 Excerpt from chromosome 21 by the Human Genome Project. This is a LibriVox recording. All LibriVox recordings are in the public domain. For more information or to volunteer, please visit LibriVox.org. Recording by Brian Dirks. Excerpt from chromosome 21 by the Human Genome Project. A-A-G-C-T-T-G-T-C-G-T-G C-C-C-T-T-G-T-G-C-T-C-G-T-G-A-A-G G-G-A-G-T-G-T-G-T-G-T-G-A-C T-T-G-C-C-T-C-T-T-C-G-T-G-T-G-T-G-T-G-T-G-C-G C-C-T-G-T-C-G-C-G-T-G-T-G-T-C-T-T-C-C-G-C-G C-G-C-G-G-G-T-A-A-T-C-G-C-T-T-G-T-G-T-G-T-G-T-G-T-G-T-G G-G-G-G-G-G-G-T-G-G-T-G-C-G-G-G-C-G-G-T-C-G-T-G-G-C T, T, G, T, G, G, C, T, T, G, G, C, G, A, G, C, C, T, G, G, G, T, G, C, G, C, C, T, G, C, A, T, C, T, G, T, G, T, T, C, C, G, G, C, T, G, T, G, G, T, G, C, C, C, T, G, C, T, T, T, T, C, T, T, G, T, T, G, G, C, T, G, G, C, T, T, G, G, T, G, C, G, T, G, G, T, C, C, C, A, G, G, G, G, G, C, T, A, G, T, T, C, G, G, T, C, G, C, T, G, C, C, T, G, G, G, T, C, T, T, T, C, C, G, C, G, T, C, T, C, C, C, G, G, C, C, T, G, G, G, T, C, T, C, C, G, T, G, C, T, G, T, T, T, T, G, T, T, T, T, G, T, C, C, T, T, T, T, T, T, G, G, T, T, G, G, C, G, T, G, G, G, C, T, C, A, C, T, G, G, C, C, G, G, T, A, C, C, C, G, T, T, G, C, T, G, G, A, G, C, C, C, T, C, C, T, G, G, C, C, T, C, G, C, C, C, C, T, T, T, T, T, T, T, G, G, T, C, G, C, C, C, T, G, C, C, C, C, T, C, T, T, G, G, C, C, T, G, C, A, C, T, G, C, T, G, C, T, C, G, C, T, G, G, T, G, C, C, C, G, T, T, C, T, G, G, T, C, G, C, G, G, T, G, C, G, G, T, G, T, C, C, C, G, G, C, C, T, G, G, C, G, G, G, C, G, G, G, G, C, end of excerpt from chromosome 21 by the Human Genome Project. 1st Part of the Entry for Copper, Author Unknown Copper, Symbols C, U, Atomic Wave 63.1, H equals 1, or 63.6, O equals 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 chalcos and ice. As according to Pliny the Roman supply was chiefly drawn from Cyprus it came to be termed Ice Cyprium, which was gradually shortened to Cyprium and corrupted into Cuprum, whence comes the English word copper, the French Quivre 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. JFW Hamp gives 8.945 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 1,000 to 1,200 degrees C. CT Haycock and F. H. Neville give 1,080.5. P. DeJean gives 1,085 as the freezing point. The molten metal is C-green in colour, and at higher temperatures in the electric arc it vaporises and burns with the green flame. G. W. A. Carl Baum succeeded in subliming the metal in a vacuum, and H. Moisson distilled it in the electric furnace. Molten copper absorbs carbon monoxide, hydrogen and sulfur 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 vapor. Its specific heat is 0.899 at 0 degrees C and 0.0942 at 100 degrees. The coefficient of linear expansion per 1 degrees 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. Matheson as 96.4 at 13 degrees 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. Sulfuric 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 sulfate and sulfur 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. 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 kephalopoda and acidia, as hemocyanin, a substance resembling the ferruginous hemoglobin, and of a species of limulus. In straw, hay, eggs, cheese, meat and other foodstuffs, 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 turecin, the colouring matter of the wing feathers of the tiraco. Native copper, sometimes termed by miners malibu or virgin copper, occurs as a mineral having all the properties of the smelted metal. It crystallises in the cubic system, but the crystals are often flattened, elongated, rounded or otherwise distorted. Twins are common. Usually the metal is arboricent, 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 mined 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 in veins and amygdales. The largest known deposits are those in the Lake Superior region, near Kawinor Point, Michigan, where masses upwards of 400 tonnes in weight have been discovered. The metal was formerly worked by the Indians for implements and ornaments. It occurs in a series of amygdaloidal dulmrites, or diabeteses, 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, anonsite, and other zeolitic minerals. Pseudomorphs after calcite are known, and it is notable that native copper occurs Pseudomorphus after aragonite, at Korokoro in Bolivia, where the copper is disseminated through sandstone. Ours. The principal ores of copper are the oxides, cuprite, and malachonite. The carbonates, malachite, and chesulite. The basic chloride, atachonite, the silicate, chrysocolor, the sulfides, calcocite, calcopyrite, erubicite, and tetrahedrite. Cuprite occurs in most cupiferous mines, but never by itself in large quantities. Malachonite 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, chesulite, or azurite, an ore containing, when pure, 55.16% of the metal. Atachonite occurs chiefly in Chile and Peru. Chrysocolor contains in the pure state 30% of the metal. It is an abundant ore in Chile, Wisconsin, and Missouri. The sulfur compounds of copper are, however, the most valuable from the economic point of view. Calcocite, red-ruthite, copper-glance, or vitreous copper, contains about 80% of copper. Copper pyrites, or calcopyrite, 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 sulfide, 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 loads usually run east and west. They occur both in the chelasse, or clay slate, and in the grawen, or granite. Irrubicite, boronite, or horse-flesh ore is much richer in copper than the ordinary pyrites, and contains 56 or 57% of copper. Tetrahedrite, phalaerites, or grey copper, contains from 30 to 48% of copper, with arsenic, antimony, iron, and sometimes zinc, silver, or mercury. Other copper minerals are persilite, bolite, stromaerite, cubanite, stanite, tenantite, implictite, wolfsburgite, famatonite, and inargyte. Metallurgy, copper is obtained from its ores by three principal methods, which may be denominated, one, the pyrometallurgical or dry method, two, the hydrometallurgical or wet method, and three, the electrometallurgical 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 practiced 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 sulfurated 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 affected by calcination or roasting. The roasted ore is then smelted to a mixture of copper and iron sulfides, 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 sulfide, 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 sulfide, 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 Pyrritic 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. One, calcination. Two, smelting in reverberatory furnaces to form the matte. Three, roasting the matte. And four, subsequent smelting in reverberatory furnaces to fine or white metal. Five, treating the fine metal in reverberatory furnaces to coarse or blister copper, either with or without previous calcination. Six, 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 consisting 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. One, calcination. Two, smelting in blast furnaces to form the matte. Three, roasting the matte. Four, smelting in blast furnaces with coke and fluxes to black or coarse metal. Five, 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. In the process of oxidation a certain amount of cupra's 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 affected by stirring the molten metal with a pole of green wood, polling. 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 color by pouring water upon the surface of the molten metal and removing the crust formed. Japan copper is purple red in color and is formed by casting into ingots weighing from six ounces to a pound and rapidly cooling by immersion in water. The color 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 sulfur as is necessary to combine with all the copper present when the calcined ore is smelted. The process is affected either in heaps, stalls, shaft furnaces, reverberatory furnaces or muffle furnaces. Stool and heap roasting require considerable time and can only be economically employed when the loss of the sulfur 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 sulfur and where it is desirable to convert the waste gases into sulfuric acid. Reverberatory roasting does not admit of the utilization of the waste gases and requires fine ores and much labor 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 labor but permit the utilization of the waste gases. Reverberatory furnaces of three types are employed in calcining copper ores. One fixed furnaces with either hand or mechanical rabbling. Two furnaces with movable beds. Three 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 Ohara modified by Alan Brown, the Hickson, the Keller Gaylord Cole, the Ropp, the Spence, the Wethy, the Parks, Pierces turret and Brown's horseshoe furnaces. Blakes and Brunton's furnaces are reverberatory furnaces with a movable bed. Furnaces with rotating working chambers admit of continuous working. The fuel and labor costs are both low. In the White Howl revolving furnace with lifters, a modification of the oxland, the ore is fed and discharged in a continuous stream. The Brookner cylinder resembles the Elliot and Russell black ash furnace. Its cylinder tapers slightly towards each end and is generally 18 feet long by 8 feet 6 inches in its greatest diameter. Its charge of from 8 to 12 tons of ore or concentrates is slowly agitated at a rate of 3 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 planes or prisms which regulate the fall of the ore. In the Gerstenhofer and Hassenclaver-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. Olivier and Perret furnace ore is passed from plate to plate by hand, Meletra or by mechanical means Spence and McDougall. The McDougall 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 Herrschoff who uses a large hollow revolving central shaft called 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 large numbers of inventions. In some the rakes are attached to rigid frames with a reciprocating motion, in others the crossbars 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 greater favour. The McDougall airshoff, 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 the passage of the rakes. The consumption of fuel however does not exceed one of coal to ten of ore. The quantity of ore which these large furnaces with a hearth area as great as 2,000 feet and over will roast, varies from 40 to 60 tons a day. Shaft calcining furnaces like the Gerstenhofer, Hassenklaver and others, designed for burning pyrites, fines, have not found favour in modern copperworks. 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 mat. 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. Today 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 economically in both fuel and labour than the reverberatory. The Welsh method finds adherence only in Wales and Chile. In America the usual method is to roast ores or concentrates, so that the mat yielded by either the reverberatory or cupola furnace will run from forty-five to fifty percent in copper, and then to transfer to the Bessemer converter which blows it up to ninety-nine percent. 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 forty-three feet six inches from bridge to bridge and of a width of fifteen feet nine inches, 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 two hundred 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 mat is made, the growth of the furnace in size has been steady. Richard Pierce's reverberatories in 1878 had an area of half of fifteen feet by nine feet eight inches, and smelted twelve tons of cold charge daily, with a consumption of one ton of cold to two point four tons of ore. In nineteen hundred the furnaces were thirty-five feet by sixteen feet, and smelted fifty tons daily of hot ore, with the consumption of one ton of cold to three point seven 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 arrival in size, equipment and performance. They are round stacks, designed on the model of iron blast furnaces, twenty-nine feet 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 mat of forty to forty-five percent, 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 forty-six percent of silica, it is only through the utmost skill that it can be made to run as low on an average as naught point three percent in copper oxide. As the mat contains an average naught point two percent of silver, it is still treated by the zea vergal 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 richer 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 ten million tonnes 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 thirty-six inches diameter at the two air 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 one hundred and eighty inches by fifty-six inches at the two air 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 and Montana company at Great Falls Montana which have put through five hundred tonnes of charge daily pouring their melted slag and mat into large wells of ten feet in diameter. A combined brick and water-cooled furnace has been adopted by the Iron Mountain Company at Keswick, California for matte concentration. In it the cooling is affected by water pipes interposed horizontally between the layers of bricks. The Mount Lyle smelting works in Tasmania which are of special interest will be referred to later. 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 Holway 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 Borden at Rio Tinto and Claude Votain 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 affected without any embarrassment till metallic copper commenced to separate and chill in the bottom tuyere. To meet this obstacle P. Manhei proposed elevated side tuyere which could be kept clear by punching through gates in a wind box. His invention was adopted by the Vivians at the Eggy Works near Saag, Vaucuse, France and at Leghorn in Italy. But the greatest expansion of this method has been in the United States where more than 400 million pounds of copper are annually made in Bessemer converters. Vessels of several designs are used, some modeled exactly after steel converters, others barrel shaped, but all with side tuyere elevated about 10 inches above the level of the bottom lining. Practice however in a treating copper matte differs essentially from the treatment of pig iron in as much 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 three to six 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 three to fifteen pounds 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 cubic feet 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 is prohibitive. Under favorable 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 tenths to 6 tenths of a cent per pound of refined copper. Pyrritic smelting. The heat generated by the oxidation of iron and sulphur has always been used to maintain combustion in the kilns or stools for roasting pyrites. Pyrritic smelting is a development of the Russian engineer Semenikov's treatment proposed in 1866 of copper matte in a Bessemer converter. Since John Holway'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 sulphurated 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 three into one when however a hot blast is used on highly sulphurated copper ores. A concentration of eight of ore into one 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 pyrritic 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 lower fuel consumption as three of coke to one hundred of charge. Two types of pyrritic 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 work of the Mount Lyle company Tasmania. There the blast is raised from six hundred to seven hundred degrees Fahrenheit in stoves heated by extraneous fuel and the raw ore smelted with only three percent of coke. The ore is a compact iron pyrritis containing copper 2.5%, silver 3.83 ounces, gold 0.139 ounces. It is smelted raw with hot blast in cupola furnaces, the largest being two hundred and ten inches by forty inches. The resulting matte runs twenty-five percent. This is re-concentrated raw in hot blast cupolas to fifty-five percent 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 ninety-seven point six percent of the copper, ninety-three point two percent of the silver, and ninety-three point six percent of the gold. Pyrritic 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 Nichols and James process. Nichols 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 one-quarter to one percent 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 sorts 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. Sulfuric 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 chilli salt, pita 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 lexivated. 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 Schaftner in 1862 and the subsequent recognition of the solubility of the oxide in the same solvent by James Douglas and Steri 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 degrees centigrade 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. Sulfur 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 mats. The precipitation of the copper from the solution in which it is present as sulphate or as cuprous and cupric chlorides is generally affected 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 sulphurated 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. Ours in which the copper is present as sulphate are generally lixivated 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. One, conversion of the copper into a soluble form. Two, dissolving out the soluble copper salt. Three, 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 affected 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 Mijden Peak in Serbia. Joseph Perino's process which consists in heating the ore with iron nitrate to 50 to 150 degrees centigrade 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, stool 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 oxidising roasting of the ores and a subsequent chloridising roasting with either common salt or Abrams out in reverberatory or muffle furnaces. The bulk of the copper is thus transformed into cupric chloride, little cuprous chloride being obtained. This method had long been proposed by William Longmade, Max Schaftner, Becky and Hoped, 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 Dutch 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 affected by allowing the ferrous chloride liqueurs 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 sulphurated ores, are lixivated 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 parietes. 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, a 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 practiced on a commercial scale. Today, 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 sulfate, 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 sidepipes with leaden connections 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 wild dynamo was passed, apparently with a current density of five or six amperes per square foot, 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 sulfuric 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 economise 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, who found that using the low current density of 1.8 ampere per square foot of cathode, and an electrolyte containing one and a half pounds of copper sulfate and half a pound of sulfuric acid per gallon, all the gold, platinum and silver present in the crude copper anode, remain as metals undissolved in the anode's slime or mud, and all the lead remains there as sulfate, formed by the action of the sulfuric acid, or SO4 ions. He found also that arsenic forms arsenic 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 sulfate, 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. Cooprous oxide, sulfide 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 sulfate and stannic oxide. Zinc, iron, nickel and cobalt pass into solution more readily indeed than does the copper. Of the metals present 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 electronegative 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 much 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 neutralised, partly by chemical action on certain constituents of the slime, partly by local action between different metals of the anode, both of which affect solution independently of the current, and partly by the peroxidation or aeration of ferrous sulfate 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 electropositive 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. Eubel has found that with neutral solutions a 5% solution of copper sulfate gave no good result, while with a 20% solution the best deposit was obtained with a current density of 28 amperes per square foot. With solutions containing 2% of sulfuric 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 square foot. The maximum current densities for a pure acid solution at rest were for 15% pure copper sulfate solutions, 14 to 21 amperes, and for 20% solutions 18.5 to 28 amperes per square foot. But when the solutions were kept in gentle motion these maxima could be increased to 21 to 28 and 28 to 37 amperes per square foot, 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 affected in the neighbourhood of the electrodes and partly to the neutralisation of the tendency of liquids undergoing electrolysis to separate into layers due to the different specific gravities of the solution flowing from the opposing electrodes. Such an irregular distribution of the bath, with strong copper sulfate 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 recognised 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 patterns of Fe and A.S. Elmore who sought to improve the character of the deposit by burnishing during electrolysis of E. Dumoulin and Sherrod Cooper Coles who prefers to rotate the cathode at a speed that maintains a peripheral velocity of at least 1000 feet per minute. Certain other inventors have applied the same principle in a different way. H. Thoferne 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 square foot 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 higher current density as 280 amperes per square foot 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 reduction 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 JW Swan has succeeded in depositing excellent copper at current densities exceeding 1000 amperes per square foot. 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 particles 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 square foot. 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 electronegative 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 1.5 to 2 pounds of copper sulfate and from 5 to 18 ounces of sulfuric acid per gallon of water and the space between the electrodes is from 1.5 to 2 inches whilst the total area of cathode surface in each tank may be 200 square feet more or less. The anodes are usually cast copper plates about say 3 feet by 2 feet by 3 quarters or 1 inch. The cathodes are frequently electrodeposited copper deposited to a thickness of about 1.32 of an inch 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 a joining rows of tanks and an overhead traveling 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 Thofferen multiple system, not the jet system alluded to above. There are 600 tanks about 8 and a quarter feet by 4 and a half feet by 3 and a quarter feet deep arranged in three groups of 200 tanks in series. The connections 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 bipolar 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 has calculated that the cost of refining is eight shillings 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 pure than many cast anodes having flat surfaces and being held in place by guides dissolve with great regularity and by a space of only 5 eighths of an inch between the electrodes so that the potential difference between each pair of plates may be reduced to 0.15 to 0.2 volt. J. A. W. Borscher in Germany and A. E. Schneider and O. Zonner 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 the tank with the result that the bubbling of the air upwards aspirates solutions through the vertical pipe from below at the same time aerating 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 electrolyzed with insoluble anodes to recover the copper. The yield of copper per ampere in round numbers one ounce of copper per ampere per DM 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 one pound of copper should require about 0.15 4 electrical horsepower 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 horsepower 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 horsepower hours per pound of copper refined. Many attempts have been made to use crude sulfide of copper or matte as an anode and recover the copper at the cathode the sulfur and other insoluble constituents being left at the anode. The best known of these is the Marchesi process which was tested on a working scale at Genoa and Stolberg in Renish Prussia. As the operation proceeded it was found that the voltage had to be raised until it became prohibitive while the anodes rapidly became combed 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 Nizhny Novgorod in Russia. Siemens and Halsk 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 electrolyzed. In this process 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 affected enormous economy in time, space and labour and have consequently increased the world's output. With pyritex melting a sulfuretted 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 mats 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 sulfuretted ore will not occupy a space of more than 25 feet by 100 feet and whereas in making metallic copper out of a low grade sulfuretted ore, one day's labour used to be expended on every ton of ore treated. Today 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 ounces of silver and 272,150 ounces of gold. The recovery of these valuable metals has contributed in no small degree to the expansion of electrolytic refining. End of copper from Encyclopedia Britannica 11th edition