields of ebonite between the electrodes, so that the full current-density was only attained at the centres of the jets, these ill effects could be prevented. One of the chief troubles met with was the formation of arborescent growths around the edges of the cathode, due to the greater current-density in this region; this, however, was also obviated by the use of screens. By means of a very brisk rotation of cathode, combined with a rapid current of electrolyte, J. W. Swan has succeeded in depositing excellent copper at current-densities exceeding 1000 amperes per sq. ft. The methods by which such results are to be obtained cannot, however, as yet be practised economically on a working scale; one great difficulty in applying them to the refining of metals is that the jets of liquid would be liable to carry with them articles of anode mud, and Swan has shown that the presence of solid particles in the electrolyte is one of the most fruitful causes of the well-known nodular growths on electro-deposited copper. Experiments on a working scale with one of the jet processes in America have, it is reported, been given up after a full trial.

In copper-refining practice, the current-density commonly ranges from 7.5 to 12 or 15, and occasionally to 18, amperes per sq. ft. The electrical pressure required to force a current of this intensity through the solution, and to overcome a certain opposing electromotive force arising from the more electro-negative impurities of the anode, depends upon the composition of the bath and of the anodes, the distance between the electrodes, and the temperature, but under the usual working conditions averages 0.3 volt for every pair of electrodes in series. In nearly all the processes now used, the solution contains about 1½ to 2 ℔ of copper sulphate and from 5 to 10 oz. of sulphuric acid per gallon of water, and the space between the electrodes is from 1½ to 2 in., whilst the total area of cathode surface in each tank may be 200 sq. ft., more or less. The anodes are usually cast copper plates about (say) 3 ft. by 2 ft. by ¾ or 1 in. The cathodes are frequently of electro-deposited copper, deposited to a thickness of about 1/32 in. on black-leaded copper plates, from which they are stripped before use. The tanks are commonly constructed of wood lined with lead, or tarred inside, and are placed in terrace fashion each a little higher than the next in series, to facilitate the flow of solution through them all from a cistern at one end to a well at the other. Gangways are left between adjoining rows of tanks, and an overhead travelling-crane facilitates the removal of the electrodes. The arrangement of the tanks depends largely upon the voltage available from the electric generator selected; commonly they are divided into groups, all the baths in each group being in series. In the huge Anaconda plant, for example, in which 150 tons of refined copper can be produced daily by the Thofehrn multiple system (not the jet system alluded to above), there are 600 tanks about 8¼ ft. by 4½ ft. by 3¼ ft. deep, arranged in three groups of 200 tanks in series. The connexions are made by copper rods, each of which, in length, is twice the width of the tank, with a bayonet-bend in the middle, and serves to support the cathodes in the one and the anodes in the next tank. Self-registering voltmeters indicate at any moment the potential difference in every tank, and therefore give notice of short circuits occurring at any part of the installation. The chief differences between the commercial systems of refining lie in the arrangement of the baths, in the disposition and manner of supporting the electrodes in each, in the method of circulating the solution, and in the current-density employed. The various systems are often classed in two groups, known respectively as the Multiple and Series systems, depending upon the arrangement of the electrodes in each tank. Under the multiple system anodes and cathodes are placed alternately, all the anodes in one tank being connected to one rod, and all the cathodes to another, and the potential difference between the terminals of each tank is that between a single pair of plates. Under the series system only the first anode and the last cathode are connected to the conductors; between these are suspended, isolated from one another, a number of intermediate bi-polar electrode plates of raw copper, each of these plates acting on one side as a cathode, receiving a deposit of copper, and on the other as an anode, passing into solution; the voltage between the terminals of the tank will be as many times as great as that between a single pair of plates as there are spaces between electrodes in the tank. In time the original impure copper of the plates becomes replaced by refined copper, but if the plates are initially very impure and dissolve irregularly, it may happen that much residual scrap may have to be remelted, or that some of the metal may be twice refined, thus involving a waste of energy. Moreover, the high potential difference between the terminals of the series tank introduces a greater danger of short-circuiting through scraps of metal at the bottom of the bath; for this reason, also, lead-lined vats are inadmissible, and tarred slate tanks are often used instead. A valuable comparison of the multiple and series systems has been published by E. Keller (see The Mineral Industry, New York, 1899, vol. vii. p. 229). G. Kroupa has calculated that the cost of refining is 8s. per ton of copper higher under the series than it is under the multiple system; but against this, it must be remembered that the new works of the Baltimore Copper Smelting and Rolling Company, which are as large as those of the Anaconda Copper Mining Company, are using the Hayden process, which is the chief representative of the several series systems. In this system rolled copper anodes are used; these, being purer than many cast anodes, having flat surfaces, and being held in place by guides, dissolve with great regularity and require a space of only 5/8 in. between the electrodes, so that the potential difference between each pair of plates may be reduced to 0.15–0.2 volt.

J. A. W. Borchers, in Germany, and A. E. Schneider and O. Szontagh, in America, have introduced a method of circulating the solution in each vat by forcing air into a vertical pipe communicating between the bottom and top of a tank, with the result that the bubbling of the air upward aspirates solution through the vertical pipe from below, at the same time aërating it, and causing it to overflow into the top of the tank. Obviously this slow circulation has but little effect on the rate at which the copper may be deposited. The electrolyte, when too impure for further use, is commonly recrystallized, or electrolysed with insoluble anodes to recover the copper.

The yield of copper per ampere (in round numbers, 1 oz. of copper per ampere per diem) by Faraday’s law is never attained in practice; and although 98% may with care be obtained, from 94 to 96% represents the more usual current-efficiency. With 100% current-efficiency and a potential difference of 0.3 volt between the electrodes, 1 ℔ of copper should require about 0.154 electrical horse-power hours as the amount of energy to be expended in the tank for its production. In practice the expenditure is somewhat greater than this; in large works the gross horse-power required for the refining itself and for power and lighting in the factory may not exceed 0.19 to 0.2 (or in smaller works 0.25) horse-power hours per pound of copper refined.

Many attempts have been made to use crude sulphide of copper or matte as an anode, and recover the copper at the cathode, the sulphur and other insoluble constituents being left at the anode. The best known of these is the Marchese process, which was tested on a working scale at Genoa and Stolberg in Rhenish Prussia. As the operation proceeded, it was found that the voltage had to be raised until it became prohibitive, while the anodes rapidly became honeycombed through and, crumbling away, filled up the space at the bottom of the vat. The process was abandoned, but in a modified form appears to be now in use in Nijni-Novgorod in Russia. Siemens and Halske introduced a combined process in which the ore, after being part-roasted, is leached by solutions from a previous electrolytic operation, and the resulting copper solution electrolysed. In this process the anode solution had to be kept separate from the cathode solution, and the membrane which had in consequence to be used, was liable to become torn, and so to cause trouble by permitting the two solutions to mix. Modifications of the process have therefore been tried.

Modern methods in copper smelting and refining have effected enormous economy in time, space, and labour, and have consequently increased the world’s output. With pyritic smelting a sulphuretted copper ore, fed into a cupola in the morning, can be passed directly to the converter, blown up to metal, and shipped as 99% bars by evening—an operation which formerly, with heap roasting of the ore and repeated roasting of the mattes in stalls, would have occupied not less than four months. A large furnace and a Bessemer converter, the pair capable of making a million pounds of copper a month from a low-grade sulphuretted ore, will not occupy a space of more than 25ft. by 100ft.; and whereas, in making metallic copper out of a low-grade sulphuretted ore, one day’s labour used to be expended on every ton of ore treated, to-day one day’s labour will carry at least four tons of ore through the different mechanical and metallurgical processes necessary to reduce them to metal. About 70% of the world’s annual copper output is refined electrolytically, and from the 461,583 tons refined in the United States in 1907, there were recovered 13,995,436 oz. of silver and 272,150 oz. of gold. The recovery of these valuable metals has contributed in no small degree to the expansion of electrolytic refining.

Production.—The sources of copper, its applications and its metallurgy, have undergone great changes. Chile was the largest producer in 1869 with 54,867 tons; but in 1899 her production had fallen off to 25,000 tons. Great Britain, though she had made half the world’s copper in 1830, held second place in 1860, making from native ores 15,968 tons; in 1900 her production was 777 tons, and in 1907, 711 tons. The United States made only 572 tons in 1850, and 12,600 tons in 1870; but she today makes more than 60% of the world’s total. In 1879, Spain was the largest producer, but now ranks third.

The estimated total production for each decade of the 19th century in metric tons is here shown:—

1801–1810	. . . . .	91,000
1811–1820	. . . . .	96,000
1821–1830	. . . . .	135,000
1831–1840	. . . . .	218,400
1841–1850	. . . . .	291,000
1851–1860	. . . . .	506,999
1861–1870	. . . . .	900,000
1871–1880	. . . . .	1,189,000
1881–1890	. . . . .	2,373,398
1891–1900	. . . . .	3,708,901

The following table gives the output of various countries and the world’s production for the years 1895, 1900, 1905, 1907:—

Country.	1895.	1900.	1905.	1907.
United States	175,294	274,933	397,003	398,736
Spain and Portugal	55,755	53,718	45,527	50,470
Japan	18,725	28,285	36,485	49,718
Chile	22,428	26,016	29,632	27,112
Germany	16,799	20,635	22,492	20,818
Australasia	10,160	23,368	34,483	41,910
Mexico	12,806	22,473	70,010	61,127
Russia	5,364	8,128	8,839	15,240
World’s production	339,994	496,819	699,514	723,807

As the stock on hand rarely exceeds three months’ demand, and is often little more than a month’s supply, it is evident that consumption has kept close pace with production.

The large demand for copper to be used in sheathing ships ceased on the introduction of iron in shipbuilding because of the difficulty of coating iron with an impervious layer of copper; but the consumption in the manufacture of electric apparatus and for electric conductors has far more than compensated.

Alloys of Copper.—Copper unites with almost all other metals, and a large number of its alloys are of importance in the arts. The principal alloys in which it forms a leading ingredient are brass, bronze, and German or nickel silver; under these several heads their respective applications and qualities will be found.

Compounds of Copper.—Copper probably forms six oxides, viz. Cu₄O, Cu₃O, Cu₂O, CuO, Cu₂O₃ and CuO₂. The most important are cuprous oxide, Cu₂O, and cupric oxide, CuO, both of Oxides and hydroxides. which give rise to well-defined series of salts. The other oxides do not possess this property, as is also the case of the hydrated oxides Cu₃O₂2H₂O and Cu₄O₃5H₂O, described by M. Siewert.

Cuprous oxide, Cu₂O, occurs in nature as the mineral cuprite (q.v.). It may be prepared artificially by heating copper wire to a white heat, and afterwards at a red heat, by the atmospheric oxidation of copper reduced in hydrogen, or by the slow oxidation of the metal under water. It is obtained as a fine red crystalline precipitate by reducing an alkaline copper solution with sugar. When finely divided it is of a fine red colour. It fuses at red heat, and colours glass a ruby-red. The property was known to the ancients and during the middle ages; it was then lost for several centuries, to be rediscovered in about 1827. Cuprous oxide is reduced by hydrogen, carbon monoxide, charcoal, or iron, to the metal; it dissolves in hydrochloric acid forming cuprous chloride, and in other mineral acids to form cupric salts, with the separation of copper. It dissolves in ammonia, forming a colourless solution which rapidly oxidizes and turns blue. A hydrated cuprous oxide, (4Cu₂O, H₂O), is obtained as a bright yellow powder, when cuprous chloride is treated with potash or soda. It rapidly absorbs oxygen, assuming a blue colour. Cuprous oxide corresponds to the series of cuprous salts, which are mostly white in colour, insoluble in water, and readily oxidized to cupric salts.

Cupric oxide, CuO, occurs in nature as the mineral melaconite (q.v.), and can be obtained as a hygroscopic black powder by the gentle ignition of copper nitrate, carbonate or hydroxide; also by heating the hydroxide. It oxidizes carbon compounds to carbon dioxide and water, and therefore finds extensive application in analytical organic chemistry. It is also employed to colour glass, to which it imparts a light green colour. Cupric hydroxide, Cu(OH)₂, is obtained as a greenish-blue flocculent precipitate by mixing cold solutions of potash and a cupric salt. This precipitate always contains more or less potash, which cannot be entirely removed by washing. A purer product is obtained by adding ammonium chloride, filtering, and washing with hot water. Several hydrated oxides, e.g. Cu(OH)₂·3CuO, Cu(OH)₂·6H₂O, 6CuO·H₂O, have been described. Both the oxide and hydroxide dissolve in ammonia to form a beautiful azure-blue solution (Schweizer’s reagent), which dissolves cellulose, or perhaps, holds it in suspension as water does starch; accordingly, the solution rapidly perforates paper or calico. The salts derived from cupric oxide are generally white when anhydrous, but blue or green when hydrated.

Copper quadrantoxide, Cu₄O, is an olive-green powder formed by mixing well-cooled solutions of copper sulphate and alkaline stannous chloride. The trientoxide, Cu₃O, is obtained when cupric oxide is heated to 1500°–2000° C. It forms yellowish-red crystals, which scratch glass, and are unaffected by all acids except hydrofluoric; it also dissolves in molten potash. Copper dioxide, CuO₂H₂O, is obtained as a yellowish-brown powder, by treating cupric hydrate with hydrogen peroxide. When moist, it decomposes at about 6° C., but the dry substance must be heated to about 180°, before decomposition sets in (see L. Moser, Abst. J.C.S., 1907, ii. p. 549).

Cuprous hydride, (CuH)_n, was first obtained by Wurtz in 1844, who treated a solution of copper sulphate with hypophosphorous acid, at a temperature not exceeding 70° C. According to E. J. Bartlett and W. H. Merrill, it decomposes when heated, and gives cupric hydride, CuH₂, as a reddish-brown spongy mass, which turns to a chocolate colour on exposure. It is a strong reducing agent.

Cuprous fluoride, CuF, is a ruby-red crystalline mass, formed by heating cuprous chloride in an atmosphere of hydrofluoric acid at 1100°–1200° C. It is soluble in boiling hydrochloric acid, but it is not reprecipitated by water, as is the case with cuprous chloride. Cupric fluoride, CuF₂, is obtained by dissolving cupric oxide in hydrofluoric acid. The hydrated form, (CuF₂, 2H₂O, 5HF), is obtained as blue crystals, sparingly soluble in cold water; when heated to 100° C. it gives the compound CuF(OH), which, when heated with ammonium fluoride in a current of carbon dioxide, gives anhydrous copper fluoride as a white powder.

Cuprous chloride, CuCl or Cu₂Cl₂, was obtained by Robert Boyle by heating copper with mercuric chloride. It is also obtained by burning the metal in chlorine, by heating copper and cupric oxide with hydrochloric acid, or copper and cupric chloride with hydrochloric acid. It dissolves in the excess of acid, and is precipitated as a white crystalline powder on the addition of water. It melts at below red heat to a brown mass, and its vapour density at both red and white heat corresponds to the formula Cu₂Cl₂. It turns dirty violet on exposure to air and light; in moist air it absorbs oxygen and forms an oxychloride. Its solution in hydrochloric acid readily absorbs carbon monoxide and acetylene; hence it finds application in gas analysis. Its solution in ammonia is at first colourless, but rapidly turns blue, owing to oxidation. This solution absorbs acetylene with the precipitation of red cuprous acetylide, Cu₂C₂, a very explosive compound. Cupric chloride, CuCl₂, is obtained by burning copper in an excess of chlorine, or by heating the hydrated chloride, obtained by dissolving the metal or cupric oxide in an excess of hydrochloric acid. It is a brown deliquescent powder, which rapidly forms the green hydrated salt CuCl₂, 2H₂O on exposure. The oxychloride Cu₃O₂Cl₂·4H₂O is obtained as a pale blue precipitate when potash is added to an excess of cupric chloride. The oxychloride Cu₄O₃Cl₂, 4H₂O occurs in nature as the mineral atacamite. It may be artificially prepared by heating salt with ammonium copper sulphate to 100°. Other naturally occurring oxychlorides are botallackite and tallingite. “Brunswick green,” a light green pigment, is obtained from copper sulphate and bleaching powder.

The bromides closely resemble the chlorides and fluorides.

Cuprous iodide, Cu₂I₂, is obtained as a white powder, which suffers little alteration on exposure, by the direct union of its components or by mixing solutions of cuprous chloride in hydrochloric acid and potassium iodide; or, with liberation of iodine, by adding potassium iodide to a cupric salt. It absorbs ammonia, forming the compound Cu₂I₂, 4NH₃. Cupric iodide is only known in combination, as in CuI₂, 4NH₃, H₂O, which is obtained by exposing Cu₂I₂, 4NH₃ to moist air.

Cuprous sulphide, Cu₂S, occurs in nature as the mineral chalcocite or copper-glance (q.v.), and may be obtained as a black brittle mass by the direct combination of its constituents. (See above, Metallurgy.) Cupric sulphide, CuS, occurs in nature as the mineral covellite. It may be prepared by heating cuprous sulphide with sulphur, or triturating cuprous sulphide with cold strong nitric acid, or as a dark brown precipitate by treating a copper solution with sulphuretted hydrogen. Several polysulphides, e.g. Cu₂S₅, Cu₂S₆, Cu₄S₆, Cu₂S₃, have been described; they are all unstable, decomposing into cupric sulphide and sulphur. Cuprous sulphite, CuSO₃·H₂O, is obtained as a brownish-red crystalline powder by treating cuprous hydrate with sulphurous acid. A cuproso-cupric sulphite, Cu₂SO₃, CuSO₃,2H₂O, is obtained by mixing solutions of cupric sulphate and acid sodium sulphite.

Cupric sulphate or “Blue Vitriol,” CuSO₄, is one of the most important salts of copper. It occurs in cupriferous mine waters and as the minerals chalcanthite or cyanosite, CuSO₄·5H₂O, and boothite, CuSO₄·7H₂O. Cupric sulphate is obtained commercially by the oxidation of sulphuretted copper ores (see above, Metallurgy; wet methods), or by dissolving cupric oxide in sulphuric acid. It was obtained in 1644 by Van Helmont, who heated copper with sulphur and moistened the residue, and in 1648 by Glauber, who dissolved copper in strong sulphuric acid. (For the mechanism of this reaction see C. H. Sluiter, Chem. Weekblad, 1906, 3, p. 63, and C. M. van Deventer, ibid., 1906, 3, p. 515.) It crystallizes with five molecules of water as large blue triclinic prisms. When heated to 100°, it loses four molecules of water and forms the bluish-white monohydrate, which, on further heating to 250°–260°, is converted into the white CuSO₄. The anhydrous salt is very hygroscopic, and hence finds application as a desiccating agent. It also absorbs gaseous hydrochloric acid. Copper sulphate is readily soluble in water, but insoluble in alcohol; it dissolves in hydrochloric acid with a considerable fall in temperature, cupric chloride being formed. The copper is readily replaced by iron, a knife-blade placed in an aqueous solution being covered immediately with a bright red deposit of copper. At one time this was regarded as a transmutation of iron into copper. Several basic salts are known, some of which occur as minerals; of these, we may mention brochantite (q.v.), CuSO₄, 3Cu(OH₂), langite, CuSO₄, 3Cu(OH)₂, H₂O, lyellite (or devilline), warringtonite; woodwardite and enysite are hydrated copper-aluminium sulphates, connellite is a basic copper chlorosulphate, and spangolite is a basic copper aluminium chlorosulphate. Copper sulphate finds application in calico printing and in the preparation of the pigment Scheele’s green.

A copper nitride, Cu₃N, is obtained by heating precipitated cuprous oxide in ammonia gas (A. Guntz and H. Bassett, Bull. Soc. Chim., 1906, 35, p. 201). A maroon-coloured powder, of composition CuNO₂, is formed when pure dry nitrogen dioxide is passed over finely-divided copper at 25°–30°. It decomposes when heated to 90°; with water it gives nitric oxide and cupric nitrate and nitrite. Cupric nitrate, Cu(NO₃)₂, is obtained by dissolving the metal or oxide in nitric acid. It forms dark blue prismatic crystals containing 3, 4, or 6 molecules of water according to the temperature of crystallization. The trihydrate melts at 114.5°, and boils at 170°, giving off nitric acid, and leaving the basic salt Cu(NO₃)₂·3Cu(OH)₂. The mineral gerhardtite is the basic nitrate Cu₂(OH)₃NO₃.

Copper combines directly with phosphorus to form several compounds. The phosphide obtained by heating cupric phosphate, Cu₂H₂P₂O₈, in hydrogen, when mixed with potassium and cuprous sulphides or levigated coke, constitutes “Abel’s fuse,” which is used as a primer. A phosphide, Cu₃P₂, is formed by passing phosphoretted hydrogen over heated cuprous chloride. (For other phosphides see E. Heyn and O. Bauer, Rep. Chem. Soc., 1906, 3, p. 39.) Cupric phosphate, Cu₃(PO₄)₂, may be obtained by precipitating a copper solution with sodium phosphate. Basic copper phosphates are of frequent occurrence in the mineral kingdom. Of these we may notice libethenite, Cu₂(OH)PO₄; chalcosiderite, a basic copper iron phosphate; torbernite, a copper uranyl phosphate; andrewsite, a hydrated copper iron phosphate; and henwoodite, a hydrated copper aluminium phosphate.

Copper combines directly with arsenic to form several arsenides, some of which occur in the mineral kingdom. Of these we may mention whitneyite, Cu₉As, algodonite, Cu₆As, and domeykite, Cu₃As. Copper arsenate is similar to cupric phosphate, and the resemblance is to be observed in the naturally occurring copper arsenates, which are generally isomorphous with the corresponding phosphates. Olivenite corresponds to libethenite; clinoclase, euchroite, cornwallite and tyrolite are basic arsenates; zeunerite corresponds to torbernite; chalcophyllite (tamarite or “copper-mica”) is a basic copper aluminium sulphato-arsenate, and bayldonite is a similar compound containing lead instead of aluminium. Copper arsenite forms the basis of a number of once valuable, but very poisonous, pigments. Scheele’s green is a basic copper arsenite; Schweinfurt green, an aceto-arsenite; and Casselmann’s green a compound of cupric sulphate with potassium or sodium acetate.

Normal cupric carbonate, CuCO₃, has not been definitely obtained, basic hydrated forms being formed when an alkaline carbonate is added to a cupric salt. Copper carbonates are of wide occurrence in the mineral kingdom, and constitute the valuable ores malachite and azurite. Copper rust has the same composition as malachite; it results from the action of carbon dioxide and water on the metal. Copper carbonate is also the basis of the valuable blue to green pigments verditer, Bremen blue and Bremen green. Mountain or mineral green is a naturally occurring carbonate.

By the direct union of copper and silicon, cuprosilicon, consisting mainly of Cu₄Si, is obtained (Lebeau, C.R., 1906; Vigouroux, ibid.).

Copper silicates occur in the mineral kingdom, many minerals owing their colour to the presence of a cupriferous element. Dioptase (q.v.) and chrysocolla (q.v.) are the most important forms.

Detection.—Compounds of copper impart a bright green coloration to the flame of a Bunsen burner. Ammonia gives a characteristic blue coloration when added to a solution of a copper salt; potassium ferrocyanide gives a brown precipitate, and, if the solution be very dilute, a brown colour is produced. This latter reaction will detect one part of copper in 500,000 of water. For the borax beads and the qualitative separation of copper from other metals, see Chemistry: Analytical. For the quantitative estimation, see Assaying: Copper.

Medicine.—In medicine copper sulphate was employed as an emetic, but its employment for this purpose is now very rare, as it is exceedingly depressant, and if it fails to act, may seriously damage the gastric mucous membrane. It is, however, a useful superficial caustic and antiseptic. All copper compounds are poisonous, but not so harmful as the copper arsenical pigments.

References.—See generally H. J. Steven’s Copper Handbook (annual), W. H. Weld, The Copper Mines of the World (1907), The Mineral Industry (annual), and Mineral Resources of the United States (annual). For the dry metallurgy, see E. D. Peters, Principles of Copper Smelting (New York, 1907); for pyritic smelting, see T. A. Rickard, Pyrite Smelting (1905); for wet methods, see Eissler, Hydrometallurgy of Copper (London, 1902); and for electrolytic methods, see T. Ulke, Die electrolytische Raffination des Kupfers (Halle, 1904). Reference should also be made to the articles Metallurgy and Electro-Metallurgy. For the chemistry of copper and its compounds see the references in the article Chemistry: Inorganic. Toxicologic and hygienic aspects are treated in Tschirsch’s Das Kupfer vom Standpunkt der gerichtlichen Chemie, Toxikologie und Hygiene (Stuttgart, 1893).