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METALLURGY A BRIEF OUTLINE OF THE MODERN PROCESSES FOR EXTRACTING THE MORE IMPORTANT METALS BY W. BORCHERS 4 t KGL. TECH. HOCHSCHULE, AACHEN AUTHORIZED TRANSLATION FROM THE GERMAN BY WILLIAM T. HALL AND CARLE R. HAYWARD Instructor in Chemistry Instructor in Metallurgy MASSACHUSETTS INSTITUTE OF TECHNOLOGY FIRST EDITION FIRST THOUSAND NEW YORK JOHN WILEY & SONS LONDON: CHAPMAN & HALL, LIMITED 1911 Copyright, 19x1, BY WILLIAM T. HALL AND CARLE R HAYWARD THE SCIENTIFIC ROBERT DRUMMOND AND COMPANY BROOKLYN, N. Y. PREFACE THE purpose of the book is to present, in concise form, the different processes used for extracting the important metals from their ores, and for refining them. Both the student and the practical engineer can thus get a broad view of modern metallurgical operations, without taking the time required for a study of the more detailed treatises on the subject. Whenever possible, self-explaining illustrations of apparatus have been used in place of descriptions. The micro- photographs are from articles by P. Goerens, Aachen, and W. Campbell, New York, published in Metallurgie. The present translation is commended to the public in the hope that such a summary of metallurgical processes will prove of assistance to English-speaking students in pointing the way to further study. The translators are indebted especi- ally to Professor H. O. Hofman, for his kind advice and assist- ance in reading the proof-sheets. WILLIAM T. HALL CARLE R. HAYWARD MASSACHUSETTS INSTITUTE or TECHNOLOGY, Boston, Mass., February, 1911 ni 365552 TABLE OF CONTENTS 23. Filter Pressing Plant of the London and Hamburg Gold Recovery Co. Ltd., London, Eng. 18 METALLURGY ELECTROLYSIS, OLD PROCESS OF SIEMENS AND HALSKE. Anodes: iron plates. Cathodes: strips of lead, suspended in a long wooden vat divided into cells by hollow partitions. The solution flows up through the cells and down through the partitions. The vats are 23 ft. long, 5 ft. wide and 3 feet deep; they are divided into 6 or 8 cells requiring about 100 amperes total current with a current density of about .05 ampere per square foot and a potential of 2 volts per cell. Capacity 1800 cu!ft. solution (13,000 gallons) per day. When the deposited gold has reached FIG. 35. the proper thickness, the cathodes are removed from the solution, dried, melted, and cupelled for crude gold (850-900 fine). ELECTROLYSIS BY MORE RECENT METHODS. For solutions low in copper, Butters' process may be used. The cathodes are of tin plate, from which the gold continually drops off to the bottom of the precipitation vat, which is built like a spitzlutte, and thence it is withdrawn by means of a small stream of liquid and filtered outside the vat. Since with solutions running high in copper the deposit is more dense, Charles P. Richmond (Metallurgie 4, 502) GOLD 19 has gone back to lead cathodes. These, when covered with a gold or copper deposit of proper thickness, are FIG. 36. Butters' Vacuum Filter. Frames made of iron rods covered with canvas. The spaces between are filled with cocoa matting. The perforated iron pipes are connected to suction tubes. Operation of the filter: The vat is filled with slime and suction applied. The solution is drawn through the hollow frame until the cake on the outside of the filter becomes so thick that filtration is difficult. The slime is then withdrawn from the vat and water admitted for washing the cakes. After washing, the cakes are removed from the filter by applying pressure instead of suction. FIG. 37. Cross-sections of Zinc Boxes. FIG. 38. removed from the solution, enclosed in filter bags, and placed in an acid bath, where they are used as anodes. The gold remains behind as slime, while copper is dissolved 20 METALLURGY and goes to the cathode, but if a high current density is used it falls off and is recovered as a slime similar to cement FIG. 39. Plan. A-B FIG. 40. Sections EF and CD. J]J FIG. 41. Section AB. copper. The gold which remains as slime in the bags is dried and fused as described above. Parting : gold as a residue. Most of the so-called parting processes GOLD 21 consist in dissolving away, either in aqueous solutions or by fusion, the metals alloyed with the gold, leaving the latter FIG. 42. Anode. FIG. 43. Cathode: in the FIG. 44. Wooden Second Part of the Process: Frame. Anode. behind. Of the many methods that have been devised, the following are still used in practice : Nitric acid parting, inquartation j Sulphuric acid parting, \ Described under Silver. Fusion with sulphur. Rossler process J Electrolytic parting and purification of other metals. See Silver, Copper and Nickel. (B) Gold Refining Solution and Precipitation of the Gold : Chlorinating Solution. Chlorine and solutions evolving chlorine, particularly aqua regia, are the principal agents used in the final refining of gold. For dissolving the gold, dilute aqua regia suffices (2 parts concentrated HC1, i part concen- trated HNOs, 3 parts H 2 O). Apparatus: porcelain vats. The gold solution is filtered to remove the insoluble residue (AgCl) and the following operations are then carried out: Precipitation of the gold by ferrous sulphate, filtration, washing, drying, and melting in a graphite crucible under a cover of glass and borax. If platinum is present, it is 22 METALLURGY precipitated by adding metallic iron to the solution from which the gold has been removed. Electrolytic Parting, used for refining gold that contains platinum. ANODES: gold platinum alloy. CATHODES: sheets of refined gold. Distance between elec- trodes 1.2 inches. ELECTROLYTE: gold chloride solution containing 25-30 oz. of gold and 20-50 oz. HC1 (sp.gr. 1.19) per cubic foot (7.5 gallons). If the anodes contain lead some H 2 SO4 is added. The large amount of free HC1 in the electro- lyte is required because the anode gold goes into solution only when the conditions are favorable for the formation of HAuCU, the reaction being: Au + 3Cl + HCl = H(AuCl 4 ). Furthermore, the greater the current density, the larger the quantity of free acid that must be kept in the bath; a high current density is required for the rapid de- position of the gold. Electrolytic solution of the gold is increased by rise of temperature. The conditions for electrolysis are, therefore: CURRENT DENSITY: 100 amperes per square foot, at times rising as high as 300 amperes per square foot. POTENTIAL: i volt. TEMPERATURE: 60 to yoC. (140 to 158 F.). ELECTRO LYZING VESSELS: stone jars or porcelain beakers immersed in a water bath. An automatic method is used for replacing the water lost by evaporation from the water bath; the wash- water from the cathodes and anode mud is used chiefly for this purpose. PLATINUM, although insoluble in HC1 when used alone as an anode, goes into solution somewhat in the presence of gold, but it is not deposited on the cathode as long as the amount of platinum in the electrolyte does not become more than double the amount of gold. The platinum may be precipi- tated from a platinum-rich electrolyte by adding NH 4 C1, and the same is true of palladium. Silver is converted into insoluble AgCl. GOLD 23 THE PRODUCTS of electrolytic refining are gold, 999.8 to 1000 fine, platinum, and silver chloride. Dissolving or Slagging the Metals Alloyed with the Gold: SULPHURIC ACID REFINING consists in repeatedly boiling the gold with concentrated H 2 SO4. The further treat- ment of the gold residue is the same as described under Chlorinating Solution (p. 21). NITRIC ACID REFINING consists in boiling repeatedly with concentrated nitric acid (see Silver, Ag-Cu parting). FIG. 45. Crystals of Precipitated Gold (X35). REFINING FUSION : gold containing small amounts of impur- ities, and also the precipitated gold as described on p. 21, which is to be melted and cast into bars, is fluxed with borax and glass, and if necessary with an oxidizing agent such as niter, and melted in graphite crucibles. Properties of Refined Gold : SPECIFIC GRAVITY: 19.3. COLOR: yellow with brilliant lustre. TENACITY: very tough, the most tenacious of metals. FRACTURE: hackly. 24 METALLURGY STRUCTURE: (Fig. 45). MELTING POINT: 1064 C. (1947 F.). VAPORIZATION at 2000 C. (3632 F.). ELECTRICAL CONDUCTIVITY: 0.6 to 0.7 referred to Ag=i. ALLOYS with most metals. The alloys of Au with Pt, Ag, Hg, Cu, Pb, and Zn are important metallurgically. CHEMICAL BEHAVIOR: not very active, has a low solution tension, and its compounds are readily broken down. In extracting gold, the most important solvents are Cl, Br, and KCN. PLATINUM Sources Natural Sources : FREE, alloyed with the other platinum metals as well as with Fe, Au, and Cu. MINERALIZED, occasionally as a compound with arsenic (Sperrylite, PtAs2) but more frequently in Ni, Cu, and Au ores, although in small quantities. Other Sources : Smelter products: In matte and bullion from smelting platinum-bearing Ni, Cu, and Au ores. Extraction Concentration. Wet, mechanical methods are nearly always util- ized in the preliminary concentration of platinum-bearing gravel. Solution in Metals : In Lead. This is carried out occasionally by smelting platinum ores with PbO and PbS, whereby the Fe and Cu in the ores are converted into matte while the platinum metals enter the lead bullion, from which they are subsequently recovered by cupellation (see Silver Cupellation Methods, pp. 33-35). In Nickel ] These processes take place along with the treat- In Copper V ment of Ni, Cu, and Au ores. See these metals In Gold J and their electrolytic separation. Chemical Solution and Precipitation of the Platinum : Chlorination in aqueous solution by means of mixtures evolving chlorine, especially aqua regia, is carried out in the refining of alloys. From the resulting solution the platinum is sub- sequently precipitated either by iron (see Pt-Au parting), or by ammonium chloride. 25 26 METALLURGY Refining Chemical Solution and Precipitation is utilized both in the direct working of the ores and in the parting of alloys or of crude platinum. The following treatments are given: Purification with HC1, containing but a small amount of HNO 3 , so that only the base metals are dissolved. This treatment is FIG. 46. Surface of Fused Platinum (X33). not given unless imperative. As a rule it is only necessary to carry out Chlorination in aqueous solution as described above. If the solution carries gold, this can be precipitated by ferrous sulphate or by electrolysis. The separation of the accom- panying platinum metals takes place usually by Partial Precipitation with NH 4 C1, if necessary after the pre- vious reduction of Pd and Ir chlorides from the " ic " to the " ous " state. Ammonium chloride precipitates the platinum as ammonium chloroplatinate, (NH 4 ) 2 PtCl 6 . The latter after PLATINUM 27 Separation from the Solution by decantation, filtering, wash- ing, and drying is changed by Ignition into platinum sponge: (NH 4 ) 2 PtCl 6 = Pt + 2NH 4 C1 + 2C1 2 . Melting. In order to convert the precipitated powder or the sponge into a compact mass, it is heated by an oxyhydrogen blowpipe in a furnace made of limestone. This also removes the last traces of impurities. Properties of Platinum : SPECIFIC GRAVITY: 21.5. COLOR: grayish white, brilliant lustre. TENACITY: very ductile, very high tensile strength, capable of being drawn into the finest wires and the thinnest sheets (platinum foil). STRUCTURE : see Fig. 46. MELTING-POINT: 1745 C. (3173 F.). ELECTRICAL CONDUCTIVITY: 0.08 (Ag=i). ALLOYS readily with most metals, especially when easily fusible, also with H. CHEMICAL BEHAVIOR: not very active, low solution tension; compounds easily dissociated. For the extraction of platinum Cl is the most important solvent. SILVER Sources Natural Sources : NATIVE, alloyed with Au, Cu, Hg. MINERALIZED, principally in combination with the halogens and sulphur. Hornsilver, AgCl. Bromargyrite, AgBr. lodargyrite, Agl. Silver glance, Ag 2 S, occurs free, in sulpho salts (ruby silver, tetrahedrite), and as solid solu- tion in sulphides. Other Sources : Burnt pyrite, matte from copper and lead smelting, slags, drosses, and alloys from metallurgical processes (black copper, base bullion, zinc skimmings, etc.). OLD METAL and metallic scrap. (A) Extraction Concentration: Nearly always carried on in connection with amalgamation (p. 35 et seq.). Solution of Silver in Metals : Copper, lead, and mercury are commonly used as solvents. IN COPPER: (see Gold and Copper). IN LEAD: Silver, as it occurs free, or as it results from the decomposition of its compounds, is taken up by molten lead, and the two metals are subsequently separated. Materials poor in silver (under 10%) are smelted with lead ores (see Lead). Materials rich in silver (over 10%) are treated on a bath of molten lead (see Cupellation, pp. 33-35). With lead as the carrier, most sulphide ores are satisfac- torily treated, either directly or after roasting. Lead-free copper ores carrying precious metals are not used for this 28 SILVER 29 purpose, but are worked by treating with them copper acting as the carrier. Ores, either raw or roasted, containing anti- mony and arsenic and various other materials not too rich in copper are used. Operations : i. SMELTING of materials low in silver, with lead ore to FIG. 47. Refining Furnace. Scale, i : 100. obtain a base bullion low in silver (about i% Ag). Smelt- ing directly to rich bullion leads to high slag losses. . REFINING THE BASE BULLION. Melting in a deep-hearth reverberatory furnace at a low temperature. Removal of the mechanically-enclosed impurities and of the difficultly- fusible constituents ; the material withdrawn, usually copper-bearing, is returned to the ore- smelting furnace. Heating in an oxidiz- ing flame at a high tem- perature in order to aoo remove the larger part of the antimony. The oxi- dation product (skim- mings: Pbs(SbO4)2 + #PbO), is refined under a covering of charcoal in order to reduce part of the PbO. The refined skimmings are smelted for hard or antimonial lead, a Pb-Sb alloy with 14-20% Sb. , \ E A B \C \ 303C R ft -4*-^ D 5 10 FIG. 48. 30 METALLURGY 3. CONCENTRATION of the silver in a portion of the lead. This may be accomplished either by crystallization (Pattin- son process) or by crystallization after adding zinc, and then separating the Ag as an Ag-Zn-Pb alloy (Parkes process). The enrichment of the bullion during the crystallization of the lead takes place according to the accompanying freezing-point curve of silver-lead alloys. Since the eutectic contains about 4% silver, on cooling a bullion with about i% silver, from 340 C., lead will begin to separate at E, thus enriching the mother-metal in silver. If we cool, for example, to 306 C., the ratio of crystallized lead to the still liquid silver-lead alloy will be as BC:AB. In prac- tice it is not possible to accomplish such a sharp separa- tion between Pb and Pb-Ag alloy, for, in the skimming off of the crystallized lead or in the withdrawal of the liquid Pb-Ag alloy, some of the alloy adheres mechanically to the crystals, or small crystals remain suspended in the molten alloy. Two methods have been devised for the practical working of this process: THE ORIGINAL PATTINSON PROCESS. The base bullion is melted in iron kettles and the dross skimmed. The cooling of the lead is hastened by spraying with water. After crystallization begins, the crystals are transferred by means of a perforated iron skimmer to a neighboring kettle and the process continued until only about one- third of the original contents remains in the first kettle. To the remaining liquid, either in the same or a neighbor- ing kettle, more lead of the same silver value is added and the process repeated. Similarly the skimmed crystals, which are not yet sufficiently desilverized, are re- melted and subjected to the same treatment. Thus the kettle at one end of a row will finally hold the enriched bullion (1.5 to 2% Ag) while the kettle at the other end will receive the desilverized lead. THE LUCE-ROZAN PROCESS (Figs. 49-50, p. 32). The base bullion is melted in tilting kettles which discharge SILVER 31 into the crystallizing kettle. Steam is blown through the latter until two-thirds of the lead has been crystallized and then the enriched bullion is removed by tapping. The crystallizing kettle is heated to remelt the lead crystals and refilled with bullion of the same silver value. The crys- tallization is repeated and the process continued until the lead remaining in the kettle is sufficiently desilverized. It is remelted and tapped into molds. ZINC DESILVERIZATION is based upon the slight solubility of Zn in Pb, and conversely of Pb in Zn, and upon the high melting-point of the Zn-Ag-Pb alloy compared with that of the lead. At 350 C., Pb dissolves 0.6 Zn and at 650 it dissolves 3% Zn. The different operations of zinc desilverization are as follows : The Cu is removed from the previously-purified, melted base bullion by adding to it a small amount of Zn (2%), allowing it to cool and removing the crystallized alloy. As soon as a sufficient quantity of copper skimmings have accumulated, they are liquated, thus producing (i) liquated bullion, which is returned to the bullion that has already been freed from Cu, and (2), a copper alloy low in silver (copper dust) . The copper dust is melted and oxidized with steam, thus separating it into a rich bullion and oxides of Cu and other metals. These rich oxides are leached with H 2 SO4 so that a solution containing the sulphates of copper and of zinc results and a residue com- posed of particles of bullion which had been entangled in the oxides. The copper is precipitated from the solution by zinc scrap and the resulting solution of zinc sulphate yields zinc vitriol, ZnSC^.yH^O, on evaporation. The bullion, freed from copper, is now heated up and the greater part of the silver removed by means of a larger addition of Zn (in the form of pure zinc, and skim- mings poor in silver). The alloy which has crystallized 32 METALLURGY FIG. 50. Pattinson Plant, using the Luce-Rozan Process. SILVER 33 out (zinc crust), is removed and liquated in another kettle. The liquated lead goes back to the desilverizing kettle. The residue (rich zinc crust) is distilled, giving rich bullion and zinc. The desilverized lead is freed from zinc by blowing steam through it. The resulting mixture of zinc oxide and lead oxide which floats on the surface of the metal is skimmed, washed, and sold as pigment. If the lead still contains antimony, further oxidation is accomplished by means of an air current. The antimony is thus removed in the litharge which is formed at the same time. As a residue, there remains in the kettle pure lead, the so-called "soft lead." As apparatus for desilverizing, hemispherical or flat cast-iron kettles with stirrers are used; for the distillation of the rich zinc crust, graphite or clay retorts; and for cupelling, reverberatories with hearths made of bone-ash or cement and lime. . Lead-lined wooden vats serve for working up the so-called rich oxides, and for the poor oxides resulting from the dezincification of the lead, a jig is used together with a series of wooden slime tanks. The rich bullion is converted, by cupelling, into litharge and crude silver, both the English and German methods being used. In the German process large reverberatory furnaces are used which take the entire charge at the beginning of the operation. In the English process, small reverberatory furnaces are used, the hearths being kept full during the operation by repeated additions of bullion until, finally, they are full of crude silver. If liquated bullion is not used, the first litharge drawn off will be impure. After the removal of this, silver scrap may be added to the bullion. The base metals are oxidized and the precious metals dissolved in the bullion. Now begins the cupellation proper, namely the oxidizing smelting of the lead whereby the resulting litharge is con- stant! v allowed to flow out of the furnace. Part of the 34 METALLURGY litharge is absorbed by the hearth which, before starting, has been lined with bone-ash (now little used) or cement and lime. During the cupellation the metal surface is com- pletely covered with yellow, glowing litharge which at the last recedes from the bluish surface of the metallic silver FIG. 51. Desilverizing Kettle and Auxiliary Apparatus, with stirrer, (the blick). Since, in spite of the fact that the surface of the lead is covered with litharge during almost the whole of the cupelling operation, the oxidation takes place com- FIG. 52 Desilverizing Kettle and Auxiliary Apparatus, with siphon. paratively quickly, it must be ascribed to the formation of a higher oxide of lead, PbC>2. In fact, PbC>2 is formed readily in an oxidizing atmosphere at the temper- ature of the cupelling hearth in the presence of strongly basic oxides. PbO, however, is itself a strongly basic SILVER 35 Desilverizing Kettle and Auxiliary Apparatus. Fig. 51, with stirrer. Fig. 52, with siphon. Figs. 53 and 54, with lead pump and molds. oxide, and can form compounds with PbO2 such as, for example, the lead plumbate found in red lead : 36 METALLURGY This compound carries the oxygen of the air through the litharge layer to the lead. Without it, the oxidation of the lead would take place much more slowly, or some device other than the strong air current playing upon the melt would be necessary to keep fresh surfaces of lead in contact with ah*. The products from the true cupellation are: litharge, PbO; hearth lining saturated with litharge; lead fume, the volatilized lead oxide which has been caught in the dust chamber and contains other volatile substances; and finally the silver itself, which is known as blick silver. This last may contain small amounts of lead, bismuth, copper, etc., also any other precious metals that may have been present in the ores. The litharge, hearth, and flue dust go back to the ore smelter. Amalgamation. This includes the solution in mercury of pre- cious metals, either native or resulting from the decomposition of their compounds, the mechanical purification of the result- ing amalgam and the separation of the precious metals from the mercury by distillation. AMALGAMATION WITHOUT CHEMICALS: applicable to ores containing the precious metals in the free state or as com- pounds decomposable by mercury. It is assumed in the first case that the precious metals are not widely dissem- inated through the gangue. Coatings of other metallic compounds surrounding the particles of precious metals may also prevent the necessary contact of the latter with the mercury. The following methods are used in this type of amalgamation: AMALGAMATION DURING CRUSHING AND SUBSEQUENT MECHANICAL TREATMENT: These methods differ from ordinary ore-dressing processes only by the use of mercury, free or in the form of copper or silver amal- gam, in various parts of the apparatus used for crush- ing and concentrating; in the last case, silver-plated copper plates are used. The apparatus of the ore- SILVER 37 SECTION a-&, c-d "'- Fig, 55 FIG. 56. German Cupelling Fumace. 38 METALLURGY dressing plants requires no important changes, so that the following divisions may be made of these amalga- mation processes: HYDRAULIC METHODS in connection with sluice amal- FIG. 59. English Cupelling Furnace. gamation (Hg is kept on the bottom of the sluice in grooves of the pavement). STAMP-MILL AMALGAMATION 1 Both processes are WET-GRINDING AMALGAMATION J completed on AMALGAM CATCHERS AND AUXILIARY-AMALGAMATORS. Amalgamated copper plates used as the bottom of shallow sluices or frequently, amalgamating pans; as for example in the Laszlo system, the tailings from which it is sometimes necessary to concentrate by passing over shaking-tables of the Frue, Wilfley, or Ferraris type, especially for recovering pyrite, arsen- opyrite, etc. SILVER 39 CRUSHING AND AMALGAMATION AS SEPARATE PROCESSES. In this case the amalgamators serve chiefly for mixing FIG. 60. FIG. 61. Laszlo Amalgamator. the ore with mercury. They may be classified, according to their form into MORTAR AMALGAMATION. 40 METALLURGY WET-MILL AMALGAMATION OR PAN AMALGAMATION: Schemnitz amalagmator, Laszlo amalgamator, Ameri- can amalgamating pans. These contain amalgamated copper plates as bottoms to shallow sluices, and various sluices with devices for feeding the mercury and collecting the amalgam. AMALGAMATION WITH CHEMICALS: the ores may be amal- gamated without previous roasting by the Patio Process or Old-fashioned, Mexican Heap-Amal- gamation. This consists of the following operations: FIG. 62. Arrastra. FIG. 63. 1. PRELIMINARY CRUSHING of the ore in edge mills or stamp mills. 2. FURTHER CRUSHING in wet grinders called arrastras (Figs. 62 and 63). 3. PARTIAL DE- WATERING OF THE SLIMES in low heaps surrounded by sand dams. 4. PREPARATION OF THE AMALGAMATION HEAPS (tor- tas) by spreading out the now-thickened slime upon a paved floor surrounded by stones. Diameter of the floor: 25 to 50 ft., depth 6 to 12 in. 5. INCORPORATION (by treading in with mules) of from 2 to 5% of common salt (based on the weight of ore); 0.25 to 0.5% of CuSO 4 .5H 2 O, in the form of magistral (a product obtained by the roasting SILVER 41 of pyrite which contains CuSO 4 and CuCl 2 besides ferric salts); and 6 to 8 Ibs. of Hg per pound of Ag in the ore. 6. WET CONCENTRATION OF THE SLIMES after the completion of the amalgamation. 7. TREATMENT OF THE AMALGAM (see below). The Patio process, in spite of its great age (already in use 300 years) is a very slightly developed method which gives satisfactory results only with ores that have been prepared by weathering processes, but the mercury loss is always high (200% of the recovered silver). For the chemical reactions see the Krohnke process. KROHNKE PROCESS: suitable for ores which contain the precious metals either free or in sulphides or chlorides from which the precious metals only are to be extracted. Other metals remain in the amalga- mation residue. REAGENTS: Cuprous chloride dissolved in brine; metals which are electropositive to silver (Zn, Pb, Cu) used in the form of amalgam; and finally mercury. CHEMICAL REACTIONS: Cii2Cl 2 (or CuCl) when dissolved in a concentrated NaCl solution will quickly decompose Ag 2 S. Ag 2 S + 2CuCl * 2 AgCl + Cu 2 S. With the degree of fineness practically obtainable in crushing, not over 80% of the silver contained in simple or complex native sulphides is obtained by this method. The stopping of the reaction cannot here be attributed solely to the law of mass action. Opposing the action of the CuCl is the mechanical resistance offered by the Cu 2 S crust adhering to the grains of Ag 2 S. CuCl alone, therefore, is insufficient as a solvent of Ag 2 S in amalgamation. In the presence of metals electropositive to silver, such as Zn, Pb, and Cu, the Ag 2 S is reduced to Ag in 42 METALLURGY spite of the coating of other sulphides (Cu 2 S) or salts. Cu 2 S under these circumstances does not hinder the completion of the reaction, but acts as a metallic conductor for the exchange of energy between the more positive metal and silver: it also forms with these and the solution a galvanic couple in which Ag 2 S is the cathode and Zn is the anode. If Cu or CuCl is lacking at the beginning of the process, Zn acts upon Ag 2 S as follows: In spite of the seemingly complete separation of Ag from Ag 2 S, in this case there is another drawback which causes imperfect amalgamation. ZnS, which is insoluble, envelopes the silver and hinders its amalgamation. By using Zn and Hg in a pure NaCl solution, as much as 50% of the Ag that has been converted to the metallic state by the Zn may remain unamalgamated. Copper works more favorably and, in fact, as Krohnke pointed out, through the formation of CuCl, for CuCl causes the decomposition of Ag 2 S, but ZnCl 2 does not. But, as stated above, CuCl alone is also not sufficient of itself to completely decompose Ag 2 S, and the solution tension of Cu is weaker than that of Zn. If, now, care is taken at the outset to provide for the presence of CuCl and Zn is added, then the following reaction takes place without the formation of ZnS: 2CuCl + Ag 2 S + Zn = Ag 2 + Cu 2 S + ZnCl 2 . Cu 2 S does not hinder the amalgamation, as it is more brittle than ZnS and the mechanical movement of the pulp easily exposes the surface of the Ag to the action of the Hg. SILVER 43 OUTLINE OF THE KROHNKE PROCESS: I. CRUSHING: wet edge-mills called Chilian mills are used. 2. COLLECTING THE SLIME in a system of settling-tanks, FIG. 64. Amalgamation Barrel. Length, 6 Feet. Diameter, 5! Feet. Made of 3-inch Pitch Pine Planks. Wooden (formerly iron) rods pass through the barrel between the heads in order to prevent the charge from balling up. Capacity, 6 tons of ore besides the reagents. each of 4000 cu.ft. (100 metric tons) capacity, in which the water is removed by filtration and evaporation. 3. PREPARATION OF THE AMALGAM: Zn is heated with ten times its weight of Hg under acid water; 44 METALLURGY Pb with twice its weight of Hg without acid; then filtered. Zn amalgam contains 14 to 17% Zn. Pb amalgam contains about 45% Pb. The Zn content may be increased to 58% by heating and slowly cooling a large mass, in which case filtration is unnecessary. APPARATUS : cast-iron kettle or, for large quantities^ a sheet-metal kettle (bottom of one piece). The bottom edges should project beyond the source of heat so that any mercury which trickles out or is spilled may be recovered. 4. CUPROUS CHLORIDE SOLUTION: i part by weight of CuSO 4 .5H 2 O +0.5 to i% H 2 SO 4 + 6 to 20 parts saturated NaCl solutions + scrap copper (CuSO 4 4- 2 NaCl + Cu = Na 2 SO 4 + 2CuCl) are heated by steam for from one half to two hours (kept saturated with NaCl) in double-walled wooden vats. The space between the walls is packed with a mixture of tar and powdered lime. 5. AMALGAMATION is accomplished in barrels (see Fig. 64). The barrels are charged in succession with the following reagents, boiling NaCl solution, Zn amalgam (or Zn and Hg followed by the rotation of the barrel to produce amalgam), then CuCl solution, followed immediately by the roughly-broken lumps of air-dried slime, which disseminates through the mass, forming a thick paste. The barrels are kept turning slowly (four or five revolutions per min.) for four or five hours. Power required, 8-9 H.P. 6. WORKING UP THE AMALGAM: After completing the amalgamation, the barrels are allowed to stand for from one to three hours; then they are rotated for a short time, at first quickly, then slowly. Every- SILVER 45 thing is then allowed to run out, the barrels are washed out, and the addition of water is continued until its volume amounts to eight times that of the slimes. The thinned slimes are worked up in set- tling pans, from which the pulp gradually overflows and the amalgam is finally withdrawn at the bottom. FIG. 65. American- Amalgamating Pan. FIG. 66. Amalgam Filter. The chemical purification of the amalgam, particu- larly from copper, is accomplished Either by treatment with Ag 2 S or AgCl; By treatment with CuCb-NaCl liquor; Or by treatment with NH 3 and air. The amalgam is finally filtered (Fig. 66) , this being ultimately hastened by a centrifugal machine, and then it is distilled (Figs. 67 and 68). 46 METALLURGY CASO OR CALDRON PPOCESS: REAGENTS: Cu, NaCl, Hg. APPARATUS: copper pan. APPLICABLE: only for the purer ores carrying native FIG. 67. The Krohnke Furnace for Distilling Amalgam. FIG. 68. Krupp Furnace for Distilling Amalgam. silver or AgCl. These are seldom available and the process is little used at present. WASHOE PROCESS: REAGENTS: CuSO 4 , NaCl, Fe, and Hg. The princi- pal reagent, as in the Krohnke process, is CuCl, the product obtained by the reaction mentioned above, under 4. The action of the reagents is hastened by grinding between iron plates in iron pans which are heated by steam. SILVER 47 AMALGAMATION APPARATUS: wooden or iron vats with removable bottom plates and mullers (Fig. 65). PROCEDURE: the ore is crushed in stamp mills, partially de-watered in settling tanks and then sub- jected to wet grinding, with the above reagents, in steam-heated pans. The pulp is subsequently concen- trated and the amalgam purified and distilled. After a preliminary chloridizing roast, ores may be treated by Barrel amalgamation. Pan or vat (Tina) amalgamation. The preliminary treatment is essentially the same for both processes: Preliminary crushing (rock break- ers); drying (usually in rotary cylinders); further crushing with salt (NaCl) in stamp mills (dry stamping) or in other crushing apparatus (edge mills or ball mills). Chloridizing roast in rever- beratory furnaces (with stationary hearths and movable stirrers Pearce, Brown, and Wethey; with movable hearths Bruckner and White-Howell) or in shaft furnaces (Stetef eldt) . The amalgamation proper now takes place either in barrels, in which case there is a preliminary treat- ment with iron and water to reduce ferric and cupric compounds to ferrous and cuprous compounds and to reduce AgCl to Ag, after which mercury is added; or by pan amalgamation, as in the Washoe process; or by vat amalgamation in wooden vats with copper bottoms and copper mullers. In the latter case the copper acts as the reducing material for cupric and ferric compounds. The amalgamation procedure is the same as that described above. Solution of the Silver in Chemical Solvents. Ziervogel Process: Applicable to matte rich in silver. The matte is concentrated to about 74-75% Cu and carries at the 48 METALLURGY Mansfeld Works 0.4 to 0.5% Ag. With higher percentages, the Ag segregates in large pellets which is unfavorable for good extraction. The process consists in converting Ag to Ag 2 SO 4 and the Cu to CuO, leaching out and precipitating the Ag, and smelting the CuO for Cu. The operations are: 1. Breaking of the matte by hand. 2. Grinding in ball mills. 3. Preliminary roasting for two days in a roasting furnace with mechanical stirrers; this converts the copper to CuSO 4 . 4. Final roasting in a furnace of the same type, which pro- duces CuO and Ag 2 SO 4 . 5. Leaching out the Ag 2 SO 4 with warm water. 6. Precipitation of the silver with granulated copper in small stoneware vats fitted with filter bottoms. 7. Washing, pressing, drying, and fusing the silver. 8. Re-roasting the leached oxide and further treatment again as under 5, 6, and 7. 9. The CuSO 4 resulting from the silver precipitation is con- centrated and crystallized to form blue vitriol (CuSO 4 .5H 2 O). In the following processes, the Ag is either already present as AgCl in the ore or it is converted into AgCl by a chloridizing roast and leached with chloride or hyposulphite solution from which it is subsequently precipitated. Among these processes, the following have found limited usage : The Old Oker-Longmaid-Henderson Process : leaching of the pyrite, after a chloridizing roast, with water from the con- densation tower of the roasters. Precipitation of the Ag with Nal and the Cu with Fe. Recovery of the I by decom- posing the Agl with Na 2 S. Augustin Process : the ore is leached with a saturated brine. One gallon of saturated NaCl solution dissolves at 2oC. 0.169 oz - Ag ( IO oz - NaCl: 0.4 oz. AgCl). The Ag is precipitated from the brine by means of Cu. Kiss Process: the ore is leached with CaS 2 Os solution and the silver precipitated by Ca(SH) 2 . SILVER 49 Russell Process: leaching with 4Na 2 S 2 O 3 -301128203 solu- tion and precipitation with Na 2 S. Patera Process: this process is used extensively, especially in plants designed and built by Ottaker Hofmann. The Patera-Hofmann process consists of the following opera- tions : 1. A PRELIMINARY LEACHING with water to remove chlo- rides of the base metals. APPARATUS: wooden vats, 15-20 ft. in diameter and 5 ft. deep with removable filter bottoms of lattice work. All wooden surfaces are coated with asphaltum. 2. LEACHING OUT THE SILVER with sodium "hyposulphite" containing 0.25-0.5% Na 2 S 2 O 3 -5H 2 O. 100 g. of Na 2 S 2 O3-5H 2 O will dissolve 40 g. of AgCl, forming Ag 2 S 2 O3-Na 2 S 2 O3-2H 2 O. Same apparatus as in i. 3. PRECIPITATING of the Ag with Na 2 S or calcium poly- sulphide, the latter being prepared by boiling milk of lime with sulphur. The calcium polysulphide solution is preferred because it prevents the concentration of Na2SO4 in the solutions, the SC>4 being precipitated as CaSC>4. The Na 2 S resulting is oxidized to Na 2 S 2 O3, so that no fresh supply of this salt is required. 4. FILTRATION : the precipitate is transferred to a settling tank and from there to a filter press. 5. DRYING AND ROASTING in a reverberatory furnace. 6. CHARGING INTO A HOT LEAD BATH and cupelling the rich argentiferous lead. Nitric Acid Parting: This process, known as inquartation, is only used to-day, except for Au-Ag parting in assaying, according as the demand for the resulting AgNOs permits. Ag is dissolved by HNO 3 , leaving the Au unattacked. If the ratio of Au to Ag in the alloy is between i : 3 (hence the name inquartation) and 1:1.75, the Au remains behind in a coherent state after heating the alloy with HNOs. If the Au content is higher, the resulting Au carries Ag even after repeated treatment with HNC>3. With less Au, 50 METALLURGY T Plant for Nitric Acid Treatment of Anode Mud from Electrolytic Parting of Gold and Silver. Globe Works, Denver. FIG. 70. Cross-section of Fig. 69. SILVER 51 the Au remains behind as a powder. Apparatus necessary : glass, porcelain or stoneware vessels, less frequently platinum dishes. An apparatus for parting Au-bearing anode mud by the HNOs method is shown in Figs. 69-70. Sulphuric Acid Parting: applicable for Au-Ag alloys which contain not more than i oz. Au to 2-4 oz. Ag and less than 10% Cu. The H 2 SO 4 dissolves the Ag and Cu, forming sulphates : Ag 2 + 2H 2 SO 4 = Ag 2 SO 4 + 2H 2 O + SO 2 , if the granulated alloy is boiled with an excess of H 2 SO 4 . On a small scale, porcelain vessels are used, on a large scale, cast iron. The acid solution is diluted with water in lead-lined vats and clarified from small particles of gold thereupon the silver is precipitated on scrap copper or iron in other lead-lined vats. The precipitated silver is removed and compressed. For further treatment, see Silver Refining. For treatment of the gold, see Gold Refining (page 21). Chloridizing Fusion: conducting Cl into the melted Au-Ag alloy in order to convert the Ag into AgCl is no longer used. Electrolysis of Ag-Au-Cu alloys, see Copper. Solution of the Base Metals : In these methods the silver remains in the residue. According to the nature of the ore and the condition of working, one of the following processes is used: Freiberg Vitriolization Process: adapted to roasted matte. In contrast to the Ziervogel process the ore is so roasted that the Cu is converted to CuO and Ag to the metallic state. The CuO is dissolved as sulphate by treatment with H 2 SO 4 . The silver-bearing residue is smelted with silver- lead ores and the solution is worked up into blue vitrol (CuSO 4 -5H 2 O). Rossler's Sulphurizing Fusion. This consists in melting Au-Ag-Cu alloys with S. There remains an Au-Ag alloy, 52 METALLURGY Electrolysis Mobius Process. FIG. 71. Cross-section of Electro- lytic Cell. FIG. 72. Anode Cell with Five Anode Plates of Dore Silver. FIG. 73. Cathode and Supporting Device. FIG. 74. Modern Mobius Apparatus. Cathode, endless belt of sheet silver having below devices for greasing and brushing. This belt is drawn through a shallow vat. Anodes, bars of dore silver placed in shallow troughs having porous bottoms. SILVER 53 which is parted either by the sulphuric acid method (see above) or by electrolysis (see Silver Refining). The Cu 2 S, which contains some Ag 2 S, is converted to argentifer- ous copper by an oxidizing smelt and then electrolyzed. Hartz Vitriolization Process. This process converts precious metals carrying copper into, blue vitriol by the action of air and dilute sulphuric acid on the granulated alloy : H 2 SO 4 + O + Cu = CuSO 4 + H 2 O. The solution is clarified and converted by concentration into blue vitriol. The slime carrying the precious metal, after being washed and dried, is refined by cupellation. Electrolytic Refining: see Copper. Electrolytic Refining of Zinc Skimmings : see Zinc. (B) Refining Silver The aim in silver refining is to remove the last traces of impur- ities and to effect also the separation of the silver from other pre- cious metals. It consists usually of an oxidizing fusion in which oxidizing fluxes are utilized as well as the atmospheric oxygen, and, if other precious metals are present, of electrolysis. 1. Fire Refining of the crude silver either by melting under airblast, i.e., carrying out cupellation in small reverberatory furnaces with oxidizing flame, or FUSING WITH NITER, which is common in working up precipitated silver (cement silver) which carries small quantities of Fe or Cu (apparatus: crucibles), or by FUSING WITH Ag 2 SO 4 (Rossler's method), adapted to Ag carrying Bi. This process is also carried out in crucibles. 2. Electrolytic Refining. According to Mobius and Wohlwill, if gold-bearing silver (the so-called dore silver) in the form of small plates or bars, is used as the anode in an aqueous solution of HNO 3 (i%), AgNO 3 (0.5% Ag) and Cu(NO 3 ) 2 (4% Cu), the Ag will be dissolved and de- posited upon a sheet of pure Ag, which serves as the cathode. 54 METALLURGY The current density recommended is 30 to 20 amperes per sq.ft. The E.M.F. is 1.5-1.4 volts. The apparatus is shown in Figs. 71, 72, 73, and 74, page 52. Properties of Refined Silver : SPECIFIC GRAVITY: 10.5. FIG. 75. FIG. 76. Fused Silver. Dendritic Surface (X33). COLOR: white with brilliant lustre. DUCTILITY: tough, very ductile. MELTING-POINT: 961 C. (1762 F.). VAPORIZED readily at 1200-1500 C. (2192-2732 F.). SILVER 55 THERMAL AND ELECTRICAL CONDUCTIVITY best of all metals ALLOYS WITH MOST METALS: the most important metal- lurgical alloys are those with Au, Pb, Hg, Cu and Zn. ^^ ^&Mim FIG. 77. Electrolytic Silver (Mag. 16). CHEMICAL BEHAVIOR: not very active, although its solution tension is greater than that of gold. Silver compounds are easily dissociated and are sensitive to light (photography). As solvents for metallic Ag, concentrated HNO 3 and H 2 SO 4 are used. For AgCl other metallic chlorides and " hyposulphite " are employed. MERCURY Sources Natural Sources : FREE, usually containing Ag and less frequently Au. As CHLORIDE, HgCl, calomel (rare). As SULPHIDE, HgS, cinnabar, the most important ore of mercury. It is finely disseminated through the gangue and the per cent of mercury present is usually low. With other sulphides in tetrahedrite. Other Sources : AMALGAMS resulting from the use of mercury in extracting other metals or from contact with metals in other ways. RESIDUES and unmarketable products produced in making mercury pigments (HgO and HgS). (A) Extraction of Mercury Roasting Accompanied by Distillation : Oxidizing Roast is the method commonly used in treating mercury ores. Since the dissociation temperature of HgO is very low (400 C.), Hg results directly from roasting at high temperatures : APPARATUS: Shaft furnaces for lump or briquetted ores. (Figs. 78-79.) SELF-FEEDING ROASTERS for finely-divided ores. (Figs. 80-83.) REVERBERATORY FURNACES for easily reducible or sintering ores. To all of these furnaces, a system of iron or clay 56 MERCURY 57 CONDENSATION TUBES is attached, these tubes being partly cooled by air, and partly by water. The products of these processes are: LIQUID MERCURY, which collects under .water in the lowest part of the condenser. MERCURIAL DUST or SOOT, made up of flue dust from the FIG. 79. ore and sublimated mercury compounds (HgO, HgS, HgSO 4 ) which collect on the walls of the condenser and mechanically retain globules of mercury. 58 METALLURGY MERCURY 59 Heating with Desulphurizing Fluxes (Fe and CaO), is used only for especially rich ores, or for residues from the manufacture of cinnabar. For this purpose retorts are used such as have been already described under Amalgam Dis- tillation (see Silver). (B) Mercury Refining Filtration of the liquid mercury that collects in the condensing tubes is frequently sufficient. The use of presses is necessary in working up the " dust " or " soot." Distillation is used for amalgam (see Silver). Washing with Acids is employed when the mercury is to be used for scientific purposes. Properties of Mercury : SPECIFIC GRAVITY: 13.5. COLOR : bluish- white. TENACITY: liquid at ordinary temperatures. FREEZING-POINT: -39.4 C. (-38.9 F.). VOLATILIZES slowly at ordinary temperatures. BOILING-POINT: 360 C. (680 F). ELECTRICAL CONDUCTIVITY: about 0.017 tnat f Ag. ALLOYS: form readily with Au, Ag, Cu, Pb, Sn, Cd, Zn, with the alkalies and alkaline earths. More difficultly with Ni, Co, Fe, Mn, Cr, W and the earths. CHEMICAL BEHAVIOR: The affinity of Hg for the halogens and for S is great; slight for oxygen. The chlorides HgCl and HgCl 2 , as well as the sulphide HgS, may be sublimed without decomposition. HgO first forms at about 300 C., but dissociates very readily at 400. The best solvents for Hg are HNO 3 and aqua regia. Hg may form a univalent as well as a bivalent cation in solution. COPPER Sources Natural Sources: NATIVE COPPER: surrounded or accompanied by sulphide or oxide copper ores, clay and shale. RED COPPER ORE, cuprite, Cu 2 O, carrying 88.8% Cu. Asso- ciated with sulphide copper ores, spathic iron ore, clay or shale, and earth. BLACK COPPER ORE, thenorite, CuO (rare). COPPER GLANCE, Cu 2 S, free but not widely distributed. COPPER PYRITES, chalcopyrite, Cu 2 S.FeS.FeS 2 . This, the most important copper ore, contains 34.6% Cu. It is associated with galena, blende, shale and slate. PEACOCK ORE, bornite, (Cu 2 S) 3 .FeS.FeS 2 . Not so abun- dant as chalcopyrite. BLUE VITRIOL, CuSO 4> 5H 2 O. Formed by the weathering of sulphide ores and associated with their gangue. MALACHITE, or green carbonate of copper; HO-Cu-COs-Cu- OH; formed by the weathering of other copper minerals. Associated with similar minerals and in similar localities. AZURITE, or blue carbonate of copper; HO-Cu-COs-Cu-COy- Cu-OH; associated with malachite. Other Sources: MATTE containing copper, speiss, slag and alloys from lead and nickel smelting. SCRAP METAL and residues from metal-working processes. (A) Enriching and Concentrating Processes The percentages of Cu, as given above, refer to the pure mineral, free from gangue. By means of the gangue, however, the Cu content of most ore bodies is reduced to a few per cent. From 60 COPPER 61 these ores the copper can neither be successfully leached out by chemical means nor can it be suf- ficiently concentrated by me- chanical processes so that the ores can be successfully fused directly for metal, as this would be so impure that refining would be out of the question. It is neces- sary, therefore, to effect a chemical concentration by a fusion. Of all the metals that come into considera- tion here, copper has the greatest solution tension in fused sulphides; it will, therefore, withdraw the sulphur from other metals in so far as the total amount of sulphur present in a fused mixture of sulphides and other compounds is insufficient to form the lower sulphides of all the metals. Upon this principle are based the following Roasting and Smelting Operations : i. An Oxidizing Roast effects the removal of the sulphur from sul- phide ores, leaving, however, more than 62 METALLURGY enough, after allowing for a loss due to reaction in the furnace, to form Cu 2 Swith the copper present. In matte smelting, there SECTION A-B FIG. 85. Stall Roasting Plant (Peters). Scale, i : 150. must always be enough sulphur present to form with the copper and to form a certain amount of FeS, to prevent a high loss of copper in the slag; the FeS and Cu 2 S COPPER 63 should be present in such quantities as to correspond to the formula FeS.Cu 2 S. In crushing the ore, which is usually in the form of large lumps to make it suitable for roasting, it should be observed that the finer the ore the greater the cost of roasting, the greater the loss in flue-dust, and the greater the hindrance in subsequent leaching operations. The size of grain desirable in roasting is from 6 to 60 mm. In choosing a method of crushing, one that is quick and gives the smallest amount of fines should be selected. Hand breaking has advantages, for the fines are then under 10% as a rule. Machines which produce the smallest quantity of fines usually require the most power. If the nature of the ore requires that it be reduced to a small grain, rolls, stamp mills and ball mills are used. The following processes and and apparatus may be mentioned. HEAP ROASTING: Open heaps of ore having a trapezoidal or hemispherical cross-section are constructed upon some combustible material in order to start the roasting. (See Fig. 84.) STALL ROASTING : To serve as a protection against the weather and to prevent damage to the neighborhood, walls may be built to protect the heap. These walls at first consisted of board fences, earth banks, and finally stone walls, from which came the so-called " stall." Stall roasting is essentially the same as heap roasting except for the protective walls, which also serve to regulate the amount of air. The best stall construction is Peters' modifica- tion of the Swedish stalls. (See Fig. 85.) PYRITE BURNERS: these may be regarded as a further step in stall construction. Many different forms have been developed, among which may be mentioned: kilns (Fig- 86), lump pyrites burner (Fig. 87), fine pyrites burner, Gerstenhofer's fine pyrites furnace, Hasenclever-Helbig furnace, Olivier-Ferret furnace and the Maletra-Shaffner roasting furnace (Fig. 88). 64 METALLURGY COPPER 65 REVERBERATORY ROASTING. As compared with other reverberating furnaces, the following points are character- Fic. 89. Stetefeldt Shaft Furnace. istic: They should combine a small firebox with a rela- tively long hearth, and the length of the latter depends upon the amount of heat evolved from the roasting ore. Ores 66 METALLURGY carrying 10% or less of sulphur are seldom roasted in reverberatories. The relation between length of hearth and per cent of sulphur is as follows: With 10% sulphur, length of hearth 15 ft. With 15% sulphur, length of hearth 30 ft. With 20% sulphur, length of hearth 45 ft. With 25% sulphur, length of hearth 60 ft. FIG. 90. MacDougall Furnace. FIG. 91. Herreshoff Furnace. Hearth-furnaces worked by hand are designated as hand reverberatories. Among the SHAFT FURNACES the Stetefeldt furnace should be mentioned. It is especially suitable for chloridizing roasting. (Fig. 89). The most important reverberatories with mechanical stirrers are the following: PARKES FURNACE: this has two circular superimposed hearths with radial stirring arms attached to a central shaft. The firebox is beside the lower hearth. COPPER 67 MACDOUGALL FURNACE (Fig. 90), Herreshoff furnace (Fig. 91) and the furnaces of Humboldt and Kaufmann: these have five to seven circular hearths placed one above the other. Horizontal arms, extending from a vertical shaft,, work the ore in such a way that on the upper hearth it is slowly moved toward the periphery. Here it falls to the next hearth, is worked toward the centre, falls to the third: hearth and so on to the last. The arms are easily removed and replaced by new ones. O'HARA FURNACE: In its original construction this furnace consisted of two superimposed hearths through which rakes resembling ploughshares were drawn by means of an endless chain. The latter was run upon sprocket wheels. During interruptions in the process, the chains should be dropped into a groove in the hearth in order to protect them from the hot gases. This apparatus is untrustworthy and suffers greatly from hot gases. ALLEN FURNACE: Allen obviated the above difficulties by placing tracks in his furnace upon which carriages carry- ing the rabbles were supported. The carriages were drawn by the chain as before. BROWN FURNACE: Brown placed the tracks and supporting carriages in compartments built in the side walls of the furnace. Brown's furnace is also constructed in the form of a broken circle, the so-called elliptical furnace, with external fireboxes. (Fig. 92.) HIXON FURNACE: Hixon finally placed the supporting trucks and the dragging device (iron cables instead of chains) in a channel built in the bottom of the furnace. HOLTHOFF-WETHEY FURNACE : the rails and trucks for the tube-like rabble holders were placed entirely outside the walls of the furnace. The space beneath the hearth served mainly as a cooling hearth. PEARCE FURNACE (see Gold, page 3, Figs, i and 2): The hearth is circular. The rabbles are held by tube-like arms through which air for roasting is supplied. The 68 METALLURGY FIG. 92. O'Hara-Brown Elliptical Furnace. COPPER 69 rabble arms are driven from a central shaft and project through a slot in the inner wall of the furnace. The slot is kept closed by an iron ring which moves with the rabbles. The Pearce furnace is built both with a single hearth and with two hearths one above the other. Auxiliary heating, if necessary, is supplied by external fire boxes. The space under the roasting hearth serves as a dust chamber. WETHEY FURNACE: This is a furnace for fine pyrite and consists of four superimposed hearths. The rabbles make a circuit through two hearths. Two of these furnaces {- -3658- FIG. 93. Bruckner Furnace. Scale 1:150. may be placed side by side and the rabbles moved by the same, machinery. SPENCE FURNACE: This is a furnace for treating fine pyrite. It consists of four superimposed hearths, the rabbles being drawn back and forth over the hearths by means of horizon- tal rods. KELLER FURNACE: This is a furnace for fine pyrite. It has five superimposed hearths and its operation is similar to that of the Spence furnace. Two furnaces are usually placed side by side and operated by the same engine. Among furnaces with moving hearths the following have found favor. 70 METALLURGY BRUCKNER REVOLVING FURNACE (Fig. 93) : This furnace, which is widely used, consists of a cylinder resting upon friction wheels to which power is applied by means of toothed gears. The heated gases from an external fireplace pass axially through the cylinder. At the opposite end are the dust chamber and stack. The cylinder is charged from hoppers placed above manholes in the furnace walls, the holes being subsequently closed. After the roasting is completed the furnace is discharged through the same holes into cars. WHITE-HOWELL FURNACE (Fig. 94): The roasting hearth consists of a slightly-inclined cylinder, open at both ends. FIG. 94. White-Howe 11 Furnace. It is made of sheet iron and lined with firebrick. The furnace is fed through a tube at the higher end and is discharged into a walled chamber at the lower end. Beside this chamber is the fireplace from which the hot gases pass over the roasting ore. At the higher end is a dust chamber from which leads the flue to the stack. HOCKING- OXLAND FURNACE : This is similar in construction to the White-Howell, but somewhat differently built in the arrangement of the firebox and the chamber for collect- ing the roasted ore. ARGALL FURNACE (Fig. 95): Within a broad cylinder are enclosed four smaller, somewhat shorter cylinders slightly inclined. The large cylinder is fitted with firebox and COPPER 71 dust chamber similar to those previously described. The ore is fed into each narrow cylinder in turn as it reaches its highest point during the revolution of the broad cylinder, and is discharged at the opposite end, which lies next to the firebox. A chamber below the large cylinder receives the ore. MUFFLE FURNACES: These are used for the chloridizing roasting of copper ores, especially roasted pryite carrying copper. The construction, shown in Fig. 96, page 72, has four iron charging tubes with gates for closing. The charge is worked by hand through doors in the side of the muffle. FIG. 95. Argall Furnace. Scale, i : 150. The gases are conducted away through an iron tube near the firebox. In order to lessen the danger of overheating the charge, the muffle arch is made of double thickness. 2. Smelting for Matte consists in uniting the copper of the roasted product with part of the sulphur, uniting the rest of the sulphur with part of the iron and slagging the remaining iron together with other materials. In order to prevent high slag losses, it is best not to attempt too great a concentration of copper. A matte carrying 50% copper should be regarded as the maximum and in most cases a lower grade is produced. The concentration of the ore to matte obtained by these processes fluctuates widely, viz., between 12:1 and 3:1 or, in other words, between 12 and 3 tons of ore yield i ton 72 METALLURGY COPPER 73 of matte by the process of roasting and smelting. In roasting attention must be paid to these limits. The roasted product contains a mixture of the oxides, sul- phides, sulphates, arsenates, antimonates, and silicates of the various metals contained in the ore. PRINCIPLES OF THE SLAG CALCULATION: The slag should lie between a singulo- and a bi-silicate, preferably in the vicinity of a singulo-silicate with FeO as the predominating base. Exceptions: With high iron and zinc content: between a singulo and a sub-silicate. With high SiC>2 content : between a bi-silicate and a tri-silicate. It should be borne in mind that the furnace temperature must be raised if high silica is used, thus increasing the danger of reducing the iron compounds : FeS + Cu 2 = Cu 2 S+Fe. APPARATUS FOR MATTE SMELTING: i. SHAFT FURNACES: From the original narrow shafts with rectangular or trapezoidal cross-sections, a change has been made in favor of wider furnaces with circular, oval, or long rectangular cross-section, and recently to six-, seven- and eight-sided sections. The height of the shaft depends upon the iron content of the ore and the nature of the fuel. The height diminishes as the iron content increases. Furnaces using charcoal should be made higher than those using coke. High zinc necessitates a low furnace, as otherwise deposits of zinc oxide will be formed. The chemical action of the matte and slag have led to the adoption of cooled furnace walls. Whereas in lead smelting only a relatively small portion of the wall is cooled, in copper smelting not only the whole bosh but in some cases the entire shaft is surrounded by a water-jacket. For these relatively large cooling plates, cast iron should be avoided as much as possible. Besides wrought iron, copper is 74 METALLURGY FIG. 97. Early Type of Shaft Furnace. COPPER FIG. 98. Early Type of Shaft Furnace. Scale, i : 150. FIG. 99. Oker Furnace in the Harz. 76 METALLURGY sometimes used, especially for the inside walls. To lessen the danger of interrupting the process by an accident to the water-jacket, the latter should never be constructed of one piece, but in segments which may be independently removed and replaced. The first requisite for the use of a water-jacket is an FIG. 100. Allis-Chalmers. FIG. 101. Colorado Iron Works, Denver. Scale, i : 150. adequate supply of water. If this is available, there is no danger connected with the process. In construction, operation and repairing, the water- jacket furnace is to be preferred to a brick furnace. With good management the cost of operation is about the same for each. Great care should be taken with brick furnaces in the choice of bricks, and of mortar, and in the construction. Water-jacket furnaces are easier to design, also easier and quicker to build. Brick furnaces also require great care and attention when blowing in. If the blast pressure is too high, great difficulty may be encountered, such as burning out the COPPER 77 hearth and shaft walls. If the blast pressure is too low, wall-accretions and sows may form. Scarcely any difficulty is experienced in blowing in a water- jacketed furnace. With similar handling, the relation between the repair costs of brick and water-jacket furnaces is 2:1. Cracks FIG. 102. American Water-jacket Furnace (20 Tuyeres), Colorado Iron Works. Scale, i : 150. and thin places formed in the walls of brick furnaces tend to grow worse instead of better, for even if they are filled with clay, the matte and slag tend to eat under the clay and widen the crack. If in order to remove accretions it is necessary to lower the charge, it is often found in brick furnaces that the 78 METALLURGY FIG. 103. Sticht Shaft Furnace, Mount Lyell M. & R. Co., Tasmania. Side View and Longitudinal Section. COPPER 79 accretions stick even more firmly to the walls than before, .because of the fact that the upper part of the walls of the FIG. 104. Sticht Shaft Furnace, Mount Lyell M. & R. Co., Tasmania. End View and Cross-section. empty shaft have become very hot. In the water-jacketed furnace, however, the accretions can be readily removed, for the walls remain cool. 80 METALLURGY FIG. 105. Johnson American Water- jacket Furnace (16 Tuyeres). Side Elevation. FIG. 106. Johnson American Water- jacket Furnace (16 Tuyeres). Side Elevation. COPPER 81 SLAG AND MATTE SPOUTS : In small furnaces these are made of brasque, in large furnaces of cooled plates formed by fastening wrought-iron tubes, about one inch in diameter, to cast-iron troughs. (Hixon's method of building Ltir- mann's slag spout.) FORE-HEARTHS. To lessen undesirable reactions which result in the formation of furnace " sows," copper losses and other difficulties, it is of great importance in copper blast-furnace smelting that the matte be removed as soon as possible from the furnace. The furnaces are therefore almost always built with external crucibles so that the separation of matte and slag may take place in the fore-hearth. The first fore-hearths were made of brasque in the form of holes in front of the furnace. (See Figs. 97 and 98, PP- 74, 750 The modern fore-hearths consist either of portable iron pots or larger boxes (Figs. 104 and 106, pp. 79, 80), or of special reverberatory furnaces for acid slags. The smaller pots are lined with clay and straw, the larger ones are lined a half course of firebrick. If the fore-hearth is very large the lining is omitted entirely. The first slag chills on the sides of the hearth and the heat radiated from the walls so regulates the temperature that the slag remains as a lining. A diameter of 10 ft. is the upper limit for brick-lined hearths. Watercooling is resorted to in some cases for the whole wall of the hearth as well as for special parts, for example the matte tap, the slag spout, and in the so-called Orford fore-hearth for the partition which separates the matte from the slag. According to Peters the use of reverberatories as fore- hearths possesses the following advantages: (a) Economy in the cost of remelting for subsequent treat- ment of the matte. (b) Reduction of the necessary temperature required in the blast furnace and thus saving of fuel. METALLURGY (c) Prevention of overdriving the blast furnace, for if little matte is required at any period of the subsequent opera- tions, the time can be utilized in accumulating a small stock of especially rich or poor matte, casting it into bars of definite size and laying it to one side. A means is thus provided for regulating the amount and composition of subsequent charges in the reverberatory. (d) Assistance in the separation of matte and slag. In the unheated fore-hearth, the temperature is con- tinually falling, while in the reverberatory fore-hearth the products of the blast furnace are run into a heated chamber, which favors the separation of matte and slag. The reverberatories used as fore-hearths do not differ essentially from other copper reverberatories. The following features, however, should be provided: (a) The matte from the blast furnace should be introduced at one of the ends. (b) The slag should be withdrawn from the opposite end. (c) The fireboxes are, therefore, on the sides and should be arranged so as to close easily. (d) The molten matte should flow as directly as possible from the fore-hearth to the converters. (e) Sufficient space should be provided to allow 50 tons of matte to be withdrawn and stored in case of an inter- ruption in the subsequent processes. The matte is withdrawn from the reverberatories either into a sand bed where it cools and is broken into lumps, or else it is run into casting ladles for delivery in the molten state to the next apparatus. When a reverberatory fore-hearth is not used, the slag is allowed to flow through small fore-hearths (slag pots) in which a small quantity of matte settles. From these slag pots it is withdrawn to be made into building stone if sufficiently acid, or it flows into slag cars for transpor- tation to the dump. If the location and water supply is COPPER 83 ,84 METALLURGY favorable, it is granulated in a stream of water and by it sluiced to the dump. In smelting for matte, and also in the process of matte concentration to be described under (4), there is oppor- tunity to add any oxidized ore and smelter products (slag rich in copper) that are at hand, and work them up at the same time, but care must be taken that in the roasting or in the mixing of the charge enough sulphur is present, or made up by the addition of unroasted sulphide ores, to provide enough sulphur for all the Cu that is in the form of oxide. 2. REVERBERATORY FURNACES FOR MATTE SMELTING are used at several American smelting plants on a very large scale. The reverberatory process is used principally for finely divided ores and roasted products. According to Peters (Metallurgie, 2, 9, 1905), the grate, hearth, flues and stack must be so designed that a tem- perature of 1700 C. can be developed if from 1400 to 1500 is required to fuse the charge. The firebox and hearth must be of such size that fresh additions of coal or ore will not cause a great fall in the temperature. The hearth made of fused sand must be kept covered, to a depth of at least 8 in., with fused matte to resist the action of the fresh charge which bakes on to the sides and of the oxides that form during the operation. The hearth should be of such length as to best utilize the heat and give the cleanest possible separation of matte and slag. The following is a good example of modern American furnaces (Matthewson) : Grate: 6x16 ft. Hearth: 20x100 ft. (2000 sq.ft.). Flues: 3} feet wide. Sand bottom: 3 ft. thick, burnt in one layer. Fuel required: 57 tons per 24 hours, of which 6J tons are recovered by treating the ashes in jigs. Ore treated averages 275 tons per 24 hours with a maximum of 330 tons. It is taken red hot from the roasters every 80 minutes in 1 5-ton lots and dumped through four hoppers COPPER 85 upon the first 20 feet of the hearth back of the fire bridge. The molten matte, 8 in. deep on the hearth, corresponds to 60-80 tons. When matte is needed in the converter department, it is drawn into ladles, up to 10 tons at a time. Every 3 or 4 hours 30-40 tons of slag are withdrawn, FIG. 1 08. Mansfeld Furnaces. requiring 15 minutes for the operation. The slag is granu- lated and sluiced to the dump. The process requires little hand labor, as the charge, dumped in the hottest zone and floating upon the matte, distributes itself, and then flows 80 ft. before reaching the taphole. After the large mass has been melted down, the roof and side walls absorb so much 86 METALLURGY heat that at the end of 80 minutes a new charge can also enter the now highly superheated chamber. The process thus becomes continuous as with the blast furnace. FIG. 109. American Reverberatory Furnace (Peters). Scale, i : 150. Since the reverberatory furnaces for the different proc- esses in copper smelting (matte smelting, black copper smelting and copper refining) are of similar design, although COPPER 87 of different dimensions, they are shown together on pages 85-87 (Figs. 108-110). 1. Oxidizing Roasting and 2. Smelting for Matte are often carried on in one operation. This is exemplified in the old process, kernel roasting, and in pyritic smelting, a process brought to high efficiency during the past ten years in Australia and in the United States. Kernel Roasting: This is a prolonged heap roasting, the pur- pose being to concentrate the copper at the centre of the slowly roasted lump of ore. By this method of roasting there takes FIG. no. Mathewson Furnace, Washoe Smelter, Anaconda. Scale, 1:500. place at the contact of the oxidized surface and the unchanged sulphide a reaction similar to that which occurs in matte smelting. The copper is taken up by the unaltered sul- phide, which results in a constant enrichment of the centre in copper, leaving a crust composed of iron oxide. At the temperature maintained in this process, part of the copper is unavoidably converted into CuSC>4, and remains as such in the crust. The kernel can be enriched to about five times the copper content of the ore; the crust usually contains about 3% Cu as CuSO 4 and i% as CuO. Pyritic Smelting (Metallurgie, 3, 1906). True pyritic smelt- ing consists of an oxidizing smelt without using any fuel 88 METALLURGY except that contained in the ore itself (Fe and S). It is used for pyrite carrying copper and precious metals, but with not too much lead (2, etc.) or in which the Cu2S is combined with other sulphides (FeS, FeS2, etc., in chalcopyrite and bournite) are only slightly attacked by the customary solvents. Double sulphides may be converted by roasting at a low temperature (450-500 C.) into insoluble Fe 2 Os, soluble Cu2O, CuO, CuSO 4 and unoxidized, but free, Cu 2 S. Moreover, according to a recent investigation by O. Frolich, the double sulphide may be decomposed by heating to about 200 into the simple free sulphides without forming any oxides: Cu 2 S.FeS = Cu 2 S + FeS (cf. Metallurgie, 1908, 5, 206). The following solvents are used : Water for Sulphates. Since in weathering and roasting some basic sulphates are formed, the water is usually acidulated if it is not directly obtainable as acid mine water. Hydrochloric Acid is used ordinarily in connection with other salts, e.g. (NaCl), to increase the solubility of cuprous and silver chlorides. At the Stadtberge works in Lower Marsberg, 150 tons of acid schist, carrying 2% Cu as carbonate and as copper glance, were formerly leached daily as follows: The ore, crushed to 6-15 mm., was placed in shallow trenches which had been made water-tight with clay and lined with boards. Here it was treated for 8-10 days with a solution of acid and NaCl, whereby copper amounting to 0.5% of the weight of the ore was dissolved. The residue was then allowed to stand in the air for from 8-10 weeks in well venti- lated heaps and again leached with salt solution (liquors containing FeCl 2 waste from the precipitation vats). This dissolved i% Cu, based on the weight of the ore. After the second leaching the ore was transferred to the dumps, where 92 METALLURGY it was again subjected to the action of the atmosphere. Here almost all of the remaining copper was leached out by the action of the rain, the solution being removed in drainage trenches and conducted to the precipitation appar- atus. The reactions in the three stages of the leaching are as follows: The acid dissolves the free carbonate. The ore soaked with free HC1, FeCl 2 and FeCl 3 , as it remains in the heaps and later on the dumps, undergoes the following changes : (1) Cu 2 S + 2FeCl 3 = 2CuCl + S + 2FeCl 2 . (2) 2 FeCl 2 + 2HC1 + O = H 2 O + 2 FeCl 3 . The Feds thus formed reacts again as under (i). The CuCl may also be converted to the higher chloride : (3) 2 CuCl + 2HC1 + O = H 2 + 2 CuCl 2 . The latter now works in a similar way to Feds. (4) Cu 2 S + 2CuCl 2 = S +4CuCl. Sulphuric Acid. The most important cases where sulphuric acid is utilized have already been described under Silver. Ferric and Cupric Salts. The action of ferric and cupric chlor- ides has already been illustrated under hydrochloric acid. Ferric sulphate acts in a similar way, but cupric sulphate does not, because cuprous sulphate is so unstable that it is not formed under the conditions that prevail in leaching. In the Rio Tinto district (Fig. 112), where pyrite occurs with free Cu 2 S, the ore is arranged in long, terraced heaps 28 ft. high, placed side by side and provided with air passages. The heaps are sprayed from time to time with acid mine water, to leach out the CuSC>4 and to assist in the sulphatizing action. The ferrous sulphate and sulphuric acid contained COPPER 93 in the mine water and in the heaps apparently come from a part of the pyrite: (i) If a considerable quantity is present, it may be oxidized to ferric sulphate: (2) FIG. 112. Rio Tinto Leaching Plant. The Fe 2 (SO 4 ) 3 then reacts with the Cu 2 S to some extent as follows: (3) 2Fe 2 (SO 4 )3 + Cu 2 S = 2 CuSO 4 + S +4FeSO 4 , A or, together with the atmospheric oxygen and moisture, (4) 2 Fe 2 (SO 4 ) 3 +Cu 2 S+H 2 O+3O = 2 CuSO 4 + 4FeSO 4 +H 2 SO 4 . In any case so much FeSO 4 is formed, as compared 94 METALLURGY to the amount of H 2 SO 4 present, that under a very active oxidation much basic sulphate results and a considerable quantity of hydrated ferric *oxide forms on the surface of the ore. The removal of the copper takes place so rapidly that much FeS2 remains unaltered and the residue from leaching, still containing 49 to 50% S, may be used in sulphuric acid manufacture. Ferric chloride leaching was at one time introduced into the Rio Tinto district, but the practice has been abandoned (Db'tsch Process). Ferrous Salts, especially FeCl 2 , have been recommended by Hunt and Douglas in leaching ores which contain copper as oxide or carbonate: 3CuO + 2FeCl 2 = Fe 2 O 3 + CuCl 2 + 2CuCl. As an argument in favor of the process it was pointed out that CuCl solution requires less iron for precipitating the copper than CuCl 2 or CuSO 4 solutions. Ores which cannot be leached directly are prepared under some circumstances by subjecting them to a chloridizing or a sulphat- izing roast. In both cases an oxidizing roast is carried out; in the first case chlorides (MgCl 2 , NaCl, etc.) are added and chlorine is made available by the aid of SO 2 and O, and in the second case the temperature is kept as low as possible (450 to 500 C). A PURIFICATION OF THE LEACHED LIQUOR is necessary if the plant also sells copper as blue vitriol. The chief impur- ity is FeSO 4 , together with small quantities of Sb, As, etc. All these are substances easily precipitated if copper oxide (roasted matte) is introduced while air is being blown through the hot solution for the purpose of oxidizing FeSO 4 to Fe 2 (SO 4 ) 3 . 2CuO + 2FeSO 4 + O - 2CuSO 4 + Fe 2 O 3 . This process was brought to a high state of efficiency by Ottokar Hofmann in Kansas City (Metallurgie, 1907, 4, 582). COPPER 95 (B) Extraction of Copper from Intermediary Products In the concentration processes described under A, the copper remained either in the form of sulphide low in iron (Cu2S) often carrying precious metals, or in the form of an aqueous solution. The concentrated matte is commonly converted into copper by smelting processes, among which the so-called Reaction Process is most commonly employed, although the Roast-reduction Process has found favor if a matte containing silver is to be desilverized by leaching between the roasting and smelting operations. If the copper has been concentrated in the wet way, obviously only precipitation methods come into consideration. Reaction Smelting is carried out either in reverberatories or in converters, and causes the separation of copper according to one of the following reactions : With the sulphides of base metals, the oxidation always takes place so that the metal oxide and SO 2 are formed: Cu 2 S + 30 = Cu 2 O + SO 2 . The higher oxide is not formed in the presence of reducing agents (Cu 2 S),but the latter, on the contrary, reduces the Cu 2 O: 2 Cu 2 O + Cu 2 S - 3Cu 2 + SO 2 . The reverberatory furnaces used for this process are not so large as the newer smelting furnaces, but like the latter they have a sand bottom and are provided with tuyeres beside the firebox for blowing air upon the metal bath (cf. page 85-87). In 1880 Manhes in France and one of his assistants at the Parrot Copper Company of Butte, Montana, succeeded in blow- ing copper matte directly to metallic copper in a converter similar to that used in the Bessemer process. The success was due to a change in the method of admitting the air. The previous failures had been due to using a converter of exactly the same pattern as 96 METALLURGY that used in steel manufacture without taking into account the different conditions governing the working up of the raw materials used in copper smelting. In the copper Bessemer process a matte carrying 40 to 55% Cu is commonly used. In comparison with the iron Bessemer process (see Iron) the following differences may be noted. Copper Matte. Iron. 45 to 60% oxidizable substances of low calorific power. 3 to 6% impurities of power. high calorific Formation of large amounts of strong- ly basic FeO which slags with the converter lining. Quantity of slag small. Converter lining soon destroyed (9 charges). Converter lining lasts (200-250 charges). a long time Three different materials: slag, con- j Two materials: small amount of stantly increasing; matte, constant- ly decreasing; metal, in the second stage constantly increasing. Large amount of metal. Copper: good conductor of heal. Specific heat 0.155 (liquid, see Glaser, Metallurgie, 1904, I, 126). Iron: poorer conductor. Specific heat 0.1665 (liquid, see Oberhoffer. Metallurgie, 1907, 4, 495). Only when Manhes arranged the tuyeres so that the air no longer passed through the separated metallic copper was the blowing successful. MANHES' CONVERTER. This consists of a cylinder on a horizontal axis; the tuyeres are arranged in a row, in the cylinder mantle, being placed parallel to the axis, so that by rotating the drum the tuyeres can be made to enter the matte at different depths. This cylindrical type has been successful at both the Anaconda and Copper Queen (Bisbee) plants. Fig. 113 shows a modern arrangement of such a plant. STALMANN constructed a converter of rectangular cross- section. These converters have been introduced by Sticht at COPPER 97 FIG. 113. Horizontal Converters at the Anaconda Plant. FIG. 114. Converter Department, Mount Lyell M. & R. Co., Tasmania. METALLURGY the large plant of the Mount Lyell Mining and Railway Co., Australia. Fig. 114 shows the converter department of these c/: works. Figs. 115, 116 and 117 also show sketches of the converters as made by Mr. Sticht. Upright converters, resembling the Bessemer type, have been successfully used at the Parrot and Anaconda plants. COPPER 99 The tuyeres are placed in a semicircle above the bottom on one side of the converter (Figs. 118-121, p. 100). The converter lining, consisting of 17-28% clay and 83-72% quartz (gold bearing), is tamped into a thickness of about 1 8 in. The blowing takes place in two stages: (i) Oxidation of the FeS and slagging of FeO, usually requiring 30-50 minutes, and (2) blowing to separate the copper, requiring 30-50 minutes more. The Roast-Reduction Process : This consists of the following operations: i. OXIDIZING ROAST (dead roast) for the purpose of removing FIG. 117. Stalmann Type of Converter at Mount Lyell. Designed by R. Sticht. as much of the S as possible and converting the Cu into CuO. (Apparatus: see A, p. 61 to 71.) 2. REDUCING SMELT in a shaft furnace or a reverberatory, with other rich copper-bearing materials if desired. This is very successfully used at the Mansfeld works in conjunction with the Ziervogel process. Reverberatory furnaces are used at this plant. Precipitation of Copper from solutions of the chloride and sul- phate is almost always accomplished by iron. In the Stadt- berge Works (page 91) wooden vats are used in which the iron in the form of thin turnings rests upon a wooden frame 100 METALLURGY and the solution is agitated by wooden paddles. At Rio Tinto the copper is precipitated in trenches in which cast iron bars have been placed. (Fig. 122). FIG. 11 8. FiG.119. FIG. 120. Converters at the Anaconda Plant. FIG. 121. The precipitated copper is treated in jigs to remove large pieces of iron and in washing drums to remove COPPEk 101 fine particles of iron and iron oxide. It is then filtered and dried. If the solution carries silver, part of the copper (one-third to three-eighths) is first precipitated, carrying with it the silver, then the remaining copper is precipitated, silver m FIG. 122. Rio Tinto Precipitating Trenches. free, in a second vat. The two lots are treated differently (see below). (C) Copper Refining Pure copper, or refined copper, is made to-day almost entirely from crude copper. The attempts to produce pure copper directly from the ore have thus far proved a failure, and the same is true of most attempts to produce pure copper from matte. It is only recently that the difficulties hitherto met with in the latter process have been overcome. 102 Crude copper may contain the following impurities: S, As, Sb, Zn, Pb, Fe, Ni, Co, Ag, Au. If precious metals are not present the following process may be used : Furnace Refining consists of an oxidizing followed by a reduc- ing fusion. 1. OXIDIZING FUSION. A reverberatory furnace is usually used for this purpose, the object being to eliminate the oxidizable impurities from the copper. The S, As, Sb, Zn, and Pb are partly volatilized as oxides and partly slagged; any Fe and Ni are also slagged. The S is present in impure copper largely as Cu 2 S, and is removed only after the greater part of the other impurities has been eliminated. It is oxidized principally through the agency of Cu 2 O after a certain quantity of the latter has been dissolved in the copper. When this point is reached, the evolution of SC>2 takes place rapidly (boiling); when the reaction has nearly ceased, a wooden pole is plunged into the metal. The gases escaping from the submerged end of the pole agitate the metal, thus allowing the SO 2 to escape more easily and at the same time the spurting copper is further oxidized (dense poling). The copper resulting from this treatment contains much Cu 2 O (set copper). The result- ing slag runs high in copper and goes back to the smelting furnace. Set copper is very brittle and unfit for the ordinary purposes for which copper is utilized. It is further refined as follows: 2. REDUCING FUSION. The molten copper is heated with a reducing flame and kept covered with charcoal; a wooden pole (green wood) is introduced and poling con- tinued until the dissolved Cu 2 O has been reduced by the hydrocarbons in the pole and the charcoal in the covering. This process is complete when a thin test-bar can be bent through 1 80 without showing a crack. When broken by repeated bending, the fracture must show a fibrous, salmon-red texture with silky luster. COPPER 103 The refined copper is cast from wrought-iron ladles, which are coated with chalk or clay, into cast iron or copper CASTING MACHINE, WALKER SYSTEM. FIG. 124. molds. In large plants it is run directly from the furnace through troughs into mechanically dumping molds (Figs. 123-124). 104 METALLURGY Electrolytic Refining is almost universally used for copper containing precious metals. It is carried out in ELECTROLYZING VATS made of wood but lined with lead. FIG. 125. Anode Cathode I FIG. 126. FIG. 127. THE ANODES are cast plates of unrefined copper (Figs. 125, 126, and 127) that contains precious metals. THE CATHODES are thin sheets of copper made by deposit- ing the metal electrically on lead or greased copper plates from which it can be easily removed. COPPER 105 THE ELECTROLYTE is an acid solution of copper sulphate containing 12-15% CuSO 4 .5H 2 O and 5-10% H 2 SO 4 . DURING THE PROCESS, it is important that the constantly decreasing quantity of free acid and the constantly increas- ing quantity of Cu in the electrolyte does not get outside the above limits. More copper is dissolved from the anode than corresponds to the current density, because of the action of the atmospheric oxygen : Cu + O + H 2 SO 4 = CuSO 4 + H 2 O. The above reaction takes place under normal conditions only at the surface of the anode. If, however, the proc- ess is carried on at a low current density, a part of the copper as it goes into solution from the anode may become incompletely charged, assuming a univalent instead of a bivalent charge; in other words, the solution then con- tains cuprous ions, but these are precipitated again at the anode : It follows, therefore, that the insoluble particles remaining at the surface of the anode, which fall to the bottom as slime, are likely to contain considerable copper. The latter, however, will be easily dissolved if the electrolyte is agitated with air so that the concentration at the electrodes is equalized. THE TEMPERATURE may be raised to 40 C. (104 F.); there is no advantage in raising the temperature above this point, if the quality of the cathode copper is taken into consideration. THE CURRENT DENSITY may vary, according to the quality of the anode copper, between 4 and 15 amperes per sq.ft. THE POTENTIAL should be 0.1-0.3 volt. THE REACTIONS DURING ELECTROLYSIS are as follows: The anode copper carrying a positive charge goes easily into solution and passes to the cathode, where it loses its 106 METALLURGY charge and is deposited. Of the impurities present in anode copper, As, Sb, Bi and Sn go partly into solu- tion, but are again partly precipitated; Ag, Au, Pt, and Cu 2 S are insoluble; Fe, Zn, Ni and Co go into solution, but are not deposited on the cathode. The insoluble substance falls in a finely-divided form to the bottom near the anode, while the soluble substances that are not deposited remain and gradually contaminate the electrolyte, so that it must be drawn off and its valuable constituents recovered. PRODUCTS OF ELECTROLYTIC REFINING: Cathode copper (electrolytic copper) , 97-99% of the anode copper; Precious metals, ] Bismuth, tellurium, etc., [ from the anode mud. Blue vitriol. Impure cement copper ] / . IN r n 1 from the anode mud and the (arsenical) , finally also \ . ...... impure electrolyte. nickel vitriol. Electrolytic Treatment of Copper Matte (Process of Borchers, Franke and Giinther) has been carried out since 1907 in an experimental plant at the Mansfeld works. In previous unsuccessful attempts low-grade matte was used, but in this case matte carrying 72% Cu and upward is employed. In its essential features, the method of working is the same as that in the electrolytic refining of metallic copper that carries precious metals. The principal differences in pro- cedure and the products are as follows: ANODES, cast plates of concentrated matte (Cu2S). POTENTIAL: higher than in treating metallic copper, viz., 0.75 volt. REACTIONS DURING ELECTROLYSIS: with the proper cur- rent density, the copper is dissolved as when the anode is crude copper, and the sulphur from the Cu 2 S remains behind in the free state. COPPER 107 The other constituents of the matte act the same as in metal refining. If the current density is too low, only half of the copper is dissolved, the balance remaining as blue CuS. PRODUCTS OF THE PROCESS: Besides the products obtained from treating metal, there is in addition, Sulphur, which is recovered as yellow crystals on drying the anode mud. Although this process requires more power it possess the following advantages : ELIMINATION OF SMELTING the concentrated matte to FIG. 128. FIG. 129. black copper, and elimination of the damage due to the fumes from such work; LESSENING OF LOSSES in precious metals, which are greater in smelting for black copper than in any other stage of copper treatment. RECOVERY OF THE SULPHUR contained in the matte. Properties of Refined Copper : Specific gravity: 8.94. COLOR: yellowish-red, brilliant luster. x MECHANICAL PROPERTIES: of medium hardness and tensile strength, very ductile. 108 METALLURGY STRUCTURE of cast and electrolytic copper, granular; of rolled and hammered copper, fibrous (see Figs. 128-129), MELTING-POINT: 1084 C. (1983 F.). BOILING-POINT: is said to be in the vicinity of 2100 C. (3815 F.). THERMAL AND ELECTRICAL CONDUCTIVITY: nearly as great as Ag (0.96). ALLOYS easily with Mg, Al, Mn, Zn, Cd, Co, Ni, Hg and the precious metals, slightly with Fe, Mo, W, Cr if the metal is pure, more readily in the presence of Si. Copper also alloys easily with its own compounds, especially with Cu 2 O and Cu 2 S, and, when in the molten state, it alloys somewhat with H, CO and SO 2 . CHEMICAL BEHAVIOR. In the solid state it is fairly resistant to dry oxygen at the lower temperatures, but above 400 C it is easily oxidized. Crusts of CuO and Cu 2 O (copper hammer scale) form on the surface. In a damp atmosphere it easily forms basic salts with O and weak acids (verdigris) . In the molten state copper has a great affinity for S, the the highest of any metal except manganese. Of the acids, only the oxidizing acids, HNO 3 , hot concentrated H 2 SO4, and aqua regia, serve as solvents. With access of air, dilute and weak acids attack copper. Copper does not evolve hydrogen from acids, as its solution tension toward hydrogen is 0.329 volt. BISMUTH Sources Natural Sources: FREE, with small quantities of S and As. MINERALIZED as OXIDE, Bi 2 Os in bismuth ochre, and as SULPHIDE, Bi 2 S 3 , in bismuth glance. The sulphide occurs free, in solution or in combination, with sulphides or arsenides of cobalt, nickel, copper, lead and silver. Other Sources : LITHARGE and other products from the cupelling process. REFINERY SLAG from silver refining. Residues from the preparation of pure bismuth salts. These may be pure products for which there is no demand or salts that have become contaminated with impurities. (A) Concentration Processes Mechanical Concentration consists usually in hand picking. Chemical Concentration is used in working up litharge that contains bismuth. In the section on Silver, it was stated that at the beginning of the cupellation process, when bis- muth is present in the bullion, only the lead is oxidized; later, as the bismuth becomes more concentrated in the metal bath, it is also oxidized, forming Bi 2 Os, which enters the litharge. The latter seldom reaches a concentration of more than i to 2% Bi. It is prepared for the regular Pb-Bi sep- aration by: i. Reducing Fusion to form a Bi-Pb alloy, which is low in Bi at the start. 100 110 METALLURGY 2. Oxidizing Fusion, whereby litharge is formed which is at first low in bismuth but later becomes richer. The different lots of litharge are kept separate and operations i and 2 are repeated with the litharge low in Bi. The rich litharge (more than 20% Bi) is treated by leaching and precipita- tion, usually by 3. Agitation with Dilute HC1 (15%) accompanied by blowing steam through the solution; this results in the changing of the oxide to chloride. APPARATUS: earthenware jars with wooden covers and spigots. PRODUCTS: an acid Bids solution containing a small amount of PbCl 2 and, in suspension, some undissolved PbCl 2 which is removed by 4. Filtration. 5. Stirring the BiCl 3 solution into water and neutralization of most of the free acid with Ca(OH) 2 CK Na 2 CO 3 . This causes the hydrolysis of the Bids: BiCl 3 + H 2 = BiOCl + 2 HC1. The BiOCl is precipitated, while the PbCl 2 , not previously removed by filtration, remains dissolved in the weakly acid solution. The BiOCl is filtered, dried, and is reduced to Bi or is used in fire refining bismuth that carries lead. (B) Extraction of Bismuth Liquation : Ores carrying native bismuth were formerly subjected to liquation, in crucibles or retorts, to remove the greater part of the bismuth without melting the gangue. The process can be used only on rich ores and the gangue usually retains 4-5% Bi. The gangue is re-treated by one of the following processes : Reduction Process: applicable to oxidized ores, intermediate products and residues. Among the two latter are bismuth salts, e.g., the oxychloride. Because of the volatility of Bi 2 O 3 BISMUTH HI and of Bi itself, and because of the ready reducibility of Bi 2 O3, the temperature should be kept low during this process. It is necessary, therefore, to produce a slag of the lowest possible melting point in order that it may be as free from bismuth as possible. The silica content of such a slag should lie between a singulo and bi-silicate, in fact as near to the latter as possible. The bi-silicate slags used in other smelting processes (Cu, Pb), which contain chiefly FeO and CaO as bases, are too infu- sible for this process. In this case, therefore, Na 2 COs is used as a flux, besides materials carrying FeO and CaO. The high price of Bi permits the use of this somewhat expensive flux, especially as the fluidity of the slag is such that the Bi 2 O 3 and the reducing carbon quickly sink through it, thus reducing to a minimum the loss by dusting and volatilization. If basic bismuth salts, such as BiOCl, are to be treated, they should not be added directly to the melted charge because of their volatility. They should rather be stirred into wet Ca(OH) 2 or Na 2 COs, dried and calcined for a time in order that the Cl may be completely taken up by the Ca or Na. APPARATUS: the process may be carried out either in crucibles or in reverberatory furnaces. Melting in Crucibles does not require highly refractory materials, and hence well-baked clay crucibles may be used, although it should be borne in mind that the slag produced in the process will dissolve basic as well as acid oxides. For mak- ing these crucibles, which may be done as part of the process, it is important to use as a foundation a clay which has been well baked and subsequently ground. Such material is mixed thoroughly with the least possible amount of fresh clay for a binder. The crucibles are usually formed by hand. For heating, a simple air furnace may be used, but for large plants small continuous kilns are preferred, since they allow a better control of the temperature and are not so hard on the crucibles. The latter is an important point because the use of a graphite crucible, which more easily resists temper- ature changes, is prohibited. The corrosive action of the 112 METALLURGY FIG. 133. Borchers* Continuous Kiln. BISMUTH 113 slag permits the use of a crucible only once, and the expense of graphite crucibles would greatly increase the cost of the process. The common clay crucibles fulfil all requirements if they are carefully handled. This is difficult in air fur- naces, in which the fuel conies in direct contact with the crucible, where a cold piece of coke touching the hot wall may cause it to crack. A small continuous kiln designed by Borchers, with six chambers heated with producer gas, which is generated in a producer attached to the kiln, com- pletely overcomes the difficulties encountered in the air furnace and also results in a saving of fuel. In the continuous kiln shown in Figs. 130 to 133, the hot gases are conducted from the producer through the main flue which runs along the center of the upper part of the fur- nace. From this flue the gases can be conducted by means of fl-pipes of sheet iron, fitting into corresponding faucet pipes, to a branch flue for each chamber. Sliding doors have not been used here, as they cannot be fitted tightly enough to prevent the gas from entering chambers which are cooling. From each of the branch canals a number of small slits enter the corresponding chamber; between the slits are openings from an air flue lying beneath. The com- bustion begins in the chambers above the crucibles and the hot gases surround the crucible from top to bottom. The crucibles themselves rest upon supports, between which are openings that allow the gases to pass into the space below and thence through connecting flues to the next chamber, where they warm the filled crucibles. The gases usually pass through a second chamber before leaving the furnace. The air is not admitted directly to the combustion chamber but is first passed through two chambers containing crucibles filled with hot finished product. In cooling these crucibles the air becomes pre-heated. From Figs. 130 to 133 it is seen that six chambers are so arranged that an uninterrupted continuous process can be carried on. As soon as the con- tents of one chamber are melted, gas and air are admitted 114 METALLURGY f - into the next chamber, which has already been strongly pre- heated, and the products of combustion are conducted through an intermediate, freshly-charged chamber. The chamber in which the air first entered is now closed, the crucibles are removed, and a new charge is put in place. Melting in Reverberatory Furnaces. In constructing the rever- beratory, it should be borne in mind that bismuth in the FIG. 135. H FIG. 136. Borchers' Reverberatory Furnace. FIG. 137. molten state easily leaks out through the finest cracks or imper- fections in the walls. Furnaces of earlier construction, having heavy brickwork supports for the hearth, are being discarded because they absorb large quantities of bismuth, thus dimin- ishing the output until the furnace is torn down. Even then the recovery of the metal from the hearth material is tedious and costly. In plants designed and constructed by Borchers, fur- naces have been built with movable iron hearths lined with half a course of bricks (Figs. 134 to 137). The advantages BISMUTH 115 of these hearths are a better and quicker extraction of the bismuth and easier access to the hearths for repairs to the lining, without tearing down the side walls. The hearths can be easily changed. Moreover, firebox and flue should be isolated from the hearth walls by an air space or a water- cooled plate, as otherwise the bismuth would flow into the firebox or flue through the two ends that are heated by the fire or flue bridge. At the beginning of the process in the reverberatory furnace, some slag is first melted and then the fresh charge is ..intro- duced. This should if possible sink at once t<\$\ Bottom, to prevent loss by dusting and volatilization. Precipitation of Bismuth is based upon the decomposition of Bi 2 S3 by means of Fe. As the equation shows, this process is applicable to sulphide ores. Apparatus and operations are the same as in the reduction process. If the ores contain substances which may be con- verted into matte (Cu, Co and Ni) or into speiss (As, Sb) the slag should not be as acid as in the reduction process. The silica content may lie nearer that of a singuto-silicate. (C) Bismuth Refining Crude bismuth usually contains As and Sb, often Pb, and occasionally precious metals. Of these impurities, As and Pb, also Sb if it is present, in small quantities, may be removed by an Oxidizing Fusion. The choice of an oxidizing agent depends upon the nature and amount of the impurities. If much lead is present this must first be removed, as otherwise lead compounds (plumbates) are formed which favor the oxidation of large quantities of bismuth. The lead is best removed by MELTING WITH BiOCl in iron kettles under a neutral cover of NaCl and KC1, or by 116 METALLURGY MELTING WITH NaOH with the addition of NaNO 3 , so that As can be oxidized to Na 3 AsC>4 in the same appa- ratus. In the above operation the NaCl and KC1 or the NaOH is first melted, then the metal added, followed by the oxidizing agent, BiOCl or NaNO 3 , which is stirred in with an iron spatula. When the refining is completed, an iron hook is low- ered in the molten metal and the kettle is cooled. After the mass has solidified, the slag is dissolved in hot water, th^V^ttle is warmed to free the metal from the walls, and v Ais is then lifted out by means of the imbedded hook. Melting with Sulphur and Na 2 CO 3 or K 2 CO 3 is used in excep- tional cases when the metal is high in Sb. The apparatus and procedure are the same as above. Electrolysis is used when the bismuth carries precious metals. VESSELS : stone jars. ANODES: plates or blocks of Bi carrying precious metals. CATHODES: pure Bi. ELECTROLYTE: HNO 3 and Bi(NO 3 ) 3 , or HC1 and BiCl 3 , in aqueous solution. CURRENT DENSITY: 15-30 amperes per sq. ft. POTENTIAL: 0.5-1 volt. Properties of Refined Bismuth : SPECIFIC GRAVITY: 9.74-9.8. COLOR: bright gray, slightly reddish luster. MECHANICAL PROPERTIES: very brittle, easily pulverized. STRUCTURE: large- faced, isometric, crystalline grains having a characteristic dendritic structure (Fig. 138). MELTING POINT: 268 C. (514 F.). VAPORIZES at red heat. ELECTRICAL CONDUCTIVITY: .013 compared with Ag. ALLOYS with most metals. The alloys with Pb, Sn, Zn and Cd have very low melting points, those with Cu and Ni are very hard. BISMUTH 117 CHEMICAL BEHAVIOR: Bismuth is very resistant toward oxy- gen at ordinary temperatures, but is easily attacked in the fused state, though not so readily as lead. It is precipitated from its dissolved or fused salts by lead. It is dissolved by FIG. 138 (x8). by HC1 and H2SO4 only in the presence of oxidizing agents. Hot concentrated H 2 SO 4 dissolves Bi but the acid is partly reduced to SC>2. In all compounds of technical importance Bi is present as trivalent cation. LEAD Sources Natural Sources : GALENA (PbS, 86.57% p k) : the associated minerals and gangue are: metallic sulphides such as sphalerite, pyrite, argentite and the sulpho-salts of arsenic and antimony; car- bonates, such as cerrusite, smithsonite, limestone and dolo- mite; further, hematite and also sandstone. The silver content of galena varies between o.oi and i%; in the Rhenish provinces o.oi to .015%; in the Hartz 0.05-0.1%. CERRUSITE (PbCO 3 , 77.52% Pb) occurs as an alteration product of galena. It is found in the upper parts of galena deposits and is associated with the same gangue. Other Sources: LITHARGE: PbO from cupellation. HEARTH MATERIAL (clay, marl, cement and less often bone ash) saturated with litharge from cupellation. DROSS, the first oxidation product from lead refining or cupel- lation, a mixture of lead oxide and lead antimoniate. LEAD MATTE, an intermediate product from smelting lead ores after the roast-reduction and precipitation processes; contains lead sulphide, iron sulphide, copper sul- phide, etc. Slags which contain an appreciable amount of lead are also returned to the smelter. ALLOYS of lead with zinc, copper, bismuth, gold and silver. (A) Concentration Processes Mechanical, Especially Wet Concentration, is used for ores carrying galena. Because of the high specific gravity of galena, it is easy to concentrate such ores up to a lead content of 70-80%, 118 LEAD 119 (B) Extraction of Lead Roast-reaction Process. By this method of working, sulphide lead ores are said to be treated in such a way that a part of the sulphide is converted by oxidation into oxide and sul- phate with which the unchanged sulphide reacts to form SO 2 and Pb. The chemical reactions taking place in the roasting and the reaction-smelting are explained by the following chem- ical equations: Roasting : PbS + 30 = PbO + SO 2 PbO + SO 2 + O = PbSO 4 . If lead carbonate is also present in the ores, it is decom- posed as follows : = PbOfC0 2 . Reaction Smelting: 2 PbO + PbS - 3?b + SO 2 PbSO 4 + PbS = Pb 2 + 2SO 2 . If the unchanged sulphide and the oxidation products are not present in the above proportions, an excess of oxide or sulphide, as formed by the first equation, will remain unchanged. The excess of sulphate, however, works as follows : 3 PbSO 4 + PbS = 4PbO -f 4 SO 2 . If carbonate is still present during the reaction-smelting it will be changed as follows: Schenck and Rassbach have determined experimentally 120 METALLURGY the conditions of equilibrium between Pb, S and O. Taking into consideration the SO 2 pressures, the conditions given i ii in m Fig. 139 were obtained for the formation of PbSO 4 , PbO and Pb. Field I is the formation zone for PbSO 4 ; here the princi- pal reaction is Pb 2 + 2SO 2 = PbSO 4 + PbS; even PbO is converted into PbSO 4 in this zone: 300 100 80 28 20 700 600 500 600 700 800 FIG. 139. 900 C. Field II is the formation zone for PbO, although the forma- tion of PbSO 4 is not yet impossible. Between PbO and PbSO 4 , solutions and chemical compounds are formed. The main reactions here are: and then the following equation, PbS + 3PbSO 4 = The exact boundaries of the zones in field II, in which on the one hand PbO and PbSO 4 form, and on the other PbO alone, have not yet been determined. Field III is the zone in which Pb alone forms from PbSO 4 , PbS and PbO: LEAD 121 The roast-reaction process is applicable only to ores rich in lead with not over 4% SiO 2 . Various modifications in the methods used have sprung up because of local conditions, especially with regard to wages and cost of fuel. These processes are: CARINTHIAN PROCESS. This is carried out in small rever- beratories and therefore with small charges. A low tem- perature is maintained and the roast and reaction processes are separate. The advantages of this method are: low lead loss by volatilization; little residue; pure lead. Its disadvantages: high fuel consumption and high cost for labor. ENGLISH PROCESS. Here large reverberatories, large charges, and high temperatures are used from the start. The fuel consumption and cost of labor are low compared with the Carinthian method, but there is a high loss by volatilization, requiring expensive condensation devices. There also remains a large amount of lead-rich residue. TARNOWITZ PROCESS. This is the Carinthian method car- ried out in large furnaces. There is small loss by volatili- zation, low fuel consumption and moderate cost for labor. It yields a pure lead, but leaves a large residue in rich lead which must be treated in special furnaces. FRENCH OR BRITTANY PROCESS. The work is done in large furnaces, large charges and a long roasting period, whereby considerable lead oxide and lead sulphate are formed, which necessitates the addition of fine coal during the reaction smelting. This process has no advantages over those mentioned above, but on the contrary a high loss of lead is entailed and the life of the furnace is short. HEARTH PROCESS. The ore, with fluxes and fuel (charcoal), floats upon the lead in a crucible-like hearth, three walls of which are made high. Through the back wall, air is blown into the charge above the molten lead. The plant is simple and quickly constructed, but the 122 METALLURGY process entails a high loss of lead by volatilization, and the laborers are endangered by lead poisoning. Roast-reduction Process. By this method the ore, after a long-continued roast, is melted with reducing agents to produce lead. Roasting. Qualitatively, the reactions are the same as during the roasting period of the roast-reaction process. Quantita- tively, however, the products are quite different, as it is not intended here to leave sulphur compounds in the roasted prod- uct for the purpose of reducing the oxides that are produced in the roast. The sulphides are, therefore, roasted down to the smallest amount necessary for forming matte (i.e., if the ores contain copper). In the absence of metals which are to be concentrated into matte, sulphates also are not desired in the roasted ore. The sulphates resulting from the roast are reduced during the reducing smelt. This re-formation of sulphides is at least superfluous; in the older processes the sulphate was decomposed by silica at the close of the roasting: 2PbSO 4 + SiO 2 = Pb 2 SiO 4 + 2SO 2 + O 2 , and the temperature was raised correspondingly. Apparatus: reverberatory furnaces with hearths narrowing near the fire bridge. The newer processes were suggested by the method of Huntington and Heberlein, in which a mixture of galena and lime is first roasted at about 700 C. and the roast product after being cooled to about 500 is blown to PbO in a converter. Savelsberg dispensed with the preliminary roasting by moistening well the mixture of ore and lime and blowing it directly in a converter. Finally Carmichael and Bradford treated a mixture of ore and gypsum directly in a converter. In all three cases, the attempt is made to fulfil the con- ditions explained by Schenck and Rassbach for obtaining oxides, either directly or with the intermediate formation of LEAD 123 lead sulphate, and to bring the ore to a proper condition for treatment in the subsequent reducing smelt. Plumb- ates form at the beginning as well as at the closing stages of these treatments. The discovery of Doeltz that CaSO 4 does not react with PbS does not affect the Carmichael-Bradford process. In the plant operated in Australia the reacting substances are not CaSO 4 + PbS, but CaSO 4 + SiO 2 + PbS. Reduction Smelting. In the roast-reaction process, the unde- composed PbS serves as a reducing agent for the oxides and sulphates formed in the roast. In this process, however, the reduction is effected by carbon and carbon monoxide: 2 PbO + C = Pb 2 + CO 2 . PbO + CO = Pb + CO 2 . The direct reduction of lead silicate in the original ore, or of that produced in roasting, cannot be effected by carbon; the lead oxide, therefore, in order to be acted upon by the reducing agents must be set free from the silicate by fluxes which also form an easily -fusible slag: Pb 2 SiO 4 + CaO + FeO = CaFeSiO 4 + 2PbO. The roast reduction process is applicable to nearly all lead ores; if the silica is low, a silicious flux may be required. In contrast to the roast-reaction process, in which only one apparatus is used for both stages of the process, the roast reduction process requires two forms of apparatus. Usually the ROASTING is carried on in reverberatory furnaces or in funnel-shaped or bowl-shaped converters having tuyeres at the bottom. THE REDUCING SMELT is always carried out in shaft fur- naces. The Precipitation Process. In this process the roasting of the sulphide ores as such is avoided, because the roasting is always difficult to complete on account of the fusibility of galena, lead oxide, and the other lead compounds that result from the 124 METALLURGY roast. The sulphide ores, therefore, are fused directly with fluxes containing metallic iron : Pb + FeS. The iron sulphide resulting from this reaction exerts a strong solvent action on the other sulphides. If this fact were not taken into consideration, the first iron sulphide produced in the reaction would unite with the galena and thus prevent its decomposition by iron. Hence an excess of galena is used over the quantity required by the above equation: #PbS + Fe = FeS(PbS).r_i+Pb This production of lead matte has several advantages: THE MATTE may be roasted in low shaft furnaces, producing concentrated gases which may be utilized for making sulphuric acid. This advantage is not possessed by the roasting apparatus used in the roast-reaction and roast- reduction processes. With the roasted lead matte, a large part of the iron used in the original fusion is returned to the process, as this product consists chiefly of lead oxide and iron oxide: 2 (PbS.FeS) + 130 = 2PbO + Fe 2 O 3 + 4SO 2 . The precipitation process, because of the above facts, may be used to advantage on ores which do not contain large quantities of other sulphides. The constitution of lead matte has been cleared up by the investigations of H. Weidtmann, who found that no chemical compounds exist between PbS and FeS, but a eutectic exists having the approximate composition PbS + FeS and a melting point of 780 C. (1435 F.). Apparatus for the Roast-reduction and Precipitation Proc- esses : Roasting Apparatus: HEAP ROASTING is used as a preliminary operation for ores high in zinc and pyrite, the object being to sulphatize and leach out the zinc. LEAD 125 KILNS for roasting matte (see Copper;. REVERBERATORY FURNACES. In the early methods , the long-hearth, hand reverberatory was used almost univer- sally. The hearths were often up to 85 ft. in length and from 8 to 1 6 ft. in width. If floor space was lacking, furnaces were constructed having two hearths, one above the other, each from 40 to 50 ft. long. At the sides of the hearths were working doors, 5 to 6J ft. apart. The reason for build- ing such long hearths was because of the nature of the raw material and of the roasted products. They are readily fusible and require, therefore, a low temperature and a correspondingly long time for the roasting. In spite of the endeavor to pass the ore quickly through the furnace, a considerable quantity remains continually on the hearth, and is moved from time to time in small parcels toward the firebox, where the final roasting takes place. The construction of the hearth depends upon the special object of the roast. SLAG ROASTING. By this method the reaction between PbSO 4 and SiC>2, mentioned above, is carried out as com- pletely as possible. The portion of hearth lying nearest the firebox is constructed in such a way and of such dimen- sions that the roasted product is fused here. The slagging hearth, therefore, is made narrower and lower at this point in order to concentrate the heat and collect the fused material. At the side of the fusion hearth farthest from the firebox, the hearth widens and usually becomes higher in order that the temperature may be greatly lowered for the roasting (Figs. 140-141). SINTER-ROASTING. In this method a complete desul- phurization is not sought; the chief aim is to heat the pul- verulent roasted product only enough to sinter it. The furnaces required for this work have a short sinter-hearth near the fire-bridge, to which is joined the roasting-hearth proper a step higher, making, as it were, a terrace. The sinter-hearth does not need to be narrowed. 126 METALLURGY NON-SINTER ROASTING. In this work the product of roasting remains in a pulverulent state. The hearth of the furnace used for this method is a slightly inclined plane, sloping away from the flue-bridge toward the fire-bridge. FOR REVERBERATORY FURNACES WITH STATIONARY HEARTHS many mechanical stirring devices have been used. These have been described under Copper, pp. 63 to 71. REVERBERATORY FURNACES WITH MOVABLE HEARTHS are used in the Huntington-Heberlein process, espe- FIGS. 140-141. Reverberatory Furnace with Slagging Hearth. cially those with revolving circular hearths, having stationary rabbles, similar to the heating furnaces used in the manufacture of coal briquettes. CONVERTERS are used in the following processes: HUNTINGTON-HEBERLEIN: conical, cast-iron vessels 5-8 feet wide at the top and 5-6^ ft. deep, fitted at the bottom with grates to support the charge and to distribute the blast, which is also admitted at the bottom. Capacity 5-8 tons. CARMICHAEL-BRADFORD: conical, cast iron vessels; upper diameter 6 ft., lower diameter 4 ft., depth 5 ft. LEAD Above the bottom, which is arched downward, is placed a perforated shell arched upward which acts as a grate. The blast is admitted into the intermediate space thus formed. Capacity, 4 tons (Fig. 142). SAVELSBERG. Nearly hemispherical, cast-iron pots about 6 1 feet in diameter. Air admitted from below. Capacity 8 tons. Blast required, 250 cu.ft. per min. Blast pressure at start, 4 to 8 in. of water, later 20 to 24 in. Time of blowing, 18 hours (Fig. 143). Apparatus for the Reduction and Smelting : SHAFT FURNACES. Of the older constructions the low FIG. 142. Carmichael-Bradford Converter. Scale, i : 50. FIG. 143. Savelsberg Converter. Scale, i : 50. furnaces (Krummofen) have almost wholly disappeared, while the medium furnaces (Halbhochof en) are found only in a few plants. The high furnaces (Hochofen) are used almost exclusively in modern plants. They are, of course, much smaller than the blast furnaces used in the iron industry. The " Krummofen " are low shaft furnaces not more than 6J ft. in height, with square, rectangular or trapezoidal cross-section. They usually have only one tuyere which is placed in the back wall. Various forms of these furnaces 128 METALLURGY are described in the oldest metallurgical literature ("Agric- ola de rebus nietatticis," 1657, pp. 313-319; Schluter, " Huttenwerke," 1738, tables XX-XXX). " HALBHOCHOFEN." These are low shaft furnaces, 7-14 ft. in height from the tuyere level to the throat. They are usually of trapezoidal cross-section with from 2 to 8 tuyeres on one side of the trapezoid. Of these furnaces the Stolberg furnace has been most widely adopted and was FIG. 144. Pilz-Freiberg. FIG. 145. Upper Harz. FIG. 146. Allis-Chalmers. also used until recently in Freiburg. They are still in operation at the lead plants in Stolberg and Mechernich. " HOCHOFEN." Although the name is usually applied to blast furnaces having at least 13 ft. between the tuyere level and throat, shorter furnaces of this type are found in some lead plants. Modern blast furnaces are built with circular and rect- angular horizontal cross-sections (Figs. 144-147). The shaft is usually made of brick and the bosh and smelting zone are water-jacketed. The throat is open, except during blowing in and blowing out, when it is covered with a LEAD 129 hood. Downcomers start from the centre of the throat or from the side under the feed floor. They are usually FlG. 147. North American Water-jacket Furnace. Scale, i : 150. made of encased sheet-iron pipes. Fore-hearth: crucible furnace or Arent's siphon tap. The direct electric smelting of lead ores has not yet been successful. (C) Lead Refining The crude lead obtained in the preceding processes may con- tain various quantities of all the metals found in the ores, also metallic compounds such as sulphides, arsenides, antimonides, etc. Among the impurities are: Au, Ag, Cu, Sn, Sb, As, Bi, Co, 130 METALLURGY Ni, Fe, Zn, and S. They may be separated from the lead by the following operations : Liquation and Crystallization Processes (see Silver). In this way Au, Ag, and Cu are removed by being concentrated in a part of the lead. Oxidation of the Impurities (true lead refining). (a) Oxidation of S, As, Sb, Bi by atmospheric oxygen. (b) Oxidation of Zn, Fe, Co, and Ni by steam. Both of the above methods have already been described under Silver. Oxidation of the Lead, the so-called cupellation process for recovering precious metals (see Silver and Bismuth). Electrolysis. After several unsuccessful attempts by Keith, Tommasi and others, this process was successfully applied by Betts, in 1907, to the refining of antimonial lead and base bullion. A description and sketches of a complete plant are given in Metallurgie, 1908, 5, 68. ANODES: base bullion. CATHODES: pure lead. ELECTROLYTE : aqueous solution of silicon fluorides containing 22.7 oz. PbSiF 6 9.4 oz. H 2 SiF 6 6.7 oz. gelatin per 100 gallons. TEMPERATURE: between 17 and 57 C. (63 and 135 F.) CURRENT DENSITY: 10 to 12 amp. per square foot of cathode. POTENTIAL: 0.15-0.36 volt. Properties of Lead: SPECIFIC GRAVITY: 11.4. COLOR: bluish gray, shiny. MECHANICAL PROPERTIES: low tensile strength, high duc- tility. STRUCTURE. Ingot fracture (Fig. 148). Crystallized shell from which the molten metal was removed before freezing was completed (Fig. 149). Granular structure obtained by etching the bottom surface of a lead ingot (Fig. 150). LEAD 131 MELTING POINT: 327 C. (621 F.) VAPORIZATION takes place at red heat, though the boiling point is between 1200 and 1300 C. (2200 and 2375 F.) FIG. 148. FIG. 149. THERMAL AND ELECTRICAL CONDUCTIVITY: low, the latter 0.0756 that of Ag. ALLOYS with most metals. Its most important metallurgical alloys are those with the precious metals, for which it is 132 METALLURGY utilized as a solvent: with zinc (limited solubility), with antimony (hard lead) and with tin (solder). CHEMICAL BEHAVIOR: readily oxidized on the surface by O, H2O, and CO2 of the atmosphere. The coating of oxide and basic carbonate thus formed prevents further corrosion. In the same way the action of HC1 and H 2 SO 4 is restricted even in the presence of oxygen. Without the presence FIG. 150 (x 33)- of oxygen or oxidizing agents in the acid, it is hard to dissolve lead in acids, because the electrolytic potential is only +.151 against H. There are two common oxides, PbO and PbO 2 , the former basic, the latter acid. These unite to form a plumbate, Pb 2 PbO 4 (which, together with free PbO, constitutes red lead). Of free sulphides, only PbS is known; it dissolves readily in other sulphides (matte), but forms few chemical compounds. TIN Sources Natural Sources : OXIDIZED ORES. Tin-stone, or cassiterite, SnO 2 . In primary deposits, lode tin; in secondary deposits, stream tin. Gangue: acid eruptive rocks, particularly granite, also quartz, CaF 2 , FeWO 4 , CaWO 4 , sulphides such as Cu 2 S, PbS, Bi 2 S 3 , MoS 3 , etc., and arsenides such as FeAsS. SULPHIDE ORES. Tin pyrites, FeCu 2 .SnS4, which is much rarer than tin- stone. Other Sources : METALLIC INTERMEDIATE PRODUCTS AND BY-PRODUCTS: Hard Head and Liquation Dross (Fe-Sn alloy). Tin plate clippings (Fe covered with a coating of Sn). White Metal in the form of impure turnings: Sn, Pb, Cu, Zn, Sb. OXIDIZED BY-PRODUCTS: Slags: Silicates and Stannates. Ashes: Oxides mixed with metallic grains. (A) Concentration Mechanical Concentration. On account of the high specific gravity of tin-stone, viz., 6.8 to 7, the mechanical concentration of tin ore is carried out quite commonly, and sometimes by the simplest means. Chemical Concentration. Among the minerals accompanying tin-stone are found, as mentioned above, sulphides, arsenides and tungstates. The first two are almost always present, but of 133 134 METALLURGY the last mentioned only a few ores contain enough to cause trouble. The removal of S and As is effected by an oxidizing roast in a simple roasting furnace, usually fed by hand. For ores with much sulphide of Fe and Cu, it has been pro- posed to carry out the roasting in such a way that the sul- phates of these metals are formed, which can then be leached out, but such a process does not seem to have met with much practical application. The removal of tungstates is necessary with ores containing much tungsten, because tungstates favor the passage of tin into the slag. Even a small percentage of tungstate suffices not only to carry over considerable amounts of tin into the slag on smelting the ore, but also hinders the reduction of tin in the subsequent smelting of the slag. Hence, even in smelters that work without previous concentration, with ores containing small amounts of tungsten, it is worth while to take the slags in which considerable tin has accumulated, and remove the tungsten in order to recover the tin by another smelting. The removal of tungsten is effected by an oxidizing roast of the pulverized ore or slag, with sufficient Na2CC>3 to convert the W into Na2WO 4 . The latter is removed by lixiviation, whereupon the tin can be obtained without diffi- culty. Fuller details concerning such work will be given under Tungsten. (B) Extraction Inasmuch as tin-stone is the principal raw material, which, when necessary, is concentrated or purified by the above-men- tioned work, the chief operation of the tin smelter is of a com- paratively simple nature. It consists of a reducing fusion of the ore, and other oxidized waste products, and the subsequent working over of the slag. During the last two decades it has also been found possible by an electrolytic process to recover tin from the scrap obtained in the preparation of tin plate. TIN 135 The Smelting Operations for Oxidized Ores and Waste Products are the following. i. Reducing Fusion. Concerning the reactions that take place during this process, singular views are to be found in metallurgical literature. The ready reducibility of SnO 2 by KCN, which is preferred as a flux and reducing agent in making the blow-pipe test for tin, has led to the assumption that during smelting the carbon of the coal used for effecting the reduction is partly converted into cyanide, which then acts as in the blow-pipe test. There has never been any proof of this and, moreover, the assumption is very improbable, because, in most cases, the slag is not kept basic enough nor hot enough for the formation of cyanide. An experimental investigation by Mattonet, in the laboratory of the author, has shown that the formation of a singulo-silicate slag is most favorable for the extraction of the tin and for the slag- ging off of the other metals. According to the nature of the ore, the fuel, and the other working conditions, it has been customary to make use of either a blast furnace or a reverberatory furnace for smelting. The above-mentioned researches of Mattonet, however, have established the fact that an electric furnace can be of sendee for producing slags low in tin. In both blast-furnace and reverberatory smelting, the extraction of the tin is made difficult by the unfavorable position of the reduction temper- ature (1000 to 1100 C.) in respect to the melting point (232) and the boiling-point of the metal. The latter, to be sure, is said to be above 2100, but volatilization, of the metal takes place at temperatures far below the boiling point. The furnace charge must, therefore, be heated to 1100, and the combustion gases, which furnish the source of the heat and its transference, must be kept at least 200 hotter, i.e., at about 1300, so that the Sn is heated to about 1000 above its melting-point, at which temperature it shows a tendency to volatilize with the reaction-gases and also possesses a high solution tension which favors its passing into the slag. In 136 METALLURGY the electric process, particularly since the charge itself can take part in the heat transference, the surroundings are kept colder than the charge. THE BLAST-FURNACE PROCESS, which is the oldest method of working, is applicable if a very pure, lumpy ore, and a pure fuel, e.g., wood charcoal, are available. The smelting takes place sometimes without any additions of flux, and sometimes with slags obtained from tin ores, used in amounts varying from 25-50% of the weight of ore. In the Chinese and Malay tin districts, the work, in some cases, is still carried out with very primitive apparatus. Whereas the original blast-furnaces of the Sunda Islands consisted of pits, about 20 in. deep and 12 to 16 in. wide, dug in the earth and provided with hand-driven blast which passed through hollow tree trunks, the Chinese draft and blast furnaces are prepared by tamping clay within a barrel-shaped form made of bamboo rods and then carving out the shaft and the holes for blowing and tapping. These furnaces are built up to 6 ft. in height, with a shaft up to 16 in. wide and about 5 ft. 3 in. high. Such furnaces, when used for smelting the ore, will work with natural draft, provided the mate- rial is sufficiently coarse; and those used for smelting the slag work with a blast, which, as in the above instance, is produced by hand, using hollow tree trunks that are pro- vided with wooden pistons. (Fig. 151.) Even the larger blast furnaces, such as are used in India at works con- ducted by Europeans, as well as in Bohemia and in Alten- berg, are built so short and narrow that they can almost be classed as " dwarf furnaces." Concerning the con- struction of such furnaces, enough was given under Copper and Lead, so that it is unnecessary to go into further details here. The method of smelting tin ores which is in common use to-day is by means of the REVERBERATORY FURNACE, because the ore is usually fine grained in nature and of TIN 137 low grade, so that the use of the blast furnace is out of the question. Although formerly considerable importance was attached to the nature of the fuel, inasmuch as only those varieties of coal are suitable which produce long flames, to-day this difficulty can be overcome by the use of gas firing. Even in this case the furnace is almost always charged merely with ore and coal; when the gangue is too acid it is sometimes necessary to add limestone and small FIG. 151. Chinese Furnace for Smelting Tin Ore. amounts of fluorspar. In order to prevent mechanical losses of the powder, the mixture is usually moistened with water. After being introduced into the hot furnace, the charge is levelled off and heated strongly for from 5 to 8 hours with closed doors and nearly checked draft; after the mass has softened, it is strongly heated for from 45 minutes to an hour, the metal is tapped and the slag, which solidifies readily on cooling, drawn off. In other works, the slag is thickened by ' working in some 138 METALLURGY coal so that it can be removed through the slag door before tapping the metal, which is then covered with only little slag. According to the size of the reverberatory furnace, the amount of ore used in a single charge varies from 1600 Ibs. to 4 tons, the time of heating from 6 to 12 hours, the consumption of fuel from 60 to 120% of the ore, the labor employed from i to 3 men per furnace, and the smelting losses up to 7% Sn. As regards the construction of the reverberatory furnaces, the following points, which were partly discussed under Bismuth, should be considered: Sn penetrates readily into the narrowest fissures of masonry, and since it Is practically impossible to make an absolutely impenetrable hearth, this is not built any stronger at the bottom than is absolutely necessary for holding the heat within the smelting zone. In all the newer constructions of rever- beratory furnaces, the hearth is built upon rails, or other supports, independent of the furnace-masonry, so that it can be easily repaired without destroying the latter, and beneath the hearth is an air space which is well ventilated and in some cases even kept filled with water. The reason for keeping this space cold either by ventila- tion or by water, is to cause the immediate solidification of any tin that trickles through. On account of the large dimensions of the furnaces, it is not feasible to make movable hearths, as in the bismuth and antimony industries (cf. p. 114, Figs. 134-137). It is necessary, however, to have the fireplace and flue built in such a way that they are separated by air spaces, or some insulat- ing material, from the parts of the furnace containing the hearth. Otherwise the molten tin would tend to run toward the fire bridge or flue bridge and find a particularly favorable opportunity to penetrate through them. The smelting of tin in an electric furnace has been carefully studied by Mattonet, in the author's laboratory. Exper- iments performed in the attempt to smelt impure tin ores, TIN 139 without a preliminary roasting, in such a way that the Cu I b o I and a part of the Fe would be converted into a matte and the Pb into an acid slag, proved unsuccessful. Even 140 METALLURGY the sulphur in the ore passed to a considerable extent into the crude tin, which contained up to 2.7% S. The yield of Sn under these conditions was also unsatisfactory. Better results were obtained with roasted ores and an approximately neutral slag, in which, however, by adding some soda to the charge the melting-point was lowered somewhat and Na 2 O made a part of the basic constituents. Since at most only 450 Ibs. of Na^COs (cost $2.50) are needed for a ton of Sn, the cost of working is not increased, because more than enough Sn is prevented from going into the slag to compensate. The slag obtained was perfectly free from Sn globules (in the case of other methods of smelting it is permeated by pellets of Sn), and contains only from i to 1.5% of slagged Sn, as compared with from 2 to 5% by other processes. Direct heating by resistance is applicable for this method of smelt- ing on account of the relatively good conductivity of the charge and of the slag. In this way, the proper tempera- ture for the reduction of the Sn is maintained and is not much exceeded in the bath. Since, moreover, the blast, which in one method tends to prevent the flowing together.of the Sn globules, is absent and an acid lining of the hearth, which causes losses by slagging and by volatilization, is unnecessary, the direct yield of Sn is so good that, whereas it is absolutely necessary to recover Sn from the slag in smelting by the blast furnace or by the reverberatory furnace, in the case of electric smelting this subsequent treatment of the slag can be dispensed with altogether or limited at the most to one smelting. Consequently the smelting of rich and poor slags, as described below under 2 and 3, refers only to blast-furnace and reverberatory furnace practice. 2. Smelting of Rich Slags. The slags obtained by smelting the ore contain, as mentioned above, tin which is mechan- ically enclosed, as well as that which is slagged. As soon as a sufficient amount has accumulated, it is smelted, TIN 141 usually in a reverberatory furnace, together with other metallic waste (dross from the refining, and the "hard- head " or residue which remains on the hearth), and, in case not enough iron is present to cause the separation of the slagged tin, some scrap iron is added also. Obviously it is necessary to conduct the work with a reducing flame and with some fuel added to the charge. Since the smelting of rich slags requires a higher tem- perature than that of ores, this process is often carried out for the purpose of seasoning new hearths which are to be used for smelting ore. A considerable part of the slag adheres to the masonry of the hearth and either remains solid during subsequent smelting of ores, or else is so viscous that metal does not flow through it. 3. Smelting of Poor Slags. This is carried out in many works in blast furnaces, usually in small water-jacketed furnaces having a circular cross-section; large pieces of coal or coke are mixed with the charge. Although d fairly pure tin is obtained directly by the two former operations (the crude metal from rich slags even contains 95% Sn), in this case the crude metal is a fairly rich iron alloy (80% of Sn, 20% of Fe). Electrolytic Solution and Deposition. For twenty years this process has been carried out in working up tin-plate waste. The object, in most cases, is to recover the tin from old scrap made in the manufacture of tin-plate. According to the thickness of the sheet iron which was given a coating of Sn, the material contains from 3 to 5% of Sn. As a rule, how- ever, it is not safe to reckon on more than 2 or 3%. The recovery of the tin has been carried out most successfully in the factory of T. Goldschmidt at Essen. Most other works have made use of the same method and profited by the experi- ence gained there. The apparatus and conditions are as follows : ELECTROLYZING VESSELS: iron tanks. These are con- nected with the circuit and serve as cathodes. 142 METALLURGY ANODES: the tin-plate clippings packed not too firmly in a basket made of coarse iron wire (about 100 Ibs. of clip- pings are placed in a basket). CATHODES: sheet iron and the walls of the tank. ELECTROLYTE: this consists at first of a caustic soda solu- tion containing at the most 9% of NaOH, corresponding to 7% Na 2 O. During the electrolysis the caustic soda is partly converted into carbonate and partly into stannate. Of the original 7% of Na 2 O, 3 to 3.5% remains as the free hydroxide while i to 1.5% of Na 2 O is converted into stannate and 1.7 to 2.8% of Na 2 O unites with CO 2 . If this amount of carbonate is exceeded, the electrolyte FIG. 155. FIG. 156. must be worked over, by treating with Ca(OH) 2 , whereby CaCO 3 is precipitated and NaOH formed. TEMPERATURE: 70 C. (158 F.). POTENTIAL: 1.5 volts. Concerning the reactions that take place during the elec- trolysis there is but little reliable information. It is known that the Sn goes into solution as stannate, and therefore forms a complex anion, and that it is consequently indirectly precipitated (by Na ?) . PRODUCTS : ELECTROLYTIC TIN. Inasmuch as old tin-ware that is worked up in this way contains more or less solder, the Sn obtained by electrolysis usually contains some Pb. For this reason, the Sn is deposited in the form of TIN 143 crystalline grains and does not form coherent plates. On being dried it becomes covered with a layer of oxide, so that it must be subjected to a reducing smelt. LEAD CARBONATE: the first precipitation product in working up impure electrolytes. STANNIC ACID, or SnO 2 : the second precipitation prod- uct from impure electrolytes. (C) Tin Refining It is clear, from what has been said concerning the extraction of tin, that the principal impurity of the crude metal is iron, and that to a less extent metals such as lead, copper, tungsten, etc.> may be present. i. Liquation. By this process it is chiefly the Fe which is removed either in the free state, or in the form of a rich Fe alloy. The impure tin is carefully heated so that the melting-point of tin is exceeded as little as possible, or, in other words, so that the tin alone is liquefied, and the above-mentioned Sn-Fe alloy remains behind as a solid. For carrying out the process it is customary to make use of small, simple reverberatory furnaces, the sloping hearths of which are divided by steps into two or three sections. When there are three such sections the crude tin is placed in the middle one; the pure tin that liquates out is allowed to solidify and is then transferred into the third section, which is the farthest away from the fire-bridge. The tin that flows away from here is purified further if necessary. The residue remaining in the third compartment, the liquation dross or " hard-head," goes back into the second section, and all the residues from this in turn, including that from the crude tin, are placed in the first section, which lies nearest to the fire. The tin that flows away from the first compartment, after solidifying, is placed in the second, just as that from the second goes to the third. The dross from the first com- partment is used partly as an iron-bearing flux in smelting 144 METALLURGY tin slags, and partly for preparing Pb-Sn alloys, by simply melting it up with Pb. In the smelter certain standard alloys are prepared and sold as such, e.g., that with 50% Sn and 50% Pb) (common solder). 2. Oxidizing Fusion. Although the bulk of the Fe, W and a part of the Cu is removed by the liquation process, there still remain small amounts of these impurities in the tirj. They may be removed with the aid of atmospheric oxygen. To accelerate the oxidation process, there is placed in the iron kettle which contains the molten tin, a pole of green wood from which vapors of water and other decomposition prod- ucts of the wood are at once evolved. The melted metal is stirred up by the escape of these gases, so that new portions of it are constantly brought to the surface, where they come in contact with the oxygen of the air. The water vapor given off by the pole also tends to assist in the oxidation of the impurities. 3. Crystallization of Tin. When Sn is strongly contaminated with Pb it is almost impossible to effect a sharp and complete separation of the two metals. To be sure there has been no want of attempts in this direction. Since, however, considerable quantities of Pb-Sn alloys are used technically, it is not at all necessary in most cases to attempt such a separation it suffices to enrich the metal. This, as the experi- ments conducted by the author with the aid of Mattonet have shown, can be carried out in much the same way as in the separation of lead and silver, by the crystallizing out of the lead. From the freezing-point curve (Fig. 157) it is evident that the two metals are soluble in one another in all proportions and that there is a eutectic formed with 37% Pb and 63% Sn which melts at about 180 C. If, therefore, crude tin containing a small percentage of Pb is allowed to cool slowly, it is obvious that pure Sn will crystallize out, leaving behind a mother metal which constantly increases in Pb content, until finally the com- position of the latter is that of the eutectic alloy. In TIN 145 practical work, the cooling will not be carried out as far as this, but will be stopped while the alloy is still liquid and contains somewhat less Pb than the eutectic does. This is drawn off in the same way as in the desilverization of crude lead, leaving behind an alloy containing a little Pb. Naturally, the process does not yield perfectly pure tin the first time, simply because the Sn crystals are more or less contaminated with the mother metal, which contains Pb and adheres to some extent to the crystals. Thus, to pre- 340 320 300 280 260 240 220 180 10 20 30 I 40 50 60 70 80 90 100$ Pb 100 90 80 70 | 60 50 40 30 20 , 10 jf Sn 63$ Sn 37$ Pb FIG. 157. pare perfectly pure Sn it is necessary to repeat the proc- ess a number of times, and, as in the desilverization of lead bullion, there is no difficulty in the work. The Sn alloy obtained by this process of separation need only be brought to a content of 10% Pb, which is used very extensively as such for making terne-plate. Properties of Tin: SPECIFIC GRAVITY: 7.3. COLOR: white with a pale bluish tinge; yellow when hot. MECHANICAL PROPERTIES: low tenacity and hardness, high ductility; most ductile at about 100; at 200 146 METALLURGY so brittle that it can be pulverized. At low temperatures it becomes changed into gray tin powder. STRUCTURE: see Figs. 158, 159, 160. MELTING-POINT: 232 C. (450 F.). BOILING-POINT: 2100-2200 C. (3800-4000 F.) ; volatilizes perceptibly at 1200 C. (2200 F.). THERMAL AND ELECTRICAL CONDUCTIVITY: good, the latter about 0.13 that of Ag. ALLOYS with almost all metals. Many Sn alloys are valued FIG. 158. Surface of Impure Tin (X33). highly (with Cu = bronzes; with Cu, Pb, Sb = bearing metal, and solder; with Sb, Ni, W, etc., metal for fine- art castings). In experiments for obtaining alkali and alkaline-earth metals by the electrolysis of their fused salts, Sn has been found to be a particularly good solvent for these metals. CHEMICAL BEHAVIOR: very stable toward O and H 2 O at temperatures below its melting-point; at higher temperatures it oxidizes readily and also combines readily with S, P, As, and Sb. Cl and other halogens TIN 147 act upon it even at ordinary temperatures, forming SnCl 4 , etc. Of the mineral acids HC1 dissolves the metal best, H 2 SO 4 only slowly, and HNO 3 converts FIG. 159. Deeply Etched Under- surf ace of a Cast Tin Ingot. FIG. 160. Large Grains caused by Tempering Tin. it into insoluble metastannic acid, which is probably a polymer of H 2 SnO3. This dissolves in alkali to form salts soluble in water, called stannates. With O, S and on dissolving in acid, two series of compounds are formed, those of the stannous oxide type (SnO) 148 METALLURGY called the stannous compounds (Sn++) and those of the stannic oxide type (SnC^) called stannic compounds (Sn+ + + +). In the presence of strong bases Sn loses its basic properties and unites with O to form an anion (stannites and stannates). In acid aqueous solutions, as well as in fusions, Sn is precipitated by Pb whereas in basic solutions it precipitates Pb. ANTIMONY Sources Natural Sources : NATIVE: occurs as such rarely. As SULPHIDE: 80283, which is known as stibnite; this is the most important ore of antimony. As OXIDE: 80203, called senarmontite or valentinite. Other Sources : Waste products from the working up for other metals of ores containing antimony, particularly in the smelting of lead (dross). (A) Concentration Liquation. The ready fusibility of Sb 2 S 3 makes it possible to melt out this compound from its ores, leaving behind, to be sure, a residue rich in Sb, since a considerable amount of Sb-2S3 (up to 20%) entangled in the gangue remains behind in the liquating apparatus. In spite of this difficulty, such work is still carried out to a considerable extent in Hungarian, Jap- anese, and Chinese antimony districts, because the concentration product, the so-called antimonium crudum, is still much used in English antimony works for the preparation of pure antimony, and in German chemical factories for the manufacture of antimony preparations. If it were desired to-day to erect anti- mony works in the vicinity of the mines, naturally the liquation process would not be used at all, because the residues would have to be worked over considerably on account of the large amount of antimony they contain, and because there is not a large amount of gangue to slag off when the ores are treated directly. 149 150 METALLURGY APPARATUS: Crucibles or Tubes placed either in air furnaces of the simplest type, or upon the hearth, or in a special heating chamber of a reverberatory furnace. The cruci- bles are provided with holes in the bottom and the tubes are open at both ends so that the melted Sb 2 Sa flows out into collecting vessels. Crucibles heated in a reverber- atory furnace are placed over a tapping hole which leads from the bottom of the hearth to the outside. Reverber- atory Furnaces with deep hearths (see Fig. 47, p. 29). Oxidizing Roast with Sublimation. The ready volatility of Sb 2 C>3, and the demand on the part of chemical industries for this oxide, has led to an extensive practice of subjecting antimony ores to an oxidizing roast, in spite of the fact that sulphide ores can be worked into metal without any prelimi- nary roasting. Then again, ores can be worked with in this manner from which no antimony sulphide could be obtained by liquation and with which the expense of slagging off all the gangue would be too great on account of the large amount they contain. Now antimony on being roasted can be con- verted into Sb 2 O3, which has a basic character, or into Sb2C>5, which is acidic in nature. The latter is formed more readily in the presence of basic substances, but as Sb 2 O 3 is itself basic, considerable quantities of the higher oxide are likely to result unless special measures are taken to prevent its formation or to cause its reduction after it has once been formed. In all cases, the aim is to make Sb 2 C>3. Chemical industries demand this oxide in a condition as free as possible from Sb 2 O 5 , because the former is readily soluble in acid and the latter is not. But in extracting the antimony from the ores by the oxidizing roast, it is necessary to maintain the con- ditions for the formation of the lower oxide because this is readily volatile, whereas Sb 2 O 5 is not, and cannot, therefore, be removed as readily from the gangue by volatilization. If, on the other hand, the sublimation product is to be used for the manufacture of pure antimony, it is not necessary that it should be free from Sb 2 O 5 . ANTIMONY 151 APPARATUS: Although, in metallurgical literature, a number of forms of apparatus and processes are described which are very simple, yet in practice only low blast-furnaces or reverberatory furnaces are employed. In both cases, it is, of course, necessary to provide arrangements for properly catching and condensing the volatilized oxide. Lixiviation. It is well known that Sb 2 S 3 is soluble in aqueous solutions of alkali sulphides forming sulphantimonites, or sulphantimonates upon the addition of free sulphur. Sb 2 S 3 + aNa 2 S = 2Na 3 SbS 3 , Sb 2 S 3 + 3Na 2 S + S 2 = 2Na 3 SbS 4 . Chemical manufacturers for a long time have made use of these reactions in preparing artificial, precipitated antimony sulphide. Usually solutions of NasSbS 4 are first prepared and an intimate mixture of Sb 2 S 3 and S 2 , which is regarded as Sb 2 S5, is obtained on acidifying. If, however, the solution is to be used for the electrolytic deposition of antimony, the addition of S during the lixiviation is not only unnecessary, but undesirable. A solution of Na 3 SbS 3 can be electrolyzed directly, whereas one of Na 3 SbS 4 must first be treated with an excess of Na 2 S or of NaOH. APPARATUS: iron tanks, conical at the bottom with converg- ing sides, and provided with pipes for introducing steam, which serves to heat the solution and to stir up the ore in it. (B) Extraction Reduction Methods. For oxidized ores or metallurigcal products, the same methods are employed as in the case of the extraction of Bi (pp. in to 115) although in this case it is necessary to use cheaper fluxes because the price of Sb is to that of Bi as i : 20. 152 METALLURGY Precipitation Methods. This depends, as in the like-named processes for obtaining Pb and Bi, on causing the reaction to take place in the melted mass. As regards the apparatus and method of conducting the operation, what was said under Bi holds in the main here also. Crucibles and reverberatory furnaces are used. In reports concerning precipitation in rever- beratory furnaces a process is frequently described in metallurgical literature as the English Process, but which appears to be identical with the practice that has been followed everywhere that Sb 2 Sa has been melted with Fe. Under Bismuth it was pointed out that it is always necessary in melting charges containing easily volatile material, to provide a sufficient cover of fused material so that when a fresh charge is added it will at once sink below the surface. In smelting rich antimony ores, or the concentration product called antimonium crudum, it is clear that FeS will form the principal constituent of the slag. FeS, however, is known to be an excellent sol- vent for other sulphides and for metals, particularly Fe itself. Fe, moreover, is the agent that precipitates Sb from Sb2Sa in this process. Just as in an ordinary reducing fusion it is necessary to provide a reducing atmosphere, and solid reducing agent in the furnace and charge, so it is clear that care must be taken in the pre- cipitation process to have enough Fe present in the furnace before the antimony ore is introduced. Alternately charg- ing with Fe and Sb2Ss is prescribed, therefore, in order that the Fe may be given time to dissolve in the molten FeS, and thus attain a sufficiently fine state of division to act vigorously enough upon the Sb2S 3 with which it sub- sequently comes in contact. Electrolysis. The solution of Na 3 SbS3 obtained by extraction with alkali sulphide (cf. Concentration Work, above) can be electro- ANTIMONY 153 lyzed directly, as has been shown elsewhere, without any additions whatever to the bath, such as are recommended in the analytical estimation of Sb by electrolysis. Even although the processes that take place have not been wholly explained, it has been shown that Sb is present in the anion of the above salt as well as in Na 3 SbS 4 . The separation of the Sb on the cathode, therefore, must be the result of a secondary reaction; first the Na ions are discharged and then the Sb is precipitated from the sulphide. This partly accounts for the fact that it is difficult to deposit the Sb upon the cathode in a dense condition. The total result of the electrolysis may be expressed by the equation, 2NasSbS3 = Sb 2 (cathode) +3^282 (anode). In 1887 Borchers established the fact that large quantities of Na 2 S 2 are formed at the anode, as well as some NaHS and NasS 2 O3. The last-mentioned compound is formed readily by the oxidation of Na 2 S or of NaHS, and this explains why Na 2 S 2 O3 can be obtained by blowing air into solutions from which antimony has been deposited electrolytically. The crystallized salt is a common commercial article. Special conditions for the electrolysis are: ELECTROLYZING VESSELS: iron tanks with rectangular cross-section. ELECTROLYTE: aqueous solution of NaaSbSa. ANODES: lead plates. CATHODES: iron plates and the walls of the tanks. CURRENT DENSITY: 10 to 15 amperes per sq.ft. at the start; E.M.F., 2 volts; later 4 to 5 amperes per sq.ft. By this process not only is the solvent recovered in a com- mercial form, but also the sulphur of the ore. Inasmuch as it is difficult to prepare antimony solutions perfectly free from Fe, there is always more or less FeS formed on the walls of the electrolyzing tank, and upon the surface of the cathode, and this falls to the bottom with the metal that 154 METALLURGY drops off; hence this process does not yield directly a perfectly pure metal, but rather one that contains Fe. (C) Antimonial Lead, Hard Lead In the purification of lead bullion (cf. Silver, Desilverization, and Lead) by an oxidizing fusion, one of the first oxidation prod- ucts withdrawn is usually a mixture of PbO and Pb3(SbO 4 ) 2 . We have already seen in the above cited sections that this product is next subjected to a liquation during which a reducing flame 4, which may be regarded as containing half FIG. 164. Internal Dendritic Structure (X4o). a molecule of each of the other oxides. Sb2O 3 shows both acid and basic properties, but Sb2Os has an acid character. Sb combines directly with Cl to form SbCl 5 or SbCl 3 . HC1 dissolves Sb slowly. H 2 SO 4 oxidizes Sb to Sb 2 O 3 and an excess of this acid forms a sulphate, which hydrolyzes readily. HNO 3 oxidizes Sb to Sb 2 O 5 - Sb shows a great affinity for S ; the sulphides Sb 2 S 3 and ANTIMONY 159 Sb2Ss (the latter is unstable in the free state) have an acid character and unite with alkali- and alkaline-earth sulphides to form sulphantimonites and sulphantimo- nates, which are readily soluble in water. Tn the fused state the affinity for S is so great in the presence of alkali monosulphides that the sulphantimonites, e.g., Na^SbSs, will withdraw S from even Sb 2 S 3 , with sepa- ration of Sb, to form Na 3 SbS2, S, Cu, Fe and Pt. For the method for working this up, see Extraction. Concentration by Leaching and Precipitation, if necessary, with the aid of roasting. Directly applicable to ores, inter- mediate products and by-products which contain the Ni in the form of easily soluble salts, or free oxides and sulphides. After a preliminary roasting it is particularly suitable for sulphide and arsenide ores and intermediate products which also contain besides nickel other metals worth recovering (Cu > Co and precious metals). Contrary to the many statements in metallurgical literature, it is to be emphasized that silicates are absolutely unsuited for such work. An Old Process of Roasting, Lixiviation, and Precipitation was formerly used universally for working up mattes and speiss which contained Co, Cu and other metals besides Ni. It consists of the following separate operations: 1. DEAD ROASTING to Oxides. 2. SOLUTION OF THE ROASTED RESIDUE in acids (HC1 or H 2 S0 4 ). 3. PRECIPITATION by H 2 S or by a soluble sulphide (Na2S, etc.). PRODUCTS: Sulphides of Cu, Pb, Bi, Sb, As. These sul- phides are utilized in the works according to the predomi- nant metal. 4. OXIDATION of the residual solution with chloride of lime and neutralization with milk of lime. First of all, the iron is precipitated as Fe(OH) 3 . After the greater part of the iron has come down, the solution is filtered and the NICKEL 167 addition of chloride of lime and milk of lime continued. The cobalt is thus precipitated as hydrated Co 2 O 3 . 5. PRECIPITATION with Milk of Lime or Soda. PRODUCT: Ni(OH) 2 or NiCO 3 . Herrenschmidt's Process for working up ores carrying Co and Ni. This process is used on Herrenschmidt's own property for asbolite (18% Mn 2 O 3 , 3% CoO, 1.25% NiO, 30% Fe 2 O 3 , 8% SiO 2 , 5% A1 2 O 3 , i% CaO, i% MgO). It consists of the following operations: 1. OXIDIZING ROAST. 2. TREATMENT of the roasted residue with ferrous sulphate solution and introduction of steam and air. PRODUCTS: CoSO 4 and NiSO 4 in solution. 3. PRECIPITATION of the solution, which is kept weakly acid, with Na 2 S. PRODUCTS : a small amount of CoS and NiS. 4. THE SULPHIDES are filtered off, dried and given a sul- phatizing roast. 5. THE ROASTED PRODUCT is dissolved in water, and after treatment with CaCl 2 solution, CoCl 2 and NiCl 2 remain in solution while CaSO 4 is precipitated. 6. THE SOLUTION obtained in 5 is divided according to the ratio of Co to Ni. One part of this solution is completely precipitated with Ca(OH) 2 solution. The precipitate is converted into Co 2 O 3 and Ni 2 O 3 by means of air and chlorine. This precipitate is mixed with the remainder of the original solution so that the Ni 2 O 3 reacts with the CoCl 2 in solution to form Co 2 O 3 and soluble NiCl 2 . 7. FILTRATION: the filtered precipitate when dried is Co 2 O 3 which can be sold as such. 8. THE FILTRATE, containing NiCl 2 in solution, is precipitated by Ca(OH) 2 . 9. FILTRATION: the filtered precipitate of Ni(OH) 2 is dried and ignited, whereby it is converted into NiO. This is worked up as described below under B. The liquors containing CaCl 2 can be used again in Operation 5. 168 METALLURGY Borchers and Warlimont's Process accomplishes in a simple manner a working up of Ni ores that carry Co and Cu. It consists of the following operations: 1. SULPHATIZING ROAST at a temperature of about 500 C, in rotating iron drums, into which the air necessary for roasting is led. The drums are heated externally. PRODUCTS: SO 2 , which can be utilized in the usual way, and a roasted residue in which the Co and Cu are present almost entirely as sulphates, the Ni partly as Ni 3 S 2 , partly as NiO and only to a slight extent as NiSO 4 , while the Fe is converted for the most part into Fe 2 Os, although some FeSO 4 and Fe(SO 4 )s remain. 2. EXPOSURE of the roasted product to the air for several days, meanwhile working it about and wetting it. Hereby the FeSO 4 that is present converts any unchanged Cu 2 S into CuSO 4 . 3. LEACHING with slightly acid water. PRODUCTS: a solution of CuSO 4 and CoSO 4 , which is only slightly contaminated with NiSO 4 and FeSO 4 , and a residue which contains most of the Ni together with the original gangue of the ore and the Fe 2 O3 that has been formed. 4. PRECIPITATION with iron. PRODUCTS: cement copper, concerning the working up of which see Copper, and a weakly acid solution containing all the Co of the ore and a small amount of Ni, both in the form of sulphates. 5. PRECIPITATION of the solution obtained in 4, using CaS as precipitant; the latter is a by-product in the smelting of concentrated matte with CaO -I- C. PRODUCTS: a mixture of CoS with a little Ni 3 S 2 which is dried and used in the cobalt industry. The liquor obtained here may be evaporated to dryness and the residue used as a flux in operation i. 6. Matting of the leached residue obtained in Operation 3 adding to the charge, when necessary, substances con- NICKEL 169 taining S and free from Cu. At the same time other nickeliferous slags and by-products obtained in subse- quent operations may be added to the charge. PRODUCTS: waste slag and first nickel matte which is worked up further as described under A on page 162. As regards the method of obtaining metallic nickel see the following descriptions of Extraction and Refining. (B) Extraction The choice of a process for working up the products obtained by concentration work is determined by the character of the material (whether a concentrated matte, a concentrated speiss, or an oxide), the character of the finished product (whether pure nickel, or a nickel alloy), and finally, by the availability of a cheap fuel or other source of energy (water-power). The Roast-Reduction Process has in the past been most used because the concentration work, as described above, has yielded chiefly a concentrated matte or speiss. This process comprises the following operations: i. Dead Roast. Matte or speiss is roasted with great care in a reverberatory furnace until the last traces of S or As have been oxidized and removed. Sometimes it is necessary to add a little Na 2 CO 3 or NaNOs (or a mixture of both salts) in order to remove the last traces of As. APPARATUS: reverberatory furnaces. PRODUCTS: SO->, As2C>3, which, however, on account of the low concentration in which they are present in the gases, cannot be utilized; hence, wherever this method is employed its effect is very injurious to vegetation in the adjacent country. The roasted residue consists almost entirely of NiO. In case one or both of the above-mentioned Na salts were used to effect the removal of the As, the residue also contains Na 3 AsO 4 , which can be extracted with water. 2. Reduction. For this work, there is the residue obtained 170 METALLURGY under i as well as oxidized products obtained in concen- trating work as described under A. The reduction may consist either of a reducing roast or of a reducing smelt. The reduction by means of a reducing roast is effected by taking the NiO obtained in previous operations, stirring it up with flour paste to a plastic mass, rolling it into plates, and cutting these lengthwise and crosswise into small cube-shaped briquettes, which, after being dried in retorts or crucibles, are embedded in charcoal powder and heated within the same vessels to the reduction temperature of NiO, usually using producer gas as fuel. It is not intended hereby to effect a complete fusion, but merely to weld together the reduced Ni; the cubes, naturally, are smaller in size than the cubes of NiO that were intro- duced into the reduction furnace. The " cube nickel " thus obtained is preferred in this form by buyers for many purposes. It is, for example, suitable for purposes of alloying, but it is not suitable for working into nickel wares on account of the presence of carbides as well as some NiO, which impurities cause brittleness. A direct reduc- ing smelt of nickelous oxide, which may take place in a crucible furnace or in a shaft furnace, is usually employed when the oxidized concentration product still contains metals that should either be eliminated by fur- ther separation processes (cf. Nickel Refining), or when it contains large amounts of other oxides (CuO), which it is desired to reduce at the same time so that a nickel alloy will be formed. Roast-Reaction Smelting in reverberatory furnaces or in converters does not give rise to metal, as is the case when working with copper products. The reason for this was found, in the author's laboratory, to be as follows: The reaction Ni 3 S 2 + 4NiO = ;Ni + 2SO 2 takes place, to be sure, at temperatures above 1400 C, but NICKEL 171 only very slowly even when the heat is raised above 1600 C, and with high oxygen concentration in the blast. On the the other hand the reaction takes place rapidly even at relatively low temperatures. More- over, metallic Ni alloys very readily with its sulphides and with NiO. Reaction smelting, therefore, yields under the most favorable conditions only an alloy consisting of much NiO with little Ni; it can never lead to metallic nickel under any of the methods of working which have been tried up to the present time. Desulphurizing Fusion of Concentrated Matte. Borchers and Lehmer's process consists of smelting the concentrated matte with lime and carbon, preferably in an electric furnace. If the matte was entirely free from other metals, then this smelting can be carried out so that a fused, pure nickel will be obtained directly. Since, however, the sulphide nickel ores usually contain precious metals (Au and members of the Pt group), the product is an alloy of Ni with these metals, from which pure nickel is obtained by electrolysis (see Nickel Refining). The slag of CaS can be used either as a sulphide flux in the smelting of matte, or, as in the above-described process of Borchers and Warlimont, as a precipitant for Co and Ni solutions. (C) Nickel Refining As a raw material for the preparation of pure nickel, the crude Ni prepared as described under B is usually taken. It is pos- sible, however, to make use of concentrated matte, or even certain ores, as starting material. 172 METALLURGY Pure Nickel from Crude Nickel. The crude nickel obtained by a reducing roast, or the nickel obtained by a reducing smelt of nickelous oxide, contains, as has already been mentioned, C and NiO so that it is not suitable to be used in the manu- facture of nickel wares (foil, wire, etc.). Refining Fusion. The removal of the C offers no difficulties, as it takes place when the crude metal is melted ; but the last traces of NiO cannot be removed by means of carbon. This reduction of the NiO was first accomplished by the method of Fleitmann, who fused with Mg, which is an extremely energetic reducing agent; then Basse and Selve accom- plished the same end by mixing in as much as 3% of MnO2 before carrying out the reducing roast. The MnO 2 is reduced, together with the NiO, and serves to remove the NiO when the cube nickel is heated in carbon-free crucibles; naturally it prevents the forma- tion of NiO by atmospheric oxygen when the metal is melted. Electrolysis is employed if the nickel contains precious metals. Since the solution tension of Ni is on the other side of H, the concentration of H ions must be kept as low as possible, contrary to the practice in the electrolytic refining of Cu. This is accomplished by limiting the amount of free acids having high dissociation constants (HC1, H 2 SO 4 ) or by using acids with low dissociation constants (H 3 BO 3 , HsPO4, and organic acids). Unfavorable changes in con- centration are avoided by moving the electrodes and heating the electrolyte. When these requirements are fulfilled, the electrolytic deposition of Ni offers no difficulties worth mentioning. The conditions for carrying out the elec- trolysis are as follows: ELECTROLYZING TANKS: made of wood with lead lining. ANODES: Cast plates of impure nickel, containing precious metals with S up to 3%. CATHODES : Pure Ni. ELECTROLYTE: NiSO.j or NiCU solution, the former being NICKEL 173 preferred. Concentration 4 to 14 oz. Ni to the gallon not under 0.01% or over 0.25% of free acid. TEMPERATURE: 50 to 90 C. (122 to 194?.) CURRENT DENSITY: 5 to 28 amperes per sq.ft. POTENTIAL: i to 1.3 volts with 15 or 20 amperes per sq.ft. REACTIONS DURING THE ELECTROLYSIS: under the above conditions it is possible, even when the crude metal con- tains up to 0.5% Cu, to deposit a sufficiently pure Ni (0.1% to 0.2% Cu). It is important to have enough S present not only to precipitate the Cu as Cu 2 S, but also to make the last traces of Fe form compounds of Cu 2 S with FeS. Such compounds are attacked by the elec- trolysis less in proportion as the current density at the anode is kept low, i.e., the more the anode surface is increased in proportion to the size of cathode surface. In designing an electrolyzing plant for Ni, therefore, the anodes should be made relatively large and the cathodes small. Pure Nickel from Concentrated Matte. The matte which is obtained as end product in the concentration of sulphide ores contains up to 76% Ni. It then has 23% to 24% S, up to 0.4% Fe, 0.1% to 0.2% Cu, and small amounts of SiO 2 due to the presence of a small amount of enclosed slag. Accord- ing, to the above-cited researches of Bornemann, unlike the concentrated matte of copper smelting, it does not contain a pure sulphide of nickel, but rather a solution of Ni in NisS2. This matte already has many metallic properties and is suited, as has been demonstrated by Borchers and Giinther, for a Direct Electrolytic Treatment. The conditions, as established by Giinther are partly the same as in the electrolysis of crude Ni. The points of difference may be summarized as follows : ANODES: concentrated matte, cast into plates. CURRENT DENSITY: 23 to 26 amperes per sq.ft. POTENTIAL: 3 volts. REACTIONS DURING THE ELECTROLYSIS: Of the constituents of the matte, it is chiefly Ni that passes into solution at 174 METALLURGY the anode, leaving behind free S, which remains adhering to the electrode as a coherent but porous crust until the matte has nearly disappeared, and thus influences the progress of the electrolysis scarcely at all. In this crust there are present about 80% free S and 20% of a mixture of Cu, Fe and Ni embedded in it. Whereas the original . matte contained less than 0.2% Cu, the mixture of embedded sulphides will contain over 12% Cu with about 51% Ni and 3.5% Si + C. From this composition, and from the low Cu content of the Ni that is deposited on the cathode, it follows that Cu is combined with S in the crust and that the latter unites with the sulphides of Fe and Ni to remain, for the most part, adhering at the anode; this is confirmed by the results obtained in the electrolysis of crude Ni that contains S. Electric Smelting of Concentrated Nickel Matte with lime and carbon directly yields, as already mentioned under B, pure Ni provided the matte contains no foreign metal other than S and Ni. Pure Nickel Directly from Ores. Extraction of Ni by Means of CO, Mond's Process. This is based upon the property that Ni possesses of combining with CO to form a readily volatile compound called nickel carbonyl, Ni(CO)4. The process is applicable only to ores and roasted products that contain the Ni in the form of free oxides. The operations are as follows: 1. REDUCING ROAST in retorts heated to 300 C. with gases containing H. The NiO is changed by this treatment into porous Ni. It is important to maintain a reducing temperature as low as possible so that a large surface of the reduced Ni will be obtained. 2. PASSING OF GASES CONTAINING CO at 100 C. and 15 atmospheres pressure over the charge. The temperature of 100 C. is chosen because it can be kept constant easily by means of steam. As regards the pressure, Dewar found that Ni(CO) 4 is stable at NICKEL 175 50 under a pressure of 2 atmospheres. 100 " " 15 180 " " 30 250 " " 100 3. Deposition of Ni from the Ni(CO) 4 , which depends upon the ready reversibility of the reaction The vapors of Ni(CO) 4 escaping from the vessels under pressure can be decomposed into Ni and CO either directly, or, after a preliminary purification which con- sists in cooling and freeing from mechanical admix- tures, by simply heating at 200 C. under ordinary atmospheric pressure. The CO is collected and used again in the process. Properties of Nickel : SPECIFIC GRAVITY: 9. COLOR: light gray. MECHANICAL PROPERTIES : possesses a high degree of ductility and tenacity so that it can be rolled into foil and drawn into wire. STRUCTURE of Cast Ni: similar to that of ferrite (see iron); rolled Ni is very fine grained. MELTING POINT: 1451 C. (2644 F.). THERMAL AND ELECTRICAL CONDUCTIVITY: about 0.2 that of Ag. ALLOYS: with most metals. Important alloys are coin metal, Niand Cu; German silver, Ni, Cu and Zn; nickel steel, Fe and Ni. In the extraction of Ni the tendency to alloy with its own sulphides and arsenides, as well as with NiO, comes into consideration. CHEMICAL BEHAVIOR: The metal resists oxidation fairly well at moderate temperatures. Waste in forging and rolling is slight. Solution tension in acids and salts is greater 176 METALLURGY than H( +0.223 vo ^ ts toward H). It dissolves, therefore, quite slowly in HC1 and H^SC^, more rapidly in HNOa, forming under normal conditions a bivalent Ni cation. In the molten condition the solution tension is greatest in arsenides, next in sulphides; in arsenides the solution tension is greater and in sulphides less than that of Cu. IRON Sources Natural Sources : NATIVE, as a constituent of the earth's crust it is of rare occurrence; large lumps of meteoric iron are, however, sometimes found (up to 55,000 Ibs. in weight). As OXIDE AND HYDRATED OXIDE in the following ores: Specular Iron Ore l Hematite } Fe 2 O 3 with 70% Fe. Red Iron Ore Magnetite, or Magnetic Iron Ore: FeO.Fe 2 O 3 with 72.4% Fe. Brown Iron Ore j Limonite, Bog Ore !> Fe 2 O 2 (HO) 2 to Fe 2 O(OH) 4 . Brown Hematite J As SULPHIDE in M " Marcasite Magnetic Pyrites, or pyrrhotite, FeS with some FeS 2 , usually nickeliferous. Combined and mixed, as FeS with FeS 2 , in numerous other sulphides. The sulphides themselves do not come into considera- tion as ores of iron until after they have been utilized for the manufacture of H 2 SO4j they are then roasted to oxides so that they retain but very small amounts of sulphide. (Roasted or " burnt " pyrite, see below). SALTS. Spathic Iron Ore, or Siderite, FeCO 3 with 48.27% Fe, usually carrying some MnCOs, CaCOy and clay. Blue Iron Ore, or Vivianite, Fe 3 (PO 4 ) 2 , which seldom 177 178 METALLURGY occurs by itself, but accompanies other ores (e.g., limonite). Other Sources : METALS: Scrap from the mechanical working of iron and old metal. OXIDES: Roasted pyrite, the roasted residue from sulphide ores. Hammer scale, the oxide crust formed on iron heated to redness and detached in the mechanical working. SALTS: Basic silicates and phosphates, slags from the refining of iron. (A) Concentrating and other Preliminary Operations. For Concentrating Iron Ores it is often desirable to employ: Magnetic Concentration which is carried out largely when the ore is magnetite. Ores carrying hematite are also concen- trated by the magnet, after they have been converted into magnetite by roasting. The Briquetting of pulverulent ores and other raw material has been carried out to a considerable extent in recent years. The requirements of a good ore briquette for blast furnace work, have never been satisfied. Chemical Preparation. Dissociating Roast, usually combined with an oxidizing roast, is carried out to effect the DEHYDRATION of hydrated ores (limonite). OXIDATION of magnetite. DECOMPOSITION AND OXIDATION of carbonates (spathic iron ore). WORKING up of pyrite for sulphuric acid. The Apparatus used for this work consists of simple shaft furnaces, the so-called calcining furnaces which usually work with natural draft. They either take both ore and fuel, or the ore alone, when they are fired externally with solid or gaseous fuel. The discharge is usually through a saddle- IRON 179 180 METALLURGY shaped bottom to facilitate the taking out of the solid charge through side openings. The apparatus that serves specially for the roasting of pyrite has been described under Copper. FIG. 166. Old Siegen Kiln (B) Cast Iron Reducing Roast. Various attempts have been made to make use of this operation and produce the so-called " iron sponge," a product which is metallic, but is not fused until it comes to the refining operation. Such processes have never been adopted permanently, even for the preparation of iron for special pur- poses in which it was thought desirable to have the iron in this form. The Reducing Smelt is to-day the only process used in the iron industry for the production of cast iron. As Reducing Agent, carbon in the form of charcoal is seldom used, but rather coke and the CO formed in the furnace. The Other Fluxes are determined not alone by the nature of the gangue to be removed, but by the purpose which it is desired to accomplish, and by the demands that are to be IRON 181 placed upon the iron. Thus for slagging off the SiC>2 and clay which is frequently present in the ore, CaO in the form of CaCC>3 is added to the charge, and the amount added depends upon whether it is desired that the cast iron pro- duced shall have little or much Si. To produce a cast iron rich in Si, it is necessary to form a more basic silicate and a slag richer in AloOs than when an iron low in Si is desired. For slagging off the S from ores containing some sulphide, the slag should be kept as basic as possible with CaO. The addition of an ore rich in Mn can also serve to advantage in this case. For Fuel, coke is generally used and charcoal but seldom. In some few cases where coke is expensive and the ore deposits are in the vicinity of cheap water power, electricity is now to be considered as a possible source of heat. The blast requisite for the production of the desired amount of heat by the combustion of carbon is given a preliminary heating of 750 to 900 C. and compressed up to one atmosphere. Apparatus : BLAST FURNACES. These shaft furnaces are sometimes as much as TOO feet in height. The shaft rests upon pillars and is built of masonry with iron bands, or of an iron casing, and has an inside diameter of 26 ft. at the widest part, narrowing toward the throat, which has about two-thirds this width. The iron blast furnaces usually have internal crucibles. The hearth has an inside diam- eter of 13 ft. and a maximum height of nj ft. It contains the tuyeres for the blast, the cinder notch, and the tap hole. One tuyere is reckoned for each sq. yd. of hearth surface so that with a diameter of 12 ft. there are n or 12 tuyeres. These tuyeres are hollow bronze or copper castings which can be cooled. The boshes rest upon the hearth and reach into the shaft to about two-fifths of the whole height of the furnace, reckoning from the bottom of the hearth, and they widen as they go upward until they reach the widest part of the furnace. The boshes and hearth are 182 METALLURGY both cooled by spraying with water. Large furnaces use about 525 gallons of water per minute. Iron blast furnaces work with a closed throat, for the hot escaping gases are rich in CO and can be utilized as a source of heat and power. The "down-comer" is usually a wide irtnnnrtntlTn ~U LJ L.J LJ L LJ LJ [J LJ FIG. 167. Blast Furnace with Stoves (Friedrich-Alfred Smelter) Rheinhausen. central sheet-iron tube passing through the bell and hopper- feed in the throat. In some constructions, the gas is taken off laterally from the closed throat of the furnace. HOT BLAST. To-day regenerator stoves of about the same dimensions as the blast furnaces usually form an IRON 183 inseparable whole with the latter throughout all of their operations. The regenerative system requires the setting up of at least three stoves; to be on the safe side it is customary to allow four stoves for each blast fur- nace. In the cylindrical stoves of Cowper's design, there FIG. 168. Blast Furnace with Dust Catcher and Blast Stove. is found a shaft into which the hot gases from the throat of the blast furnace enter at the bottom for the CO to be burned by air which has been heated by passing through flues in the walls of the shaft. The remaining space of the regen- erator is loosely filled with special brick, the spaces between which form numerous little channels. It forms the real 181 METALLURGY FIG. 170 F.G.169 FIG. 172. FIG. 173. IRON 185 heat reservoir. The products of combustion are divided into innumerable gas currents by these brick, and in this part of the shaft they are carried downward and then led away from the open space at the bottom. After the stove has become hot enough, the gas and combustion air are led into a neighboring regenerator, and now the air intended for the blast furnace is passed through the hot chamber in the opposite direction to that previously taken by the gases from the blast furnace. This heated air collects above the part of the regenerator that is filled with brick and passes downward through the combustion chamber and thence into the main pipe for the blast. Since this air must be heated to between 750 and 900 C., the iron pipes through which it passes are provided with a fire- brick lining. Even the bustle pipe leading to the sep- arate tuyeres of the blast furnace is lined or else placed in a hollow tube with non-conducting layer of air inter- vening. REACTIONS IN THE BLAST FURNACE. The changes within the more important temperature boundaries, which the solid charge experiences as it passes from the throat of the furnace downward, are the following: With furnaces that are 75 to 100 ft. high between the throat and the bottom of the hearth, the temperature at the top ranges between 150 and 300 C., whereas in the smelting zone the temperature may reach 1500. 150 to 400 C. Drying of the charge. Breaking down of hydrates. Begin- ning of reducing reactions. 400 to 1000 C. According to the studies of Boudouard on the equilibrium of the reaction 186 METALLURGY the speed of the reaction in the direction of the upper arrow increases from about 400 to 700, but from there on the tendency for the reaction to take place in the opposite direction becomes more and more pronounced. Here the efficiency of CO as a deducing agent rapidly diminishes. As regards the concentration of the reaction gases, it has been found that the ratio of CO:CO2 must not be less 25m. As reducing agent FIG. 174. FIG. 175. than i : i at the most favorable reaction temperature (700). At such a concentration of CC>2, the reducing action of the CO stops almost entirely even at 700 C. In this zone, therefore, the greater part of the oxides of iron, namely Fe 2 O 3 , Fe s O 4 , and FeO, are reduced by CO, but at about 800 there is a dissociation of the car- bonates, CaCO 3 , FeCOs and MnCO 3 , into CaO, FeO, IRON 187 MnO, and CC>2, and the effect of this is to greatly increase the concentration of the CO2- From 1000 upward the C begins to dissolve in solid Fe. Disregarding other influences (Si, Mn, etc.) it has been established that the solubility of C in Fe increases as the temperature rises (cf. also p. 190). From 1300 upward solid C is practically the only reducing agent, particularly for FeO, MnO, SiC>2, phosphates, and silicates; FeS also reacts with CaO and C to form Fe and CaS. The blast, as it passes through the, furnace in the opposite direction, of course loses its O rapidly by reason of the combustion of the C. The more the blast has been heated before it comes in contact with the coke, the greater will be the formation of CO(C+-^ = CO 2 and CO 2 + C = 2CO). The CO, after it is formed, as the above equations show, will be changed repeatedly into CO2 and back again to CO. The gas finally leaving the throat of the furnace has a composition of 22 to 28% CO, 16 to 8% of CO 2 , 63 to 57% N, and a low per cent of hydrocarbons and free hydrogen. With the present-day dimensions of blast furnaces, about 500 tons of pig iron are produced in a furnace during a day of 24 hours. The weight of a single charge of ore (ore and slagging fluxes) reaches 15 tons/ for which 3 to 7 tons of fuel are required. The charge remains in the furnace for from 18 to 26 hours. Aside from the content of the ore and the quality of the other constituents of the charge, the capacity of the furnace is dependent particularly upon the nature of the pig iron that it is desired to produce. A furnace capable of producing 100 tons of white cast-iron, can yield at the most but 80 tons of gray iron, and still less of other kinds, e.g., metal run- 188 METALLURGY ning high in Mn or Si. Furthermore, in respect to the con- sumption of fuel, the production of white iron is the most favorable. For such iron, 100 to no tons of coke are required to produce 100 tons of the metal. The coke consumption rises with other qualities of iron. Concern- ing the nature and value of the Products of the Blast Furnace, we may summarize them as follows : i. PIG IRON, an alloy of Fe and compounds of Fe and Mn with C, Si, P, S, and O, from which on solidifying individual constituents, such as C in the form of graphite or amorphous carbon, carbides, silicides, etc., may separate out. The total C content is over 2% and usually under 5%. The nature of the ore, the conditions of working, and the qualities required by the purchaser, have led to the production of quite a number of different commercial grades of pig iron, but the qualities, as in the case of most pure irons, arejdependent largely upon the amount and jnanner of^com^ina^iQiL-QLthejC present. The latter^ on" its part, is influenced by other constituents found in the iron, particularly Si and Mn; a short discussion of the relations between Fe and C will explain best the nature of the different products obtained in the iron industry. If we start with a perfectly liquid melt, it is known from the interesting researches of Roberts- Austen, Wiist and Goerens, and others, that all the C can exist combined with Fe as Fe 3 C (which is called cementite) until the amount of C present reaches 6.66%. A higher C con- tent scarcely comes into consideration in practical work; with a lower C content there may exist in the fusion a solution of this carbide in Fe. If the temperature falls, it may happen when the amount of Fe 3 C present is large that, just before the whole melt solidifies, a decomposition of the Fe 3 C into 3Fe + C takes place. Directly after such decomposition free C is therefore embedded as pure graphite in pure Fe, and as the melting- IRON 189 point of the latter is higher than that of the solution (about 1500 C-) it begins to solidify within the liquid melt and thereby envelopes the solid graphite. As long as the lat- ter possesses this specifically heavy shell it remains uni- formly distributed in the melt. When, therefore, the cool- ing does not takeplace toojslowly, a pig iron is obtained ^hich containsTarge crystals of graphite uniformly dis- tributed through it. When, however, the temperature is kept high for a long time, the crust of Fe gradually becomes saturated again with C, becomes liquid and the graphite now rises to the surface and forms a scum called " kish." The melt, in which all this has taken place, finally solidifies on further cooling at 1130 C., with about 4.2% C chemically combined with Fe. From molten iron with less than 4.2% C and more than 2% C, crystals first separate from the melt which are to be regarded as solid solutions of Fe 3 C in Fe (mixed crys- tals). The saturated solution, or in other words the mixed crystals richest in C, contain 2% of C, they correspond, therefore, to a solid solution of about 2 Fe 3 C + i5Fe. The mother metal from which these crystals are deposited thus becomes enriched in Fe 3 C until the above-mentioned 4.2% C content is reached. Then the remaining metal solidifies at 1130 C. At the concentration 4.2% C, lies a eutectic point between the melting points of the saturated mixed crystals, 2Fe 3 C -f- i5Fe (2% Fe) and cementite, Fe 3 C (6.66% Fe), and this eutectic corresponds closely to the composition Fe 3 C + Fe2. As Wtist and Charpy both showed independently in 1905, this eutectic solidifies under all conditions, no matter what other substances may be present, at 1130 C. Under the conditions corresponding to the line aB (Fig. 176), there exists a solid mass of free mixed crystals em- bedded in eutectic; at the point B there is only eutectic present; and at the line BC graphite and cementite in eutectic. 190 METALLURGY In every cast iron with more than 2% C, there tends to take place after solidification a decomposition of Fe 3 C into 3Fe + C (graphite or amorphous carbon, temper carbon) provided the Fe 3 C has already crystallized from the solution or has had an opportunity to crystallize. Substances which favor the crystallization of this compound from the liquid or solid solution (e.g., Si, Al?) naturally favor as well the progress of the reaction Fe 3 C = 3Fe + C; substances 1500 1400 1300 1200 1100 1000 900 800 M A \ ~^ ^ ^ ] jiqui< I \ Liq lid ^> \ S / N5 ixedC I V S : 3C \ \ / / Graj hite Mix I edCrj e n +F< stals )3 C 5 / f + ll quid ^ Mixlci. ^rsst. H-Knt. Kl Hectic . Re, O-j-T? lit. / /r G / Mixe ICrys tals - 1- Gra phite 4-Te mper ^arbo i \ / O^ x/ VOO 600 500 1 Ft rrite \- Gri phite -fTe mper Darbo v i 0.5 1,0 1,5 2,0 2,5 3,0 3,5 4,0 4,5 5,0 5,5 # C. FIG 176. which increase the solvent power of Fe for Fe 3 C (e.g., Mn) increase the stability of the Fe a C and hence tend to 'prevent the formation of free C. From this point of view the nature of the two principal kinds of cast iron appears in the follow- ing light: WHITE IRON is a supercooled solution and is, therefore, to be regarded as representing a meta-stable system between Fe 3 C and Fe, in which the reaction Fe 3 C = IRN 191 3Fe + C has not been allowed to take place. The chilled white iron consists of cementite and pearlite (see Fig. 178), also containing, when the cooling was very rapid, some mixed crystals of the concentration FeaC : Fe n . If such an iron is kept for a long time at about 1000 C. the compound FesC begins to crystallize from the solid solution and then the Fe 3 C decomposes rapidly, although the free C does not appear as graphite, but as the amor- phous, readily-oxidizable, temper carbon (see Fig. 180, P- 195)- GRAY IRON represents the more stable system Fe Fe 3 C C. It has had time, at the different temperatures and con- centrations, to reach more or less completely a state of equilibrium, i.e., during the cooling FeaC has had opportunity to crystallize out and dissociate into 3Fe and C and the latter is now to be found in the form of graphite, although it is not impossible that some temper carbon may have been formed in the vicinity of 1000. (Fig. 179.) As has been pointed out, the separation of C is favored by the presence of Si, even when the C content is low, and to promote this separation from i to 2.5% Si, or even 5% Si in special cases, is often added to the iron. * Whereas gray iron is used for foundry purposes as well as for the production of wrought iron, white iron is used almost exclusively for the latter purpose. A distinction is further made between foundry iron and Bessemer iron (Si high, P low), open hearth iron, basic Bessemer iron (high in Mn and P), spiegeleisen or ferromanganese (rich in Mn, low in P), and charcoal iron (rich in C). In the many constituents of the different kinds of pig iron, only a few of the particularly important solidification phenomena have been explained. The first accurate equilibrium diagrams between Fe and C were those of Roberts-Austen and of Roozeboom. Subsequently Heyn gave different diagrams for the meta-stable and stable 192 METALLURGY systems, while Benedicks improved the original diagram of Roozeboom. Since Wiist and Charpy have determined the freezing-point of the system Fe Fe 3 C to be 1130 for all cases, and Wiist and Goerens have proved that graphite is always produced by the decomposition of Fe 3 C, Goerens' diagram (Fig. 176) may be taken as correct for the stable and for the super-cooled system Fe - Fe 3 C. To understand the terms used in metallography, the following definitions will serve: FIG. 177. Ferrite (X5oo). FERRITE, chemically pure iron: a-iron, which is magnetic and free from C, passes at 780 into /?-iron, which is non-magnetic and also practically incapable of dissolving C. This second form of iron exists only between 780 and 800 C. Above 880 it changes to ^-iron, which is likewise non-magnetic, but is capable of dissolving C or Fe,C (Fig. 177). CEMENTITE is iron carbide, Fe 3 C (Fig. 178). AUSTENITE and MARTENSITE are solid solutions of Fe 3 C in f-iron. They exist as mixed crystals of varying con- centrations. The saturation point of these solutions lies at 2% C corresponding to the composition 2Fe 3 C -f IRON 193 i5Fe. A eu Lectio between these saturated mixed crystals and Fe 3 C lies at 4.1% C, corresponding to a composition TROOSTITE: a colloidal solution of Fe 3 C in Fe. SORBITE: mixtures of Fe, Fe 3 C, and solid solutions of Fe 3 C in Fe. PEARLITE: the eutectic between ferrite (Fe) and cementite (Fe 3 C). Its C content is 0.9%, corresponding to Fe 3 C + 2oFe (Figs. 178 to 180). FiG. 178. White Cast Iron. Cementite (White) with Lamellar Pcarlite (X5). GRAPHITE (Fig. 179): concerning its formation see p. 188. TEMPER CARBON is non- graphitic C which separates out from white iron by keeping it for a long time at a tem- perature near 1000, during which time the finely-divided cementite changes into a mixture of ferrite, pearlite and temper carbon (Fig. 180). This form of carbon is more readily oxidizable than graphite or carbide carbon (cf. Malleableizing, p. 198). 2. SLAG. SiO 2 and A1 2 O 3 are almost invariably present in the 194 METALLURGY gangue of iron ores and in the ash of the fuel. Again, the oxides of the alkaline earth metals, particularly of lime, are always present in the slag, since CaO is used as a slag- ging flux (cf. p. 181). The chief constituent of blast furnace slag, therefore, is a calcium-aluminium silicate, and this silicate is approximately neutral or weakly basic, according to the impurities in the raw materials and according to the desired nature of the iron to be produced, and it ss more or less rich in other metallic oxides (FeO, MnO, etc.). Fie. 179. Gray Cast Iron. Graphite (Black); Cementite (White) with Lamellar Pearlite The basic slags crumble readily after exposure to the atmosphere, but when it is desired to utilize the less basic slags they are usually granulated by allowing them to flow into water as they are tapped from the furnace. The more basic slags, particularly of foundry iron, when they contain suitable amounts of SiC>2, A1 2 O3 and CaO, are used in the preparation of Portland cement. FURNACE GAS. In discussing the reactions of the blast furnace, it was mentioned that a gas escapes from the IRON 195 top of the furnace containing up to 28% of CO (p. 187). One cubic foot of such a gas possesses a heating value of 89 B.T.U. For one ton of pig iron, there are about 160,000 cu.ft. of this gas. A part, half at the most, serves to heat the blast that is to be used in the furnace, and there remains from 80,000 to 90,000 cubic feet of the gas for other purposes. Until recently this gas was used for heating boilers which served as a source of power for FIG. 180. Cast Iron showing Temper Carbon. (Black); in Ferrite (White), Surrounded by Lamellar Pearlite. moving the raw materials and the products of the fur- naces; to-day the gas is being utilized chiefly in gas engines. The blast-furnace industry, however, requires but a quar- ter of the power that can be produced thus, and if this additional power is not needed for other work in the plant (e.g., for rolling milJs) it can serve for outside pur- poses (e.g., for electric plants). The amount of power thus available can be easily computed. After making the deduc- tions as just outlined for the gas that is utilized in connec- 196 METALLURGY tion with the blast-furnace work, there remains 75% of 80,000 cubic feet of gas, or from a 3oo-ton furnace, 60,000X300=18,000,000 cubic feet of gas in 24 hours; 125 cubic feet of the gas yield approximately i horse power hour. Thus 18,000,000 cubic feet from a 3oo-ton furnace will yield 18,000,000 hourly about = 6000 horse power hours. 125X24 (C) Iron Alloys and Iron Compounds Besides pig iron, certain alloys of iron and iron compounds to be used in the preparation of different kinds of forgeable iron and steel are obtained by smelting in blast furnaces. Spiegeleisen and Ferromanganese, see Manganese. Ferro-Chrome, see chromium. Ferro-Molybdenum is obtained in the process of Borchers and Lehmer by smelting molybdenite (MoS 2 ) in an electric furnace with lime, carbon, and Fe or a sulphide or oxide of iron. Ferro-Vanadium is prepared by a reducing fusion of ferrous vanadate with carbon and enough Fe in the flux so that the resulting alloy will contain not more than 25% Vd. The fer- rous vanadate is obtained from sodium vanadate, which is prepared by an oxidizing roast of vanadium ores with Na^COs or NaCl, or by fusing lead vanadate with soda; the sodium vanadate is extracted with water and the clarified solution is treated with FeSO.i solution. Ferro-Silicon can be produced like pig iron in the blast fur- nace by smelting iron ores with a gangue or flux rich in SiO 2 . In such cases the aim is to produce a slag rich in AUOa. By the blast furnace method the Si content may be made as high as about 16%. Richer alloys, with 25%, 50% or more of Si, are usually obtained in electric furnaces by the reduction of mix- tures of iron ores and sand. Ferro-silicon has also been obtained as a by-product in the manufacture of CaC 2 . In this case a coal with an ash high in SiO2 was used. Iron ore was then added to the charge and a ferro-silicon with about 20% Si obtained. IRON 197 (D) Forgeable Iron The equilibrium diagram, Fig. 176, p. 190, contains a point a, which corresponds to 2% C, and represents a solid solution of the composition 2Fe 3 C + i5Fe. This is the saturation point of f-iron for Fe 3 C and represents, at the same time, the upper carbon limit for the forgeability of iron. Although in practice the upper limit for a forgeable iron is taken as 1.6% C, still in the light of the relations between Fe and C which have been sufficiently cleared up scientifically, we may regard as forgeable iron all alloys lying between pure Fe and the saturated solution of 2Fe 3 C in i5Fe. At S the C content amounts to 0.9%, corresponding to a composition of Fe 3 C + 2oFe (pearlite). Although with a C content considerably lower than 0.9%, there is an increase in hardness when the metal is heated and then rapidly cooled, this increase in hardness by such treatment begins to be very marked at 0.9% C. At this point a solution is reached from which, on slow cooling, pure Fe (ferrite) can no longer crystallize out as an independent constituent of the alloy, but is contained only in the eutectoid called pearlite, Fe 3 C + 2oFe; with higher C content the eutectoid is formed in the presence of Fe 3 C. The characteristic hardening of steel has been explained by assuming a transformation of pearlite into martensite by heating and maintaining this solid solution by quenching; on the other hand, the transformation of steel into the malleable con- dition has been attributed to a transformation of martensite into pearlite, either by reason of slow cooling or by keep- ing the quenched steel at moderate temperatures (below 700) for a long time. The more the C content falls below 0.9%, the more ferrite exists in the presence of pearlite. It is, however, easy to understand why a steel with less than 0.9% C can be hardened. When 0.9% C is present, the sample can contain 100% of pearlite, and therefore a steel with 0.3% C contains 33.3% of pearlite which can be transformed into martensite. On the other hand, it is plain that the hardening power and malleability 198 METALLURGY will not disappear when the C content is above 0.9%, for beside the cementite, which now occurs in ever-increasing amount, it is still possible for pearlite to be formed and it is not until 2% C is reached that the amount of pearlite is so small in comparison with the amount of cementite present that the malleability prac- tically ceases. The property of hardening can be caused with very low C content by the presence of other substances such as, for example, Cr and W. The transformation of pig iron into forgeable iron must evidently consist in diminishing the amount of C present and removing as completely as possible the other undesirable constituents of the crude iron. In all cases this is accomplished by oxidation processes. In one case (malleableizing) only C is removed, but in all other cases, other constituents are removed simultaneously. At the same time it is impossible to prevent a part of the Fe itself from being oxidized and passing partly into the slag and partly into the purified iron as FeO, so that a further reduction is necessary to remove the oxide. Malleableizing. In 1903 to 1905 F. Wlist and some of his students proved experimentally that by exposing white iron for a long time to a temperature of about 1000 (annealing) there is a decomposition of the dissolved Fe 3 C into3Fe + C in such a way that the C is not deposited as graphite, as in the formation of gray iron, but entirely in an amorphous condition such that it is readily susceptible to oxidation. As oxidizing agents, oxide iron ores, usually Fe 2 Os, but sometimes FeCO 3 , are employed. The iron objects to be malleableized, already in the form of finished castings, are packed in this ore, gradually brought to a temperature lying between 900 and 1000 and kept there about a week. Until recently it has been held in metallurgical literature that the oxidation of the C begins at the surface where the -oxidized ore is directly in contact with the iron, and that then the solid C migrates from the interior to the carbon-free outer layers, where it is oxi- dized in turn. The author, in his Inorganic Chemistry, opposed IRON 199 this view in 1893; in 1908 Wiist submitted indisputable experimental proof that the decarburization is brought about by the aid of CO 2 , which is obtained partly by the contact of particles of iron containing C with the Fe 2 O 3 , and partly from the combustion gases of the furnace. Wiist showed, more- over, that these gases diffused into the interior of the castings, where the most favorable conditions exist (cf. diagram 174) for the progress of the reaction CO 2 + C = 2CO, while the CO thus formed is again converted into CO 2 by the oxidized ore, and can once more take part in the reaction. FIG. 181. Charcoal Hearth. Production of Wrought Iron. These methods of refining iron are the oldest in the development of iron metallurgy. Among them are: The Charcoal Hearth Process. This is still used in some places (Sweden, Styria, Russia) where a sufficient supply of charcoal can be obtained cheaply. The " hearths " or " forges " contain a hollow pit, usually lined with iron plates, and are provided with a tuyere in the high back wall. The cast iron is melted upon charcoal but with a blast of air sufficient for the necessary oxidation, and directed so that the liquid iron must drop through the blast zone. A 200 METALLURGY glance at the diagram, Fig. 176, shows that the melting ^ pointjofiron rises as its purity increases. The iron, that was at first liquid, collects in the bottom Tfearth as a pure viscous metal (blooms) which is freed from any included slag by mechanical working (hammering, rolling), and welded to a compact mass. Puddling. This is essentially the same kind of work, but differs by melting the cast-iron in a reverberatqry^jurnace, in which the oxidation is assisted by the presence of iron oxides, added in the form of slags rich in Fe 3 O 4 , and waste from the mechanical working of forgeable iron (e.g., FIG. 182. Longitudinal Section of Puddling Furnace. hammer scale). The contact of these oxides, and of the O from the flame gases, is aided by stirring up the slag with rabbles. As the pure iron solidifies it is broken up from time to time and turned over, etc., so as to expose the whole charge to the refining action. As in the last-mentioned proc- ess, the metal is finally freed from slag by hammering, and is thereby welded together into a conglomerate of Fe crystals. The reverberatory furnaces consist of beds or hearths sup- ported by iron plates, with a free space underneath so that they are exposed to the cooling action of the atmosphere. The side walls are hollow iron plates, which can be cooled with water, and are lined with a basic slag which has a IRON 201 high melting point, and is rich in FeO. The older, single- hearth reverberatory furnaces have a large grate area and the waste gases which pass off at a temperature of 1200 to 1300 C. are used, as a rule, for making steam. The newer furnaces have regenerative gas firing and double hearths. According to the requirements placed upon the iron, 10 to 20 charges of 500 to 700 Ibs. each are worked up in a day in the older furnaces; in the newer ones, 24 charges of 1000 Ibs. each. Loss of metal, 6 to 15%. For heating the furnaces and for power sufficient to hammer FIG. 183. Plan of Puddling Furnace with Flat Grate, burning Anthracite Coal. or roll the blooms, the coal consumption equals the weight of the finished product. The forgeable iron produced by either of the last two processes is called wrought iron. Since the temperature of the furnace hardly reaches 1300, it is evident that the crystals of iron which are formed during the refining cannot fuse together but become welded. Oxidizing Fusion of Cast Iron Above the Melting-point of Pure Iron. Production of Steel. The oxidation is accom- plished by atmospheric oxygen, as well as by the oxygen of iron oxides, by blowing in converters; or by smelting in regener- ative furnaces, or in electric furnaces. The blowing consists in conducting a blast of air through the 202 METALLURGY molten cast iron until the impurities have been burned away; the heat of combustion serves to maintain a temperature so high that the purified iron remains liquid. The substances present have the following heats of combustion per pound: 81=14,090 B.T.U.; P= 10,740 B.T.U.; = 4297 B.T.U.; Mn = 3io3 B.T.U., and Fe = 2435 B.T.U. The rise in temperature caused by these elements is in the following order: By the combustion of i% of the entire weight of the charge, Si increases the temperature, according to recent calculations of Wiist, 287 C. ; the same amount of P, 185 C. ; of Mn, 61 C.; of Fe, 44 C.; and of C., 8.8 C. (CO escapes as a gas and takes along with it the greater part of the heat of combustion). Therefore, for purification by this process, an iron rich in Si or in P is desirable. An iron containing large amounts of both these elements could not be refined satisfactorily in a converter, because a cast iron rich in Si yields an acid slag and necessitates an acid lining in the melting apparatus. The SiO 2 formed by the combus- tion, and that present in the converter lining, will tend to prevent the slagging off of P2O 5 , for it would set P 2 C>5 free from any Fe 3 (PO 4 ) 2 , in accordance with the equation: Fe 3 (P0 4 ) 2 + 3Si0 2 = 3 FeSiO 3 + P 2 O 5 . Free P 2 Oo, however, is readily reduced by the large excess of hot iron present and the P will thus pass into the iron again in the form of phosphide. The process, therefore, is conducted according to one of two principles, the older one being named, after its discoverer, the Bessemer Process. In this case pig irons high in Si (up to 2% Si) are " blown " in tilting, pear-shaped vessels called Bessemer converters. These are made of wrought-iron plates bolted firmly together and are lined with an infusible silicious rock termed ganister. They are fitted with tuyere boxes at the bottom. In the Basic Bessemer or Thomas-Gilchrist Process, a cast iron rich in P is used (up to 2.5% P) with as low Si as possible. The IRON 203 converters, in this case, are given a basic lining consisting of calcined dolomite with the addition of lime. The slags formed consist of a basic calcium phosphate, with P 2 O5 up to 25%; this forms a valuable by-product, the so-called Thomas slag, which is used as a fertilizer. The converters are built in different sizes, up to 10 ft. in height and 10 ft. in width (measuring the shell without lining). They will hold from 1.5 to 25 tons of pig iron, which is blown in from 15 to 20 minutes. The lost metal amounts to iotoi5%. As adjunct to the Bessemer plant, there remain to be mentioned the PiG-lRON MIXER, which is a revolving drum on a horizontal axis; it is made of iron plates with a silicious lining, and has a capacity up to 600 tons. By this mixing, a fairly uniform material is guaranteed for the converter and a certain preliminary purification takes place, inasmuch as Mn combines with S and the MnS separates out as slag. Open Hearth Process. This process was rendered possible by the discovery of regenerative firing by Friedrich and Wilhelm Siemens and first utilized by Emil and Pierce Martin in a Siemens furnace for the production of steel (1865). In the Siemens-Martin Process the oxidation of the impurities in the iron is effected partly by the oxygen of the air and partly by the oxides added, which is either rust on scrap iron, or in the form of hammer-scale, or pure ore (magnetite, hematite). According to whether the raw material is rich in Si or in P, the work is either carried on " acid " or " basic," as in the Bessemer process, and the lining of the hearth is governed in the same way. While the Siemens-Martin process was formerly used almost exclusively for working up pig iron or wrought iron into steel, it has since been used with much success for smelting pig iron with iron ores; or, in other words, a method of working has developed in which considerable amounts of steel are obtained directly from ores. Such processes are the recent ones of Bertrand-Thiel and of Talbot. In 204 METALLURGY the former, the work is carried out with two Siemens furnaces with basic hearths in such a way that in one furnace a preliminary refining takes place with the production of a slag FIG. 1840. Converter. rich in P 2 O 5 , and in the second the finishing work is carried out with a fresh addition of ore. The Talbot process works with one large furnace (up to 250 tons charge) in which the IRON 205 liquid pig iron is run upon the previously-introduced ore and lime. After the conclusion of the violent reaction that takes place, the first slag is removed in order to smelt the metal further with another addition of ore. FIG. 1846. Converter. APPARATUS: Reverberatory furnaces with open hearths and with regenerative firing. The hearth rests upon iron plates and consists either of lump quartz or of calcined dolomite. The other walls of the smelting space are also of acid or basic material, at least on the inside. The side walls contain apertures for working, charging, and tapping; the end walls have openings for admitting the 206 METALLURGY gas and air which are alternately introduced, from one side and then the other, after passing through the heating FIG. 185. Siemens Open Hearth Furnace. FIG. 186. Siemens Open Hearth Furnace. chambers lying beneath the hearth. Each furnace, therefore, has two pairs of heating chambers which consist of shaft- like structures latticed with firebricks to aid in the trans- IRON 207 ference of heat. One chamber of each pair serves for heating the combustible gases, and the other for heating the air required in the combustion of these gases. During the operation of the furnace, therefore, the current of gas is sent into one chamber from the bottom, and the air is sent into the other chamber. Both of these currents pass through the narrow openings in the brickwork and are heated by contact with the hot bricks. The two streams are kept separate until after they leave the top of the chambers, when they are introduced into the smelting space and combustion at once takes place. Since a high temperature is required in the smelting zone, the heat transference cannot be carried very far here. Hence, the hot waste gases, as they leave this zone, are led through channels in the opposite wall, into the second pair of heating chambers in which, as they pass down- ward and out at the bottom, they heat the bricks and walls. After an interval of 30 to 50 minutes the direc- tion of the gas currents is reversed and the air and com- bustible gases are now conducted through the chambers that have just been heated by the escaping gases. The furnaces are sometimes made so that they will tilt and are then suited for very large charges. The ordinary furnaces with fixed hearths have a capacity of from 12 to 40 tons, but the tilting furnaces will hold up to 300 tons. From three to six charges are worked through in 24 hours. Metal losses, 6 to 8%. Bessemer and Open Hearth Treatment in Succession (Duplex Process) has been practiced in some places (e.g., in Wittko- witz) for working up a pig iron containing too much P to be satisfactorily purified by the ordinary Bessemer process, and too much Si for the basic Bessemer process. Electric Smelting, like the Siemens-Martin process, can be used for working up either scrap or pig iron. In 1878, soon after the discovery of the dynamo, Wilhelm Siemens proposed to smelt iron by electricity, and since that time there has been 208 METALLURGY no end ot experiments toward solving the problem, but up to 1900 these were mostly concerned with the question of electrodes. In heating by the electric arc, carbon electrodes are indispensable. When, however, the electric arc springs from a carbon electrode to the mass to be heated, it is un- avoidable, if this mass consists of a metal like iron, that some carbon will be taken up by it. The taking up of carbon, in spite of the use of two carbon electrodes, was first prevented satisfactorily by means of the FIG. 187. Heroult Furnace. HEROULT FURNACE, consisting of an electric arc and resist- ance furnace in which both electrodes are introduced from above into the smelting hearth. An oxidizing slag (mag- netite and a basic flux) is kept upon the metal. The electrodes are placed far enough apart, and the layer of slag on the other hand is kept so thin that the current passes from one electrode, with the formation of an arc, first to the slag, then to the metal, from the other end of the metal again to the slag, and thence to the other elec- IRON 209 trode with the formation of a second arc. Evidently, then, as sources of heat there are: two electric arcs playing directly upon the surface of the slag, two layers of slag resting directly upon the metal, and the metal itself as heating resistance. The slag, containing the substances capable of entering into chemical reaction, is heated very hot and so is the metal itself, which is well protected from loss of heat. The C from the electrodes is oxidized in the slag layer. A glance at the whole arrangement shows, and this has been confirmed in practice, that the oxidizing agent in the refining slag is brought by the intense heat to its maximum reaction velocity in respect to the impurities present in the bath of iron. As a matter of fact, it is possible to accomplish in this furnace, and in that of the following ones, a much more efficient purification of the iron than in any of the previously described furnaces. The potential required for the Heroult furnace is no to 120 volts. GIROD FURNACE. This arc-resistance furnace was designed by Paul Girod-Ugine and, as compared with the Heroult furnace, excels in simplicity of construction and operation. One pole of the electric arc, consisting of one or more blocks of carbon, is introduced from above into the middle of the smelting hearth. The slag floating upon the metal forms the other pole. It obtains its contact through the metal bath, which is placed in circuit with iron rods, or rings, that are introduced from below into the furnace near the periphery of the metal bath. These connecting pieces, which are cooled somewhat outside the furnace, are so dimensioned that they can just conduct the current without too great resistance; in other words, they are given as much load as possible so that there is a good distribution of the current all over the bath. The ends, drawn out somewhat where they are in contact with the bath, are kept just hot enough by the current to prevent any undesired conducting away of heat from the bath. As sources of 210 METALLURGY FIG. 188. Girod Furnace. IRON 211 heat in the Girod furnace, therefore, there are: one or more electric arcs over the middle of the bath, a slag layer upon the metal, and the metal itself. The current passes with the formation of an arc from one or several upper electrodes through the refining slag to the middle of the metal, whence it radiates uniformly in all directions, heats the bath, and keeps it in motion, and is carried away in the outer periphery. For the Girod furnace, a potential of 55 to 65 volts is necessary. INDUCTION FURNACES, a discovery of de Ferrantis in the year 1887, were first adapted by Kjellin, and almost at the same time by Colby (1900), for use as furnaces for iron smelting. A closed electro-magnet, built in a right angle, forms a transformer with the fused, or ready- to-be- fused, iron which rests in a ring-shaped gutter that is placed perpendicularly to the iron core of the magnet. The most effective construction has proved to be that in which the winding of the electro-magnet lies concentric to the iron-bath, whether this be within or without the field of the primary circuit. The secondary current produced in the iron of the gutter by an alternating current in the primary circuit is at once transformed into heat and the iron is quickly melted. Since the furnace works without electrodes, there is no danger of contamination from electrode material, but notwithstanding this, the majority of these induction furnaces have not proved satisfactory. The oxidizing slag (iron oxides, etc.) owing to its low conductivity, takes but little part in the formation of the secondary current and in the resulting heat transference, but obtains its heat from the iron. The slag, therefore, is considerably colder than that of the Heroult and Girod furnaces. In those furnaces, the slag acts as one pole of the arc and, because of its poor conductivity, causes the production of much heat in the passage of electricity through it, and this high temperature is particularly favorable for increasing the velocity of the reactions that 212 METALLURGY FIG. 189. Rochling-Rodenhauser Furnace. n FIG. 190. Rochling-Rodenhauser Furnace. IRON 213 tend to take place between the oxygen of the slag and the impurities of the iron. This difficulty is overcome in the furnace of ROCHLING and RODENHAUSER by giving direct resistance heating to a part of the bath. Both vertical arms of the magnet's iron core H (Fig. 191), are provided with windings, A, of the primary circuit, and with melting gutters C; these last unite in a broad middle hearth, D. At the ends of this middle hearth are two protective walls FIG. 191. Rochling-Rodenhauser Furnace. of magnesia or dolomite, both of which materials conduct well at high temperatures. Back of these walls and con- nected with a secondary current J5, are pole plates E from which currents are said to pass through the charge con- tained in the broad hearth. CAPACITY OF ELECTRIC FURNACES. Whereas the Heroult and Girod furnaces work advantageously even with cold material, the induction furnaces are practically restricted to work with molten and partly purified iron. Starting 214 METALLURGY with cold metal as raw material, in a small Heroult or Girod furnace (2 to 2.5 tons charge) about 900 kilowatt hours = 1230 horse-power hours are to be reckoned per ton of finished steel; in large furnaces (8 to 10 tons), 700 kilo- watt hours = 95 1 horse-power hours. In the first case, there- fore, about 0.14, and hi the second 0.108 horse-power year per ton of steel. If the furnace can be charged with liquid metal, then the expenditure of energy is diminished by at least one-third. The Girod method of introducing the electric current gives greater velocity and uniformity when working with cold metal than does the Heroult method. The consumption of power by the induction furnace is less only when the iron is already melted and partly purified; when cold metal is used, more power is required than with the Heroult and Girod furnaces. In all processes for the production of steel, it is possible to remove all the impurities only by an over-oxidation, so to speak, i.e., as the concentration of the impurities diminishes, the more Fe will be oxidized to FeO, which partly passes into the slag (metal loss) and is partly held in solution by the iron. If to this iron, containing FeO, the various substances C, Si, etc., were added which are necessary for the production of special kinds of steel, then, particularly with C, chemical reaction would take place during the solidification, and in the case of C this reaction would be accompanied by the evolu- tion of a gas (CO) and the result would be a porous metal. FeO, remaining dissolved in the iron, makes the metal red-short; (i.e., brittle when hot). After the fusion is completed, there- fore, there is in all cases a Reduction by the addition of Mn in the form of a Mn-Fe alloy, of C for the recarburization, of Si in the form of silicides, and finally of Al for the removal of the last traces of FeO and for accelerating the reactions between FeO and the other substances charged. The addition of the Al takes place usually in the ladles, or even after pouring into the molds, IRON 215 The Finishing of Forgeable Irons. In the refining methods which have been described, it is practically impossible to stop the process when the metal contains the desired amount of C or Si corresponding to a certain quality of iron. The impurities, are, therefore, removed completely and then substances are added which will impart to the iron the desired properties. The simplest and quickest way to do this, is by Alloying in the Bath, which is accomplished in the pro- duction of steel by adding the necessary substances them- selves to the bath of pure Fe, or in the form of an alloy or chemical compound, either in the smelting furnace or in the ladle. For the recarburization of iron up to a certain definite point, it is sufficient to add pure coke or graphite to the ladle as the iron is poured from the converter (Darby- Phoenix Process). It is more usual, however, to add spiegel- eisen (see Mn) with high C content. Similarly, for introducing Si, a ferro-silicon is added. Also for introducing other metals, such as Mn, Cr, W, etc., it is customary to add these in the form of an iron alloy on account of its being easier to prepare such an alloy of sufficient purity than the pure metal itself; W and Cr as vrell as Ni and other metals, are also added in the pure state. For producing varieties of steel of very uniform compo- sition, particularly when it is not necessary to keep the C content low, the pure, analyzed metal is melted with the required additions (Ni, W, Cr, etc.) in graphite crucibles. In this way crucible steel is prepared, and recently the melting has been done in an electric furnace. By Welding, accomplished by pressing together (hammering or rolling) rods of different kinds of iron at a welding heat, it is possible to prepare different qualities of iron and steel which apparently are uniform but when examined under the microscope show that different materials have been united mechanically. Particularly for the so-called dou- ble shear steel (for cutlery, etc.) it is important to have hard particles embedded in a softer and sufficiently elastic 216 METALLURGY matrix; in this way the sharpness of instruments is main- tained. Cementation. This process, like that of malleableizing, retains the form of the objects subjected to it, but the nature of the process is exactly the reverse. The objects are embedded in charcoal powder and heated for a long time to a tempera- ture of about 900 C. ; in this way the C content can be raised to about 1.2%. The working conditions are favorable for the reversal of the reaction that takes place during mal- leableizing; 2CO = C + CO 2 . The CO 2 , as fast as it is formed, is reduced by the bed of C to CO again and can serve anew for the introduction of C into the iron. It is also possible to carburize iron objects superficially by dip- ping them in baths or exposing them to the action of certain gases. These compounds, or gases, on being heated, give rise to C or carbides. The process, called case hardening, is accomplished by heating the metal and plunging it into some substance like yellow prussiate of potash (potassium ferrocyanide) K 4 Fe(CN)e = 4KCN + FeC 2 + N2, or by heat- ing in a current of illuminating gas. Properties of Iron : SPECIFIC GRAVITY: 7.86. COLOR: grayish- white with high luster. MECHANICAL PROPERTIES: tough, very ductile. STRUCTURE: see ferrite, page 192, Fig. 177. MELTING-POINT: 1512 C. (2754 F.) BOILING-POINT: 2600 C. (?) (4712 F.) ELECTRICAL CONDUCTIVITY: about 0.14 that of Ag. MAGNETIC PROPERTIES: Fe is the most paramagnetic of metais. ALLOYS readily with most of the earth metals, slightly with Pb and Cu. In the presence of Si, Fe will take up more Cu; or, in other words, it dissolves copper silicide more readily than it does pure Cu. Fe also alloys easily with its own compounds with metalloids (C, Si, P, S, O) (Cf. pig-iron, page 188). IRON 217 CHEMICAL BEHAVIOR: in dry air it is fairly stable toward O at low temperatures. At 300 C. it is more readily oxi- dizable. At a moderate red heat it combines with S and P, and at higher temperatures with C and Si. The halogens act energetically upon Fe even at ordinary temperatures. On account of its high electrolytic solution tension (+0.344 toward H) it dissolves easily in dilute mineral acids with evolution of H, and in HNOs with reduction of the acid and evolution of nitric oxide. In concentrated H 2 SO4 or HNOs, it is so difficultly soluble that iron ves- sels can be used for transporting these acids. In the tech- nically-important iron compounds, the Fe is either bivalent (ferrous compounds) or trivalent (ferric compounds). Only in the carbides and silicides, which are formed and are stable at high temperatures, does Fe show a different valence. The most important carbide has the compo- sition Fe 3 C (cementite), and silicides are known corre- sponding to the formulas Fe 2 Si, FeSi, Fe 3 Si 2 , and FeSi 2 . The properties of the technically-important varieties of iron have already been discussed sufficiently in the sections denning the terms cast iron (white and gray), wrought iron, and steel. CHROMIUM Sources Natural Sources : CHROME-IRON ORE, or chromite, FeCr 2 O4 usually accompanied by serpentine. This is the most important chromium ore. Of the other minerals, the best-known is CROCOITE, lead chromate, PbCrO 4 , now scarcely used at all for the production of metallic chromium. Ferro-Chrome A Reducing Smelt of the chrome -iron ore gives directly a com- mercial Fe-Cr alloy. As flux, only charcoal or coke powder is required. At the beginning of the reaction, which takes place with evolution of CO, there is danger of the mixed ore and carbon falling apart on account of the difference in specific gravities; to prevent this, a little colophonium or pitch is added to the charge. On being heated, the pitch glues together the ore and carbon, forming a coke at a high temperature, and then finally the whole mass unites to one large lump. The charge is regulated, of course, by the nature of the ore. To one ton of ore, from 250 to 330 Ibs. of charcoal powder and 125 to 155 Ibs. of powdered colophonium or pitch is added. APPARATUS: Crucibles in reverberatory furnaces, using forced draft or regenerative firing, or in electric furnaces. THE CRUCIBLES are made of graphite or of clay. When many of the former material are required, the cost of operating is materially increased. Clay crucibles can be used but once, 218 CHROMIUM 219 but they are cheap. In the simple furnaces the con- sumption of fuel is very great and it is hard to reach the necessary temperature. In this respect the regenerative furnaces are more advantageous. Since in using clay crucibles the heating chambers must ^be cooled enough so that there is no danger of breaking the crucibles in recharg- ing the furnace, the author usually combines two furnaces FIG. 192. Section A, B, C, D, E, F. FlG I93 . Section G, H, K, L 5 M. FIG. 194. Section N, O, P, R. FIGS. 192-4. Crucible Furnace with Regenerative Chambers. in such a way that it is possible to work alternately, the furnace with the fresh charge receiving its heat from the slowly-cooling furnace with the finished charge. The furnace block contains the gas producer and the two regenerative, reverberatory furnaces. (Figs. 192 to 194.) The producer is a simple shaft furnace with horizontal grate for coke. It sends its gas into one of the main flues of the adjacent reverberatory furnace. From the main flue, the gas can be 220 METALLURGY led into one or the other branching flues of each furnace by uniting two faucet tubes with a n -shaped tube of sheet metal. The regenerators are there only for preheating the air. The work is earned out as usual, changing the direction of the gas every half hour or hour. The change in direction of the producer gas is effected by the above-mentioned faucet tubes, of the air and waste-gases by valves. Pure Chromium To prepare Cr comparatively pure and free from Fe, I. The Iron Must be Separated Chemically from chromium. This is accomplished by the following operations: 1. Oxidizing Roast with Alkaline Fluxes. Purpose: changing FeO to Fe 2 Os, Cr 2 O3 into CrOs and union of the latter with CaO or Na 2 O. 2FeCr 2 O 4 + 7 O + 4Na 2 CO 3 = Fe 2 O 3 + 4Na 2 CrO 4 + 4CO 2 . Contrary to the views prevailing in chemical and metallurgical literature, the mass should not melt, for when melting takes place the surface of attack for the O upon the FeCr 2 O 4 is greatly lessened. The firing, therefore, does not need to give a very high temperature. Simple reverberatory furnaces (Figs. 195 to 197) such as are used in the LeBlanc process for the manufacture of soda, are suitable. To prevent the charge fusing together, CaCOs has been added or used to replace a part of the Na 2 CC>3. 2. Lixiviation. Purpose: separation of soluble chromate, Na 2 CrO 4 , from insoluble Fe 2 Os and gangue. As sol- vent, hot water is used and if CaCrO 4 is present, Na 2 CO 3 or Na 2 SO 4 is added. In this case, the work is performed in closed iron drums at 120 to 130 C. The filtered Na 2 CrO 4 solution is evaporated to a concentration of 1.5 sp.gr. 3. Transformation of Na 2 CrO 4 into Na 2 Cr 2 7 . If H 2 SO 4 is added to the concentrated, aqueous solution of the chromate the following reaction takes place: 2Na 2 CrO 4 + H 2 SO 4 = Na 2 Cr 2 O 7 + H 2 O + Na 2 SO 4 . CHROMIUM 221 The Na 2 SO 4 is deposited, to a large extent, as a crys- talline powder. The liquor is drawn off from the sediment, and the solution concentrated, whereby practically all of the remaining Na 2 SO 4 separates out. If care was taken to FIG. 195. p n n n a o Q 3 FIG. 197. FIGS. 195-7. Roasting Furnace. permit from i to 2% of neutral chromate to remain in the solution, the evaporation may take place in iron vessels. After the dehydration is almost complete, the melted is poured into flat pans in which it solidifies. 222 METALLURGY 4. Reduction of Na 2 Cr 2 7 with S is accomplished by melt- ing the mixture in cast-iron kettles, which are set in masonry over a grate. Na 2 Cr 2 O 7 + S=Na 2 SO4 + Cr 2 O 3 . The melt is ladled out, broken up after it has become cold, and leached with water; the Na 2 SO 4 passes into solution, while Cr 2 O 3 remains undissolved and is separated by decanting and filtering off the solution. II. Reduction of the Cr 2 3 . By a Reducing Roast. If it is not necessary for the Cr to be obtained in a fused condition, it suffices to mix the Cr 2 O 3 with wood charcoal, or powdered coke, and heat the mixture in crucibles. Even in the regenerative gas furnace, the reduced metal is not fused, but is to be found at the bottom of the crucible in the form of a powder. By a Reducing Fusion with C. This can be accomplished only in electric furnaces. It is advantageous to add a little Al 2 O 3 to the mixture of Cr 2 Os + 3C and to use as flux some CaF 2 or some AlF 3 .3NaF. The A1 2 O 3 is reduced to metal with the Cr at the temperature required for the melting, but it then acts as reducing agent upon some of the remaining Cr 2 O 3 . APPARATUS: Heroult or Girod furnaces. See Iron, pp. 208, 210. By Reducing Fusion with Al. A mixture of Cr 2 O 3 +Al 2 is kindled by an ignition powder of 3BaO 2 + A1 2 ; the mass then continues to fuse with great evolution of heat and there is formed A1 2 O 3 and Cr 2 . This is GOLDSCHMIDT'S THERMITE PROCESS. APPARATUS: crucibles made of MgO or else lined with MgO. It is best to embed the crucibles in" sand. Electrolysis of Chromium Solutions. The fact that chromium can be deposited electrolytically from aqueous solutions of chromous chloride was discovered by Bunsen in 1854. This scientist showed that a relatively pure metal could be pro- duced by using high current densities (at least 70 amperes per square foot of electrode surface) and concentrated CrCl 2 solutions. Both conditions are difficult to maintain, for the CHROMIUM 223 surface of the electrode is constantly increasing by deposition of Cr and the concentration of the solution is constantly diminish- ing. For this reason the author in 1887 to 1900 used, instead of the CrCl 2 solution, a paste of CrF 3 crystals packed in a linen bag and suspended in a vessel of water through which SO? was constantly passed during the electrolysis to effect depolariza- tion. The cathodes consisted of Pt foil, the anodes of C plates. The current density was gradually increased during the elec- trolysis, to compensate for the gain in electrode surface. The Cr grew upon the cathode foil in a crystalline condition. By the more recent studies of G. Glaser, the limits of con- centration and of current density have been established for CrCb solutions. ELECTROLYTE: CrCl 2 solutions containing 13 to 20 oz. Cr per gallon. ANODES: rods of carbon. CATHODES: Pt foil. CURRENT DENSITY: 85 to 170 amperes per square foot of electrode surface. With densities of 70 amperes per square foot, the metal contains perceptible amounts of CrO, and with 8 amperes per square foot, only CrO is deposited. TEMPERATURE: 50 C. (122 F.), at the most. At higher temperatures the Cr is not deposited in a crystalline con- dition, but in the form of a black powder. Properties of Chromium : SPECIFIC GRAVITY: 6 to 7. COLOR: light gray with high luster. MECHANICAL PROPERTIES : hard and brittle. STRUCTURE: coarsely crystalline. MELTING-POINT: 1515 C. (2760 F.) (?). BOILING-POINT: 2500 C. (4532 F.). ALLOYS readily with Fe, Mn and W; difficultly soluble in most other metals. Some melted products, formerly regarded as alloys, have proved to be mixtures upon more accurate metallographical investigation In these the Cr lies finelv divided in the other metal. 224 METALLURGY CHEMICAL BEHAVIOR: At low temperatures the metal is fairly stable but at high temperatures it combines energet- ically with most of the metalloids, particularly with C and Si. It is more readily dissolved in caustic alkali solutions than in acids. With the former, it forms salts in which the Cr is present in the anion; with the latter, salts in which the Cr exists as a bivalent or trivalent cation. S: TUNGSTEN Sources. Natural Sources : AN OXIDE occurs, called Tungstite, WO 3 . THE SALTS, called tungstates. .__. r These tungstates are more widely Wolframite, FeWO 4 . ,..,., . , , o i- iv ^ iTr^ 1 distributed than the oxide and Scheelite, CaWO 4 . ., ., , , ~. , [ accompany cassitente (cf. Tin). Other Sources: SLAGS obtained in smelting tin ores that contain tungsten. (A) Ferro-Tungsten. In the chapter on Iron, it was mentioned that tungsten is used in the manufacture of a special kind of steel, and that such steels could be prepared by the addition of an iron alloy rich in W (80 to 85% W). This alloy, called ferro-tungsten, is prepared by a Reducing Fusion of wolframite, or scheelite, with powdered quartz and glass; these fluxes slag the bases other than iron that occur in the ores, particularly the alkaline earths. In working with wolframite it is well to remember that the ore often contains considerable amounts of hiibnerite, MnWO 4 . When scheelite is used, it is obvious that Fe must be added to the charge. APPARATUS AND OPERATIONS are the same as with ferro chrome (p. 218). 225 226 METALLURGY (B) Pure Tungsten. Since scheelite, on account of its low Mn content, is better suited for preparing ferro- tungsten than is wolframite, which almost always contains considerable Mn, the latter is used in large quantities in -the preparation of pure tungsten. In working up this ore, or tin slags likewise containing W as tungstate, there are a great many points of resemblance to the process for obtaining pure chromium. I. Separation of W from Fe, Mn, Ca, etc. 1. Oxidizing Roast with Alkaline Fluxes. Purpose: transform- ation of W into a soluble alkali tungstate, Na 2 WO 4 , of Fe and Mn into insoluble oxides, and of Ca into insoluble carbonate. 2FeWO 4 + O + 2Na 2 CO 3 = 2Na 2 WO 4 + Fe 2 O 3 + 2CO 2 . APPARATUS AND OPERATIONS as with Chromium. 2. Lixiviation. Purpose: separation of the soluble tungstate from the substances that are insoluble in water. As solvent, the weak wash waters of the previously roasted product are used. The roasted product, as it comes from the furnace, is leached hot in this liquor. By thus quenching it, the roasted product is broken up where it has been sintered and put into a condition more favorable for leaching. When the solution has a concentration of 10 to 12% tungstate, it is drawn off and concentrated by evaporation. Hereby certain contaminating salts, such as Na 2 SO 4 , which dissolve in water with the tungstate, separate out. When the solution is sufficiently concentrated it is cooled and brought to crystal- lization. 3. Precipitation of the So-called Tungstic Acid, W0 3 . If the concentrated solution of tungstate is heated by introducing steam, and then hydrochloric acid is allowed to run in, or if the powdered crystals of tungstate are shaken into the TUNGSTEN 227 hot solution of hydrochloric acid, the WOs is deposited as a heavy, yellow powder which can be purified by decantation, nitration, and drying. Na 2 WO 4 + 2HC1 = 2NaCl + H 2 O + WO 3 The precipitation and first washing take place in stone vats, the last washing in filter bags. II. Reduction of W0 3 . Reduction without Fusion. Mixtures of about 265 Ibs. WOs, 26 to 33 Ibs. wood charcoal or coke, and 15 to 9 Ibs. of powdered colophonium, or pitch, are placed in crucibles and heated as hot as possible in gas furnaces with blast or regenerative firing. In this case the reduced metal is not melted. APPARATUS AND OPERATIONS as in the reduction of Chromium (p. 222), Reducing Fusion. If the same charge as used for the above reduction is placed in an electric furnace, some difficulty will be experienced in fusing it. The melting point of W lies very high (2800 to 2850 C.) and although a temperature of 3500 C. is reached in the electric arc, this region of high temperature is of very limited area. Hence the electrodes must be brought as close to the metal as possible, or else the latter must be made one pole in the circuit of the arc, and by so doing it is very hard to avoid the taking up of C. Carbon is dissolved with the formation of the carbides W 2 C and WC. Metal containing carbide can be fused readily in the electric arc. Properties of Tungsten : SPECIFIC GRAVITY: 19. COLOR: gray, crystalline as powder; in the fused condition it is a nearly white, lustrous metal. MECHANICAL PROPERTIES : Very hard. MELTING POINT: 2800 to 2850 C. (5070 to 5160 F.). BOILING POINT: 3700 C. (7000 F.). ALLOYS with other metals to about the same degree as Cr does. 228 METALLURGY CHEMICAL BEHAVIOR: It is oxidized by O only at high temperatures; likewise the halogens act vigorously only when the metal is hot. It is insoluble in most acids. Strongly oxidizing acids produce WOs; this oxide is an acid anhydride and combines readily with basic oxides to form tungstates. CADMIUM Sources Natural Sources : THE SULPHIDE, CdS, called greenockite, is of rare occurence and is usually accompanied by ZnS in the blendes of Silesia and of North America. THE CARBONATE, CdCO 3 , is found, as a rule, only with the smithsonites of Silesia and of North America. Other Sources: The zinc dust formed in working up zinc ores containing cadmium. (A) Extraction Since there are not enough cadmium ores available to support an independent industry, the production of cadmium forms merely a side-issue in zinc smelters that work zinc ores con- taining cadmium. As will be seen under the section on Zinc (p. 234), when such ores are roasted CdO is formed together with ZnO and it is the former oxide that is first reduced to metal. The Cd is found, therefore, partly as metal, and partly as carbonate and oxide, together with large amounts of Zn and ZnO (70 to 80%) in the first zinc dust that is obtained. The preparation of Cd from this raw material consists in a repeated Reduction with Fractional Distillation, first in clay and finally in iron retorts. The method of working is the same as for the reduction of ZnO, except that the temperature is kept lower. For details, see Zinc. 229 230 METALLURGY (B) Refining The crude cadmium, containing more or less zinc, can be puri- fied further by repeating the above-mentioned Reduction with fractional distillation until finally the Zn content is brought very low. It can also be freed from Zn by Electrolysis. ANODES : crude Cd containing Zn. CATHODES : pure Cd. ELECTROLYTE: CdCl 2 or CdSO 4 . CURRENT DENSITY: 6 to 15 amperes per square foot. E.M.F.: 2.8 to 3.5 volts. Properties of Cadmium : SPECIFIC GRAVITY: 8.6 to 8.7. COLOR: white, strongly lustrous. MECHANICAL PROPERTIES: soft, ductile. STRUCTURE: It shows a tendency to form dendrites on the surface, otherwise granular. The crystal grains grow considerably upon long-continued heating. ALLOYS with most other metals. It forms some alloys with remarkably low melting-points; e.g., 4 pts. Bi, i pt. Sn, 2 pts. Pb, i pt. Cd = Wood's Metal. 15 pts. Bi, 4 pts. Sn, 8 pts. Pb, 3 pts. Cd = Lipowitz Metal. The former alloy melts at 71 C. (160 F.) and the latter at 60 C. CHEMICAL BEHAVIOR : At low temperatures it is stable in the air but unites readily with the halogens. At high tempera- tures it burns readily in O or S. It is readily soluble in HC1, H 2 SO 4 , and HNO 3 , also in alkali hydroxides. E.M.F. toward H= +0.420 volt. ZINC Sources Natural Sources : ZINC BLENDE, or sphalerite, ZnS. Gangue and accompanying minerals are 1. Galena in Germany in the ore deposits at Ems and the Upper Harz, also in Bohemia and Hungary, and in Australia. 2. Dolomite, limonite, and also smithsonite, often in clay, in Rhineland, Westphalia, Belgium, Upper Silesia, Northern Spain, Algiers, and North America. 3. In gneiss with pyrites in Sweden. SMITHSONITE* or Zinc Spar, ZnCO 3 . Gangue: often with sphalerite and galena, always with zinc silicates in lime- stone, dolomite and also limonite in Rhineland, Belgium, North Spain, England, North America. CALAMINE, H 2 Zn 2 SiO5. WILLEMITE, Zn 2 SiO 4 . The gangue and accompanying miner- als are the same as with smithsonite. RED ZINC ORE or zincite ZnO. Gangue; the same as with franklinite. FRANKLINITE: (Zn, Fe, Mn) (Fe 2 , A1 2 , Mn 2 ) O 3 . It is found in New Jersey near the village of Franklin (whence the name) associated with zincite, willemite, rhodonite, and tephroite, etc. Other Sources : ZINC DUST: the first condensation product in the reduction of zinc with sometimes as much as 90% Zn. * Sometimes called calamine, but incorrectly. 231 232 METALLURGY FLUE DUST from smelting furnaces. FURNACE CALAMINE, the deposit in shaft furnaces in which ores containing Zn are smelted. DROSS, waste metal obtained in the refining of zinciferous metals, and waste from foundries, plating works, etc. ZINC SKIMMINGS, alloys of Ag, Pb, Cu and Zn obtained in the removal of Zn and A from lead bullion. (A) Extraction of Crude Zinc Roast Reduction Work : i. The Roasting effects the transformation of ZnS and other sulphides, as well as of ZnCO 3 and other carbonates, into oxide. Any water held mechanically or chemically is also driven out. The principal reactions that take place are ZnCO 3 = As compared with the other sulphides, sphalerite is diffi- cult to roast; basic sulphates as well as oxide are formed and the former are hard to decompose. Both the oxide and the basic sulphates form a coating over the sulphide and prevent the access of air. A complete roasting, which is necessary on account of the injurious action of sulphides upon the retorts, can be accomplished only with the applica- tion of external heat. ROASTING APPARATUS: HEAPS AND OPEN KILNS for drying, and rarely for calcining, material containing carbonates. SHAFT FURNACES for burning carbonate material. REVERBERATORY FURNACES with fixed hearths, worked by hand or in some cases provided with mechanical stirrers. Recently for roasting blende that contains pyrite, rever- beratory furnaces with revolving hearths have been some- times used. ZINC 233 MUFFLE FURNACES, first introduced by Liebig and Eichhorn by arranging heating flues in the fine pyrites burner of the Maletra-Schaffner System, and then, with success in zinc smelters, by Hasenclever. In furnaces of this type, four men (2 for each 12 hours) can roast from 4 to 4.5 tons of X&&&- ^iwi? FIG. 199. Hasenclever Furnace. ore in a day. From 300 to 450 Ibs. of coal are required per ton of ore. 2. Reduction of the Roasted Product. To carry out the work to the best advantage, the properties of zinc oxide and of zinc necessitate the removal of the metal from the reduction apparatus in the form of vapor and the condensa- 234 METALLURGY tion of the vapor in receivers; in other words a distillation is combined with a reduction. Although the reduction of zinc oxide by carbon will take place at a distinct red heat, still the reaction under these conditions is so slow and incom- plete that to obtain satisfactory results the reduction tem- perature of the zinc oxide must be kept between 1000 and 1300 C.; the melting-point of Zn lies at 415 C. and the boiling-point at 930 to 950 C. FLUX: An abundance of carbon to prevent the formation of CO? during the reduction; this gas will oxidize Zn to ZnO at red-heat in the receivers and results in the forma- tion of zinc dust. Zinc silicates do not require the addition of any flux to combine with the silica, for the ZnO contained in the silicate is reduced and the SiO2 remains behind as such or in the form of acid silicates, and tends to make the residue fusible; the more difficultly-fusible the residue the better for the apparatus in which the work is conducted. BEHAVIOR OF THE CONSTITUENTS OTHER THAN ZnO that are present in the charge: Zinc sulphide is, to be sure, not reducible by carbon, but usually iron compounds are present with which it reacts to form FeS and Zn; FeS has a very injurious action upon the walls of the apparatus. Cadmium compounds behave similarly to the corre- sponding zinc compounds except that CdO is more readily reducible and melts at a lower temperature; it passes over, therefore, with the first of the zinc dust. Of the iron and manganese compounds, Fe 2 O 3 and Mn 2 Oy are reduced to FeO, Fe, and MnO. FeO and MnO unite with silicates to form readily-fusible slags and these are dangerous for the retorts. FeS is unacted upon, but melts and injures the walls. Lead oxide in the free state is easily reduced and is volatilized to some extent with the Zn; combined with silica, it forms fusible silicates which injure the walls. ZINC 235 Free silica is useful, since the melting-point of the residue increases in proportion to the SiC>2 content. Under the different modes of working that prevail in dif- ferent countries and in connection with different mines, the usual apparatus and processes followed may be classified into three groups: the Silesian, the Belgian, and the Rhenish methods. In all cases, muffles, or tube- shaped retorts, closed at one end, are used as reduction vessels. The retorts are either oval, semi-oval, or cylindrical in form. The size and shape of the retort are regulated by the following circumstances. ZINC CONTENT: the greater the zinc content, the smaller. the retort. REDUCTIBILITY : the harder it is to reduce the ore, the smaller the retort. FUEL: the influence of the fuel, now that gas furnaces are used almost universally, steps into the background. Formerly, furnaces with a great many small retorts required a fuel that would give a long flame, and furnaces with large retorts could get along with a short-flame fuel. These factors have led to the development of the following systems of retorts with furnaces suitable for them: In Silesia, where the ores are poor and where to some extent coal giving short flame is available, the work is carried out in large muffles. The Silesian muffles have the fol- lowing inside dimensions: height 26 in., breadth 6 to 8 in.; length 27 to 85 in.; thickness of walls, f in. in front and 2\ in. in back. THE CONDENSERS for Silesian muffles are metal tubes, metal boxes, and also clay tubes, conical or bellied in form. CAPACITY: up to 220 Ibs. ore, together with about 40% as much of reduction carbon. ARRANGEMENT IN THE HEATING CHAMBERS: in a row, as many as 36 pieces side by side; two heating chambers lie with the vertical back walls opposite one another so that one furnace may hold as many as 72 muffles. 236 METALLURGY FIRING: usually regenerative. CONSUMPTION OF FUEL: one Silesian furnace with 60 muffles, each of 220 Ibs. capacity, will work up 6 tons of ore in 24 hours, using thereby 6.1 tons of heating carbon and 2.4 tons of reducing carbon. YIELD: i ton of zinc; hence 6.1 tons of heating carbon and 2.4 tons of reducing carbon are used per ton of zinc. IN BELGIUM where the ores are rich and long-flame coal is available, small retorts are used which are either FIG. 200. Silesian. FIG. 201. Rhenish. FIG. 202. Belgian. Types of Muffles. Scale i : 50. cylindrical tubes 6 to 10 in. in diameter, or oval tubes 6J to 7 in. wide and 7! to n in. high; these retorts are 40 to 57 in. long and the walls are f in. thick in front and i J in. thick in the rear. CONDENSERS FOR THE BELGIAN RETORTS: conical clay tubes 1 6 in. long, and with a diameter of 3 in. in front and 6 in. in back. The wide end is shoved into the retort and the narrow end is fitted into a sheet-iron condenser, or nozzle, which serves to catch the zinc dust. ZINC 237 238 METALLURGY CAPACITY OF THE BELGIAN RETORTS: about 44 Ibs. ore with 40 to 60% as much reducing carbon. ARRANGEMENT OF THE BELGIAN RETORTS IN THE HEATING CHAMBERS: the retorts are placed one over another in six or eight rows; every two or three vertical ZINC 239 series open into a common niche in which the condensers lie. The back end of the tube rests against a supporting wall, and the front end in the wall of the niche; the rest lies free. In one furnace-block, with, as a rule, the back walls of two adjacent heating chambers lying against one another, there are from 56 to 400 of these tubes or retorts. FIRING: Siemens' regenerative or recuperative firing of different systems; direct coal firing is seldom used in modern plants; in the United States natural gas is also ^employed as fuel. FIG. 205. Part of Silesian Furnace. CONSUMPTION OF FUEL: furnaces with 56 or 70 retorts require respectively for i.o and 1.35 to 1.44 tons of ore, 1.8 and 2.5 to 2.55 tons of fuel. YIELD: One ton of zinc produced required 4.5 to 5.5 tons of fuel, and 1.3 to 1.9 tons of reducing carbon. IN ZINC SMELTERS OF THE RHENISH PROVINCES the attempt has been made to combine the advantages of large muffles with the possibility of arranging them as is customary with the smaller retorts. Semi-oval retorts are used, varying in size between the Silesian muffles and Belgian retorts, and so chosen that, with the full burden, they can be placed free in the heating chamber according to the Belgian method. 240 METALLURGY ZINC 241 The dimensions of the retorts in common use in the Rhenish provinces are: breadth up to 6\ in., height up to i if in., length up to 55 in. CONDENSERS FOR THE RHENISH RETORTS: bulging clay tubes open at both ends, of which the wider end is FIG. 207. Belgian Furnace with Siemens, Regenerative Chambers. inserted into the end of the retort and, at the beginning of the work, the front ends are stuck into the sheet iron prolongs, so as to catch the zinc dust. CAPACITY OF THE RHENISH RETORTS: 55 to 75 Ibs. of ore with 40 to 50% as much reducing carbon. ARRANGEMENT OF THE RETORTS IN THE HEATING CHAMBERS: the retorts rest above one another in three 242 METALLURGY horizontal layers; every two vertical columns open into a front niche in which are the receivers. The retorts lie as in the Belgian type. The furnace block may con- tain 40 to 80 retorts in three series, i.e., 120 to 240 retorts in all. FIG. 208. Rhenish Furnace with Recuperative System. FIRING: Siemens' regenerative firing is less used than the so-called recuperative firing. CONSUMPTION OF FUEL: a large furnace of this type with 240 retorts and recuperative firing requires 8.5 to 9.2 tons of fuel and 3.2 tons of reducing carbon in 24 hours for operating with 8 tons of ore with 52 to 54% of zinc. ZINC 243 YIELD: 2.5 tons of fuel and i ton of reducing carbon are required to produce one ton of zinc. FIG. 209. Retort Press (C. Mehler, Aachen). Scale i : 40. THE MAKING OF THE RETORTS is carried out as a rule at the smelter. The work consists of the following operations : BURNING of a part of the clay that is to be used in making the retorts. 244 METALLURGY GRINDING. PULVERIZING the burnt clay. MIXING the powdered clay with 50 to 100% as much fresh clay. STAMPING AND PRESSING the mixture into retorts after it has stood for some time. DRYING AND BURNING THE RETORTS: Enough retorts must be kept in the burning furnace to replace those that are injured during the reduction of the zinc so as to avoid the dangeir of injury that would be likely to take place in the cooling and reheating. The hot retorts, therefore, are introduced from this furnace directly into the retort-furnace as they are required. LIFE OF MUFFLES. The retorts last, according to the amount of work done in the smelter, from 14 to 30 days; this corresponds to a daily replacement of 3 to 7%. It is worth mentioning that many experiments have been carried out in the attempt to utilize electricity for heating in the roast-reduction work. Thus the Cowles Brothers in 1885 proposed the use of retorts which, by having carbon contacts at both ends, were connected with an electric circuit using the charge as the heat-producing resistance. The constant changes taking place in the body of the charge by the reduction process and the fact that as the temperature rises the retort and the walls take part more and more in the conduction of the current, have given rise to difficulties which have not yet been overcome. The Precipitation Process, consisting of the smelting of sulphide ores with iron or fluxes containing iron, has been tested in different directions. In the blast-furnace, the possibility of such a process is altogether out of the question, because the reaction gases can never be kept absolutely free from SO 2 and CO 2 . Thus the Zn would never be obtained as a fused metal. Experiments in which the ZnS was exposed to the action of ZINC 245 molten Fe in a rotary, closed converter, also proved fruitless. The requisite mixing of the ZnS and Fe, which is necessary for the rapid fulfilment of the desired reaction, could not be accomplished by revolving the drum. Apparently the problem has been solved, however, by heating a mixture of finely-divided ZnS with equally fine Fe in an elec- tric furnace in which a bath of FeS is formed b the reaction and this kept as a heating resistance. The FeS holds Fe as well as ZnS in solution, so that it serves to maintain the desired state of subdivision, and the reaction between the Fe and ZnS continues to take place smoothly. The Direct Electrolytic Process for working up crude zinc from ores by making the latter the anode in an aqueous solution of zinc salt, has never met with success. (B) Zinc Refining By Liquating in reverberatory furnaces, in which the bottom of the hearth falls away from the fire-bridge toward the opposite wall, it is customary to purify the crude zinc, obtained from the Roast-Reduction Process, which usually contains Pb. From the fused zinc a specifically heavier alloy of Pb or Fe with Zn separates out at the bottom and is removed from time to time from the deepest part of the hearth. On top of the metal bath float the more infusible impurities which were mechanically enclosed by the crude zinc; these are skimmed off. The zinc itself is usually dipped out with iron ladles, poured into plates, and rolled at a suitable temperature. Distillation is employed with alloys rich in lead, containing precious metals (skimmings rich in Zn from the Zn-desilveriza- tion). APPARATUS: Crucibles in air-furnaces; this is the oldest process for small works. 246 METALLURGY Bottle-shaped distilling vessels in tilting air- furnaces, (Figs. 210 to 212.) Bottle-shaped or tube-shaped distilling apparatus made of fire-clay with graphite lining, heated in reverberatory furnaces. (Figs. 213 and 214.) Electrolytic Separation of Zn alloys has not met with success in practice (e.g., for working up skimmings rich in Zn). Lixiviation and Electrolytic Deposition has been attempted at some places without meeting with commercial success, but a practical process has been worked out at the works of Brunner, Mond & Co., at Norwich, England, using the patents of Hopfner. It consists of the following operations : Chloridizing Roast of Zn ores, or pyritic residues containing Zn, in reverberatory furnaces. Lixiviation with water. Purification of the liquor by CRYSTALLIZING out of the sulphate or decomposing it with CaCl 2 . PRECIPITATION of Fe and Mn by means of Ca(OCl) 2 . PRECIPITATION of the more electro-positive metals with zinc dust. Electrolysis of the Liquor. ANODES: carbon plates. CATHODES : rotating Zn plates. ELECTROLYTE: ZnCl 2 solution with 0.08 to 0.12% free HC1. CURRENT DENSITY: 10 amperes per square foot. E.M.F.: 3.3 to 3.8 volts. THE APPARATUS is very complicated, because the anode plates must be placed in special cells, and each be provided with piping for carrying away the chlorine set free during electrolysis. Although the practical results have never been encouraging, the sulphatizing roast of Zn ores would seem more advantageous ZINC 247 C 5 CD n H 3Hr C2 i i o C ) o 248 METALLURGY as the sulphate liquors can be electrolyzed without the use of diaphragms. Moreover, useful by-products can be produced at the anodes and in this way the E.M.F. may be utilized to the best advantage. Properties of Zinc : SPECIFIC GRAVITY: 6.9 to 7.2. COLOR: bluish- white with metallic luster. MECHANICAL PROPERTIES: very ductile at temperatures between 100 and 150 C., can be rolled into sheets, ham- FIG. 215. Structure of Bar Zinc. mered and pressed; at 200 C. it becomes so brittle that it can be pulverized. STRUCTURE: there is a formation of dend rites on the surface of the metal, but the interior is coarse grained. MELTING-POINT: 419 C. (786 F.) BOILING-POINT: 940 C. (1724 F.) ELECTRICAL CONDUCTIVITY: 0.27 that of Ag. ALLOYS with most of the metals. The alloys with Au, Ag and Pb (zinc desilverization) have been mentioned. ZINC 249 Furthermore there are many alloys in practical use, such as those with Cu (brass), with Cu and Sn (bronze), with Cu and Ni (German silver), and with Sn and Sb (bearing metal). CHEMICAL BEHAVIOR: Zinc is oxidized superficially when exposed to the atmosphere, but the layer of oxide is thin and protects the metal within so that zinc objects made of thin foil will keep for a very long time. It is attacked by the halogens at ordinary temperature and by the other metalloids at higher temperatures. The metal will reduce H 2 O at a moderate red heat. On account of its high electrolytic solution tension (E.'M.F. toward 11=4-0.770 volts), it is easily soluble in HC1, H 2 SO 4 , HNO 3 , and in caustic alkali solutions. With the acids it forms salts in which the metal is present as a bivalent cation; with the alkalies it forms zincates in which the Zn is in the anion. MANGANESE Sources Natural Sources: As OXIDES AND HYDRATED OXIDES in the following ores: Braunite, the normal manganic oxide, Mn 2 Os. Hausmannite, mangano-manganic oxide, Mn 3 O 4 (this can be regarded as Mn 2 MnO 4> manganous manganite). Pyrolusite, black oxide of manganese, manganese peroxide, or dioxide, MnO 2 , the most widely distributed ore of manganese. Manganite, Mn 2 O2(OH)2. SALTS: Rhodocrosite, or manganese spar, MnCO 3 . This is an important ore used in the preparation of Mn-Fe alloys. Rhodonite, MnSiOa- Other Sources: SLAGS: particularly from pig iron mixers and the prepara- tion of alloys rich in Mn. METAL WASTE, ferro -manganese. (A) Manganese Alloys The large amounts of Mn used in the metallurgy of iron, together with the fact that it is difficult to prepare from pure ores by the relatively inexpensive smelting processes a pure metal that is stable in the air, make the most important work of manganese smelting the preparation of 250 MANGANESE 251 Iron-manganese Alloys: Spiegeleisen, a name given to alloys with between 5 to 20% Mn. Ferromanganese, which includes the richer alloys with 30 to 80% Mn. The preparation of either of these alloys takes place by a reducing smelt, such as was described under Iron, of iron ores containing manganese, or of manganese and iron ores together in a blast furnace. The working conditions are changed only in so far as is necessitated by the fact that as the Mn content of the required alloy increases, the temperature of the furnace and the content of bases in the slag (particularly of MnO) must be raised. The latter contains from 8 to 30% MnO. High slag losses of the Mn are unavoidable, because if a poorer slag were formed, the primarily reduced Mn would react, as it always does, with the oxides of iron in the ore. Furthermore, large losses of Mn occur through volatilization. (B) Manganese Compounds Among the manganese compounds which come into consider- ation for metallurgical purposes, or which can be prepared in smelters, the only ones to be mentioned are, Manganese Silicides, or alloys of manganese silicide. One of these silicides has the composition Mn 2 Si and contains approximately 20% Mn. There is only one commercial method for preparing silicides, and this consists of a reducing smelt of manganese oxides with sand in an electric furnace. Since, however, most Mn ores contain enough Fe to make the melted silicide unsuitable for some purposes, such ores are preferably smelted first for a ferromanganese and a slag free from Fe which, with the addition* of sand, will yield a silicide practically free from Fe with from 20 to 22% Mn. 252 METALLURGY (C) Pure Manganese Pure manganese cannot be obtained by smelting the purest manganese oxides with C. Even when the amount of C in the charge is most carefully restricted, the metal obtained will contain Mn 3 C and as this compound is decomposed by water at ordinary temperatures, it is not stable in moist air. A fairly stable metal, low in carbon, but one which is not perfectly pure, can be obtained by the Thermite Process of Goldschmidt. Mn 2 Os is mixed with A1 2 and the reaction started by means of ignition powder (3BaO2 + Al 2 ). The heat of reaction is so great that the reduced Mn is entirely liquefied and the slag of A1 2 O 3 sepa- rates out of itself. Properties of Manganese : SPECIFIC GRAVITY: 7 to 8. COLOR: white, lustrous. MECHANICAL PROPERTIES: very hard and brittle. FRACTURE : smooth, almost vitreous. MELTING-POINT: 1233 .(22 51 F.) VOLATILIZES considerably even at the melting temperature. BOILING-POINT: 2200 .(4000 F.) ALLOYS with Fe, and very readily with Cu and the precious metals. CHEMICAL BEHAVIOR: Mn is one of the metals that enters most readily into chemical reactions. It has a great affinity toward almost all of the metalloids and possesses a fairly high electrolytic solution tension, both in aqueous solutions of acids ( + 1.075 volts toward H), as well as in the melts that come into consideration in metallurgical work (a good desulphurizing agent for pig iron). Mn forms a series of six oxides with corresponding salts. Those oxides with the least O possess basic properties, whereas those containing more O are acid anhydrides. ALUMINIUM Sources Natural Sources : OXIDES AND HYDRATED OXIDES. CORUNDUM, pure A1 2 O 3 . The ruby and sapphire are varieties of corundum. EMERY, impure A1 2 O 3 . BAUXITE, hydrargillite, A1(OH) 3 , together with Fe(OH) 3 . SALTS : CRYOLITE, AlF 3 .3NaF. ALUM, A1 2 (SO 4 ) 3 , crystallized with a sulphate of a univalent , metal. SILICATES, particularly the feldspars and the weathering products of the feldspars, e.g. CaO.Al 2 O 3 .2SiO 2 . Extraction It is impossible to obtain a crude Al and to refine it by the smelting processes such as have been described for other metals, because the metal is too electro-positive. The only ores that come into consideration, therefore, are those which occur in a relatively pure state, or which can be purified without difficulty. Cryolite belongs to the former class and hydrargillite, or bauxite, to the latter. In the present method of preparing aluminium, cryolite is used as solvent for bauxite. I. Purification of Crude A1(OH) 3 . This is accomplished by the following operations: i. Transformation into Na 3 A10 3 . According to the old-fash- ioned way of roasting with soda 2 A,(OH) 3 + 3 Na 2 CO 3 = 2Na 3 AlO 3 + 3 H 2 O + 3 CO 2 , 253 254 METALLURGY or by dissolving in caustic soda, carrying out the reaction in iron kettles under a pressure of 5 to 7 atmospheres. In the last process it is necessary to roast the ore slightly before treating with the caustic soda, so that the FeO will be converted into Fe 2 O 3 . 2. Leaching of the Sodium Aluminate in the old process; in both cases there is filtration, leaving NaaAlOs in solution, and a residue of Fe 2 O 3 and the gangue of the ore. 3. Decomposition of the Aluminate by C0 2 . The Al(OH)s separates out from the solution and is obtained by filtration, washing and drying. II. Electrolysis of A1 2 3 in a Molten Bath. The Heroult proc- ess is based upon the electrolytic decomposition of A1 2 O 3 held in solution by melted AlF 3 .3NaF, using, at the same time, the electric current for fusing the electrolyte and deposited metal. The working conditions as carried out to-day are: ELECTROLYZING VESSELS: wrought iron vats with rectangu- lar cross-section, 3 to 4.5 ft. long, 20 to 30 in. wide. The vats are lined with carbon at the bottom and on the sides with cryolite. Both of these linings are kept cold enough by air, which is made to play about the vessels by artificial draft, so that no O enters into the metal from the bottom, and none of the side lining melts. ELECTROLYTE: A1 2 O 3 dissolved in fused AlF 3 .3NaF. ANODES: Carbon blocks introduced into the vat from above. CATHODES: at the start, the carbon bottom lining of the vessel; later, the metallic Al that separates out there. TEMPERATURE: 750 C. (1382 F.). CURRENT DENSITY: 700 amperes per square foot of cathode surface (the horizontal cross-section of the bath). BATH POTENTIAL: The loss in E.M.F. corresponds on an average to 7.5 volts per furnace. ALUMINIUM 255 CONSUMPTION OF POWER: 1400 electric horse power per ton (2200 Ibs.) of metal produced in 24 hours. PROCEDURE AND REACTIONS DURING THE ELECTROLYSIS, To start the process, the upper electrodes are lowered until they form an electric arc with the carbon bottom, then A1F 3 is charged, and the electric arc is maintained until the amount of fused cryolite is sufficient for the upper electrodes to dip into it, then the addition of cryolite is zzzzzzz^^ I FIG. 216. Electric Furnace of Heroult Type. continued until the melt is 8 to 12 in. deep in the vat. Now, as Al separates out during the electrolysis, it is replaced in the bath by a corresponding amount of A1 2 O 3 , carefully avoiding a large excess of the latter. The metallic Al separates out at the bottom cathode and O is evolved at the anodes which dip into the melt, and this O combines with the C to form CC>2. The Al is removed from time to time by tapping or ladling. 256 METALLURGY Although quite a number of other electrolytic processes have proved suitable for the production of aluminium on a large scale, they have not shown so great economy as the Heroult process. They possess a historical interest, however, and a few will be mentioned as representing the gradual development of the Al industry. BUNSP:N in 1854 deposited the first Al from a fused compound, A1C1 3 . FIG. 217. Dendritic Structure on Lower Side of Cast Ingot. ST. CLAIRE DEVILLE, a little later in the same year, proposed that the deposited Al be replaced in the bath by A1 2 O3, but soon after gave up electrolysis and proposed the decom- position of AlCl 3 .NaCl by Na, the method that was used for a long time in the technical production of aluminium. ROSE, in 1855, recommended the decomposition of cryolite byNa. BEKETOFF, in 1865, the decomposition of cryolite by Mg. GRABAU, in 1887, started with A1 2 (SO 4 ) 3 and discovered the following interesting reactions: A1 2 (SO 4 ) 3 + 2(AlF 3 . 3 NaF) - 4 A1F 3 + 3Na 2 SO 4 ; 4 A1F,; + 3Na 2 = A1 2 + 2 ( AlF^NaF) . ALUMINIUM 257 COWLES BROTHERS, in 1884, carried through the reducing smelting of Al 2 Os with C in the electric furnace but could not obtain pure Al; they worked, therefore, for the produc- tion of Al alloys with Cu and Fe: aluminium bronze and ferro-aluminium. Properties of Aluminium : SPECIFIC GRAVITY: 2.53. COLOR: white with high luster. MECHANICAL PROPERTIES: at 100 to 150 C. (212 to 302 F.), it can be forged and rolled satisfactorily; at 500 C. FIG. 218. Large-grained Inner Surface of a Partly Solidified Block : from which the Still-molten Metal has been Withdrawn. (932 F.) it can be compressed easily; at 530 C. (986 F.) it becomes so friable that it can be pulverized. STRUCTURE: according to the way it is poured, it is either dendritic or grained. MELTING-POINT: 687 C, (1270 F.). ALLOYS with most other metals. In the experiments carried out in the attempt to prepare Al, and as regards technical importance, the alloys with the following metals have been studied: Cu, Ni, Co, Fe, Sb, Sn, Zn, Mg. CHEMICAL BEHAVIOR. At ordinary temperatures Al is very resistant to the constituents of the atmosphere; even at 700 to 800 C. it oxidizes only slowly, but with the develop- 258 METALLURGY ment of a great deal of heat i kg. = 62 74 Cai. (i Ib. = 11,293 B.T.U.). It combines with halogens even at low temperatures and readily with the other metalloids; even with N, at high temperatures. In spite of its high solution tension ( + 1.276 to ward H) it dissolves comparatively quickly only in HC1 and NaOH; in H 2 SO4 it dissolves much less readily, and in HNOs very slowly. Al, however, is a very energetic reducing agent and precipitant for metal oxides and other compounds of metals, both in the solid, molten and dissolved condition. Particularly in the decomposition or oxides rich in O, so much heat is evolved that difficultly- fusible metals, like Cr and the slag of Al 2 Os, are melted or even volatilized (cf. Chromium, page 222). In the com- pounds that are of industrial importance, Al is either present as a trivalent cation (of the type ALOs) or, when combined with strongly basic oxides, in the anion (forming aluminates, of the type INDEX Acids, washing mercury with, 59 Allen furnace, 67 Allis-Chalmers furnace, copper, 76 , lead, 128 Alloying iron in the bath, 215 Aluminium, 253 , Beketoff, 256 , Bunsen, 256 , Cowles Brothers, 257 , Deville, 256 , Grabau, 256 , HeYoult, 254 hydroxide, purification of, 253 oxide, electrolysis of, 257 , properties of, 257 , Rose, 256 salts, 253 Amalgamation, gold, 2 , silver, 36 , , caldron process, 46 , , Caso process, 46 , , Krohnke process, 41 , , Patio process, 41 , , Washoe process, 46 , , with chemicals, 40 , , without chemicals, 36 Amalgamator, Laszlo, 40 , Schemnitz, 40 Amalgam catchers, 38 Anaconda copper converters, 97, 100 Antimonial lead, 154 Antimony, 149 , Borchers' furnaces, 112, 114 , concentration, 149 Antimony, electrolysis 152 , properties, 156 , refining, 155 , sources, 149 , starring, 155 ores, liquation, 149 , lixiviation, 151 , oxidizing roast with sublima- tion, 150 , precipitation methods, 152 , reduction methods, 151 Argall furnace, copper, 70, 71 Arsenic, removal from tin ores, 134 Augustin process, silver, 48 Austen, see Roberts- Austen Austenite, 192 B Barrel amalgamation, silver, 47 Base bullion, 29 Bath, alloying iron in the, 215 Beketoff, 256 Belgian treatment of zinc ores, 236 Benedicks, constitution of iron-car- bon alloys, 192 Bertrand-Thiel process, 203 Bessemer process, 202 Betts, electrolysis of lead, 130 Bismuth, 109 , Borchers' furnaces, 112, 114 , electrolysis, 116 , melting with sulphur, 116 , oxidizing fusion, 115 , precipitation, 115 , properties, 116 259 260 INDEX Bismuth ores, 109 , chemical concentration, 109 , concentration processes, 109 , liquation, 110 , melting in crucibles, 111 , melting in reverberatories, 114 , reduction process, 110 Blast furnaces, iron, 181, 182, 183 , , reactions in, 185 , lead, 128 Blast, hot, 182 Blowing copper matte, Manh&s, 95 hi converters, iron, 201 , nickel, 162 Bottoms, smelting of, nickel, 165 Borchers, continuous kiln, bismuth and antimony, 112 , crucible furnace with regenerative chambers, chromium, 219 , direct electric treatment of nickel matte, 173 , electric smelting of nickel matte, 171 , electric smelting of tin, 138 , electrolytic treatment of copper matte, 106 , reverberatory furnace, bismuth and antimony, 114 , treatment of nickel ores, 168 Bradford, converter for lead ores, 127 Brittany process, lead, 121 Broken Hill, zinc furnace, 247 Brown furnaces, copper, 67, 68 Bruckner furnace, copper, 69, 70 Bunsen, 256 Butters' vacuum filter, gold, 19 Cadmium, 229 ores, 229 , properties of, 230 Caldron process, silver, 46 Capacity of electric furnaces, 213 Carbon-iron alloys, constitution of, 188, 192 , Benedicks, 192 , Charpy, 192 Carbon-iron alloys, Goerens, 188, 192' , Roberts-Austen, 188, 192 , Roozeboom, 192 , Wiist, 188, 192 Carinthian process, lead, 121 Carmichael-Bradford converter, lead,, 126, 127 Case-hardening of steel, 216 Caso process, silver, 46 Cast iron, 180 , gray, 191 , white, 190 Cementation, 216 Cementite, 188, 192 Charcoal hearth process, 199 Charpy, constitution of iron-carbon alloys, 192 Chemical concentration, bismuth, 109 , tin ores, 133 solution and precipitation, gold, 2 , platinum, 26 Chloridizing fusion, silver, 51 roasting, muffles for, 72 Chlorinating solution, gold, 21 Chlorination barrel, gold, 4 - gold, 2 plant, 5, 6 platinum, 26 Chromic oxide, reduction of, 222 Chromite, reducing smelt of, 218 Chromium, 218 ores, 218 , properties of, 223 , separation from iron, 220 solutions, electrolysis of, 222 Colby furnace, iron, 211 Colorado Iron Works, furnaces, 76, 77 Concentrated nickel matte, electric smelting of, 174 Concentrated nickel matte, pure nickel from, 173 Concentration, of antimony ores, 149 , of bismuth ores, 109 , of copper ores, 60, 90 , of iron ores, 178 , of lead ores, 118 , of nickel ores, 160 INDEX 261 Concentration, of tin ores, 133 Converters, Anaconda, 97, 100 , blowing in, nickel, 162 for copper ores, 97, 100 for iron ores, 203, 204, 205 for lead ores, 126, 127 , Carmichael-Bradford, 126, 127 , Huntington-Heberlein, 126 , Savelsberg, 127 , Stalmann, 96, 98, 99 , Sticht, 96, 98, 99 Converting copper matte, 89 Constitution of iron-carbon alloys, 188-193 Copper, 60 , Allen furnace, 67 , Allis-Chalmers furnace, 76 , Anaconda converters, 97, 100 , Argall furnace, 70, 71 , Brown furnace, 67 , Bruckner furnace, 70 , Colorado Iron Works furnaces, 76, 77 concentrating processes, 60 converters, 96-99 enriching processes, 60 fore-hearths, 81 , furnace refining, 102 , hand reverberatories, 66 , heap roasting, 63 , hearth furnaces, 66 , Herreshoff furnace, 66, 67 , Hixon furnace, 67 , Hocking-Oxland furnace, 70 , Holthoff-Wethey furnace, 67 , Humboldt furnace, 67 , Johnson furnace, 80 , Kaufmann furnace, 67 , Keller furnace, 69 , kernel roasting, 87 leaching, concentration by, 90 , lump pyrite furnace, Lunge, 64 , MacDougall furnace, 66, 67 , Maletra-Schaffner furnace, 64 , Mansfeld furnaces, 85 , kiln, 64 , Mathewson furnaces, 83, 87 Copper matte, concentration by smelting, 89 , converting, 89, 95 , electrolytic treatment, 106 , nature of, Rontgen, 89 , oxidizing roast of, 89 , reverberatory furnaces for, 84 , shaft furnaces, 73 , smelting for, 71, 87 , muffle furnaces, 71, 72 , O'Hara furnace, 67 , O'Hara-Brown furnace, 68 , Oker furnace, 75 ores, 60 , oxidizing roast, 61 , Parkes furnace, 66 , Pearce furnace, 67 , Peters furnace, 86 , precipitation of, 99 , properties of refined, 107 , purification of leached liquor, Hofmann, 94 pyrite furnace, 63 , pyritic smelting, 87 , reaction smelting, 95 , reverberatories as fore-hearths, 81 , reverberatory furnaces, 65, 84 , roasting, 65 , smelting, 89 roasting furnaces, 66-73 operations, 61 roast-reduction process, 99 slag and matte spouts, 81 calculation, 73 smelting for matte, 71, 87 operations, 61 , solution of gold in, 1 , of platinum in, 25 , of silver in, 28 , Spence furnace, 69 , stall roasting, 63 , Stalmann converter, 96, 98, 99 , Stetefeldt furnace, 65 , Sticht converter, 96, 98, 99 , furnace, 78, 79 , Walker casting machine, 103 , Wethey, furnace, 69 262 INDEX Copper, White-Howell furnace, 70 Cowles Brothers, aluminium, 257 Cowper stoves for heating blast, 183 Crucible furnace, chromium, 219 Crucibles, melting in, bismuth, 111 Crucible steel, 215 Crude nickel matte or speiss, 162 Crystallization of tin, 144 - processes, 30, 130 Cupellation, English and German methods, 33 Cupelling furnaces, 37, 38 Cyanide leaching, 7 plant for gold ores, 12 solutions, electrolysis of, gold, 18 , precipitation of gold from, 15 D Darby-Phoenix process for recarbur- izing iron, 215 De Ferranti, induction furnace for iron smelting, 211 Distillation of mercury, 59 of zinc, 245 Deville, St. Claire, 256 Direct electrolytic treatment of nickel matte, 173 Dotsch, copper leaching, 94 Douglas, copper leaching, 94 Desulphurizing fluxes, heating with, mercury, 59 Dross, 118 E Electric furnace, aluminium, 255 , iron and steel, 211-213 smelting, nickel matte, 174 , steel, 207 , tin, 138 Electrolysis, aluminium oxide, 254 , antimony, 152, 156 , bismuth, 116 , chromium solutions, 222 , copper, 104 , matte, 106 -, gold, 22 , lead, 130 , nickel, 172 Electrolysis, silver, 53 , tin, 141 , zinc, 246 English method of cupelling silver bullion, 33 roast-reaction process, lead, 121 Extraction, aluminium, 253 , antimony, 151 , bismuth, 110 , cadmium, 229 , copper, 95 , gold, 1 , lead, 119 , mercury, 56 , nickel, 169 , platinum, 25 , silver, 29 , tin, 134 , zinc, 232 Faber du Faur's furnace, zinc, 247 Ferrite, 192 Ferro-chrome, 218 Ferro-manganese, 251 Ferro-molybdenum, 196 Ferro-silicon, 196 Ferro-tungsten, 225 Ferro- vanadium, 196 Filter press, Klein, Schanzlin and Becker, 14 , London and Hamburg Gold Recovery Co., 17 Filter tank for gold, 10 Filtration of mercury, 59 Fire refining of crude silver, 53 Fore-hearths, copper, 81 Franke, electrolytic treatment of copper matte, 106 Freiberg vitriolization process, 51 French of Brittany process, lead, 121 Fluxes, desulphurizing for mercury, 59 Forgeable iron, 197 , finishing of, 215 Furnace gas, iron, 194 Furnaces, Allen, for copper ores, 67 , Allis-Chalmers for copper ores, 76 INDEX 263 Furnaces, Allis-Chalmers for lead ores, 128 , American water-jacket, 77 , Argall for copper ores, 70 , Borchers and Mattonet, tin, 138 , antimony and bismuth, 112, 114 , crucible, chromium, 219 , blast, lead, 127 , , iron, 183 , Brown, copper, 67 , Bruckner, 69, 70 , capacity of electric, 213 , Chinese, for tin ore, 136, 137 , Colby induction, iron, 211 , Colorado Iron Works, copper, 76, 77 , De Ferranti induction, iron, 211 , Faber du Faur, zinc, 247 , Girod ; electric for iron, 209, 210 , Halbhochofen, lead, 127, 128 , hand reverberatories, copper, 66 , hearth reverberatories, 66 , Heroult, aluminium, 255 , , iron, 208 , Herreshoff, copper, 66, 67 , Hixon, copper, 67 , hochofen, lead, 127, 128 , Hocking-Oxland, copper, 70 , Holthoff-Wethey, 67 , Humboldt, copper, 67 , Johnson, copper, 80 , Kaufmann, copper, 67 , Keller, copper, 69 , Kjellin, iron, 211 , Krupp, for amalgam, 46 , Krohnke, for distilling amalgam, '46 , kriimmofen, lead, 127 , lump pyrite, Lunge, 64 , MacDougall, copper, 66, 67 , Maletra-Schaffner, copper, 64 , Mansfeld, copper, 64, 85 , Mathewson, copper, 83, 84, 87 , Mattonet, tin, 138 , muffle, for chloridizing copper ores, 71, 72 Furnaces, muffle, for gold ores, 8 , North American water-jacket, lead, 129 , O'Hara, copper, 67 , O'Hara-Brown, copper, 68 , Oker, copper, 75 , Parkes, copper, 66 , Pearce, copper, 67 - -, gold, 3 , Peters, copper, 86 , reverberatories as fore-hearths, 81 , for bismuth and antimony ores, 114 , for copper matte, 84 , -* for copper ores, 65 , for lead, 125 , for oxidation of impurities in iron, 205 , with moving hearths, 69 , Rochling-Rodenhauser, 212, 213 , Schrubko, zinc, 237, 238 , shaft, copper, 73 , , lead, 127 , Siemens open hearth, iron, 206 , Spence, copper, 69 , Stetefeldt, copper, 65 , Sticht, copper, 78, 79 , Upper Harz, lead, 128 , Wethey, copper, 69 , White-Howell, copper, 70 Fusion, chloridizing, of silver, 51 , desulphurizing of nickel matte, 171 , liquating of zinc, 245 , oxidizing of bismuth, 115 , of tin, 144 , reducing of tin ores, 135 , of wolframite or scheelite, 225 , sulphurizing of antimony, 155 , of bismuth, 116 , of silver, 51 German method of cupelling silver bullion, 33 Girod furnace, iron, 239, 210 264 INDEX Goerens, constitution of iron-carbon alloys, 188, 192 Gold, 1 amalgamation, 2 , American chlorination barrel, 4 , Butters vacuum filter for slimes, 19 , chemical solution and precipita- tion, 2 chlorinating solution, 21 chlorination, 2 barrel, 4 plant, 5, 6 cyanide leaching and precipita- tion, 7 - plant, 12 electrolysis, 18 electrolytic parting, 22 extraction, 1 filter press, 14 plant, 17 - vat, 10 leaching methods, recent, 11 muffle furnace, 8 ore dressing, 1 ores, 1 - parting, 20, 49 , Pearce turret furnace, 3 precipitation from cyanide solu- tions, 15 by zinc, 16 refining, 21 slimes, Butters filter for, 19 solution and precipitation, 21 in metals, 1, 2 , of platinum in, 25 wet-tube mill, Krupp, 13 Goldschmidt, process for tin-plate waste, 141 , thermite process for chromium,222 , for manganese, 252 Grabau, aluminium, 256 Graphite, in iron, 193 Gray cast iron, 191 Giinther, electrolytic treatment of copper matte, 106 , pure nickel from concentrated matte, 173 H Halbhochofen, lead ores, 127, 128 Hand reverberatories, copper, 66 Hard lead, 154 Harz vitriolization process, silver, 53 Heap amalgamation, silver, 40 roasting of copper ores, 63 of lead ores, 124 Hearth furnaces for copper ores, 66 process for lead ores, 121 Hearths, reverberatories with mov- ing,, copper, 69 He"roult, aluminium, 254, 255 furnace for iron and steel, 208 Herrenschmidt's process for nickel ores, 167 Herreshoff furnace for copper ores, 67 Hixon furnace for copper ores, 67 Hocking-Oxland furnace for copper ores, 70 Hofmann, copper leaching, 94 process for treating sliver ores, 49 Holthoff-Wethey furnace for copper ores, 67 Hopfner, lixiviation and electrolytic deposition, zinc, 246 Hot blast, iron, 182 Hunt and Douglas, copper leaching, 94 Huntington-Heberlein converter for lead ores, 126 Iron, 177 alloying in the bath, 215 alloys, 196 annealing steel, 198 , Basic Bessemer process, 202 , Bertrand-Thiel process, 203 , Bessemer process, 202 , blast furnaces, 181 , , reactions in, 185 , capacity of electric furnaces; 213 carbon alloys, constitution of, 188, 192 , case hardening, 216 cementation, 216 INDEX 265 Iron charcoal-hearth process, 199 Colby induction furnace, 211 compounds, 196 converters, 203 , Cowper stoves for blast, 183 , de Ferranti induction furnace, 211 , electric furnaces, 208-213 , smelting, 207 , forgeable 197 , , finishing of, 215 furnace gas, 194 , Girod furnace, 209, 210 , hardening of steel, 197 , HeVoult furnace, 208 , induction furnaces, 211 , kish, 189 , Kjellin induction furnace, 211 , malleableizing, 198 , mixed crystals, 189 , open-hearth process, 203 ores, 177 , concentrating and other pre- liminary operations, 178 , reducing roast, 180 , smelt, 180 , oxidizing fusion of cast, 201 -, pig, 188 products of the blast furnace, 188 , properties, 216 - puddling, 200 recarburization, (Darby-Phoenix) 215 , Rochling - Rodenhauser furnace, 212, 213 , separation from chromium, 220 , from tungsten, 226 , Siemens open-hearth furnace, 206 , Siemens-Martin process, 203 - slag, 193 , Talbot process, 203 , Thomas-Gilchrist process, 202 welding, 215 , Wittkowitz process, 207 Johnson furnace, copper, 80 K Kaufmann furnace, copper, 67 Keller furnace copper, 69 Kernel roasting, 87 Kiln, Mansfeld for copper ores, 64 Kish, 189 Kiss process, 48 Kjellin induction furnace, 211 Klein, Schanzlin and Becker, filter press, 14 Krohnke furnace for distilling amal- gam, 46 process of silver amalgamation, 41 Krummofen, 127 Krupp furnace for distilling amalgam, 46 wet tube mill for gold ores, 13 Laszlo amalgamator, silver, 39 Leached liquor, purification of, cop- per, 94 Leaching, concentration of antimony ores, by, 151 , of copper ores by, 90 . of nickel ores by, 166 , of zinc ores by, 246 Lead, 118 , Allis-Chalmers furnace, 128 , Arent tap furnace, 129 , Brittany process, 121 , Carinthian process, 121 , Carmichael-Bradford converter, 126, 127 , treatment in a converter, 122 , concentration of ores, 118 electrolysis, Betts, 130 , English process, 121 , French process, 121 halbhochofen, 128 , heap roasting, 124 , hearth process, 121 , Huntington-Heberlein converter, 126 , method, 122 , krummofen, 127 266 INDEX Lead, liquation and crystallization processes, 130 matte, 124 , North- American water-jacket fur- nace, 129 ores, 118 , concentration of, 118 , oxidation of, 130 , Pilz-Freiberg furnace, 128 precipitation process, 123 , properties of, 130 reaction smelting, 119 reduction smelting, 123 , reverberatory furnaces, 125 - roasting, 119, 122 roast reduction process, 122 , Savelsberg converter, 127 , Schenck and Rassbach, 119 , shaft furnaces, 127 , sinter roasting, 125 , slag roasting, 125 , solution of gold in, 2 , of platinum in, 25 - , of silver in, 28 , Tarnowitz process, 121 Lehmer's process for treating nickel matte, 171 Liquation of antimony, 149 of bismuth, 110 - of tin, 143 of zinc, 245 London and Hamburg Gold Recovery Co. filter pressing plant, 17 Luce-Rozan process, silver extraction, 30 Lunge lump pyrite furnace, copper, 64 M MacDougall furnace for copper ores, 66, 67 Maletra-Schaffner furnace for copper ores, 64 Malleableizing, 198 Manganese, 250 alloys, 250 compounds, 251 , Goldschmidt thermite process,252 Manganese ores, 250 , properties, 252 , separation from tungsten, 226 silicides, 251 Manhes converter for copper matte,96 Mansfeld furnace for copper ores, 85 kiln for copper matte, 64 Martensite, 192 Mathewson furnaces for copper ores, 83, 84, 87 Matte, copper, concentration by smelting, 89 , , converting, 89, 95 , , furnaces for, 84 , , oxidizing roast of, 89 , , smelting for, 71, 87 , nickel, concentration of, 162 , , converting of, 162 , , formation of, 162, 164 , , oxidizing roast of, 162 spouts, 81 Mattonet, smelting of tin in an electric furnace, 138 Mechanical concentration of tin ores, 133 Mehler, press for making zinc retorts, 243 Melting bismuth ores in crucibles, 111 in reverberatory furnaces, 114 Mercury, 56 distillation, 59 filtration, 59 furnaces, 57, 58 , heating with desulphurizing fluxes, 59 ores, 56 , roasting of, 56 , oxidizing roast, 56 , washing with acids, 59 Metals, solution of gold in, 1 , of platinum in, 25 , of silver in, 28 Mixed crystals in iron-carbon alloys, 189 Mond's process for extracting nickel, 174 INDEX 267 Mortar amalgamation, silver, 39 Muffle furnaces,' for chloridizing roast of copper ores, 72 , for copper ores, 71 , for gold ores, 8 Muffles, types of, for zinc ores, 236 N Nickel, 160 , blowing in converters, 162 , Borchers and Warlimont's proc- ess, 168 , concentration and elimination of copper, 164 , by leaching and precipitation, 166 , of the matte or speiss, 162 , without separation of copper, 161 , desulphurizing fusion of concen- trated matte, 171 electrolysis, 172 extraction, 169 , formation of first matte or speiss, 162 , Herrenschmidt's process, 167 matte, electric smelting of, 174 ores, 160 , concentration methods, 160 , oxidizing roast, 161, 164 , of matte or speiss, 162 , properties, 175 , refining fusion, 172 , roast reaction smelting, 170 , reduction process, 169 , smelting of the bottoms, 165 , of the tops, 165' , to a matte, 164 , solution of platinum in, 25 Nitric acid parting of silver and gold, 49 Non-sinter roasting of lead ores, 126 O'Hara furnace, copper, 67 O'Hara-Brown furnace, copper, 68 Oker process of silver leaching, 48 Oxidation of impurities in lead, 130 - of lead, 130 Oxidizing fusion of bismuth, 115 of tin, 144 roast of antimony ores, 150 of copper ores, 61 of mercury ores, 56 of nickel matte or speiss, 162 of nickel ores, 161, 164 Pan amalgamation, silver, 40, 47 Parkes furnace, copper, 66 process, silver, 30 Parting gold and silver, 20, 49, 51 Patera process, silver, 49 Patio amalgamation process, silver, 40 Pattinson process, silver, 30 Pearce turret, furnace, gold, 3 roasting furnace, copper, 67 Pearlite, 193 Peters, American reverberatory fur- nace, copper, 86 Phoenix recarburization process, iron, 215 Pig iron, 188 - mixer, 203 Pilz-Freiberg furnace for lead ores, 128 Platinum, 25 , chemical solution and precipita- tion, 26 chlorination, 26 ores, 25 , properties, 27 solution in metals, 25 Poor tin slags, smelting of, 141 Precipitation, chemical solution and, of gold, 2 , - ,of platinum, 25, 26 , concentration of nickel by, 166 of copper, 99 -of gold, 21 methods, antimony, 152 , bismuth, 115 , lead, 123 268 INDEX Precipitation methods, manganese, 252 , zinc, 244 Preliminary operations, iron ores, 178 Properties of aluminium, 257 of antimony, 156 of bismuth, 116 of cadmium, 230 of chromium, 223 of copper, 107 of gold, 23 of iron, 216 - of lead, 130 - of manganese, 252 of mercury, 59 - of nickel, 175 of platinum, 27 of silver, 54 - of tin, 145 of tungsten, 227 of zinc, 248 Puddling, 200 Pyrite furnace, Lunge, 64 Pyritic smelting, 87 R Rassbach, reaction smelting for lead, 119 Reactions in the blast furnace, 185 Reaction smelting, copper, 95 , lead, 119 Recarburization of iron, Darby- Phoenix process, 215 Reducing roast of iron ores, 180 smelt of chrome iron ore, 218 of iron ores, 180 of tin ores, 135 of tungsten ores, 225 Reduction methods, antimony ores, 151 for chromic oxide, 222 for iron, 214 for tungstic acid, 227 process, bismuth, 110 Refining antimony, 155 base bullion, 29 bismuth, 115 Refining cadmium, 230 copper, 101 - gold, 21 lead, 129 manganese, 252 mercury, 59 - nickel, 171 platinum, 26 silver, 53 - tin, 143 - tungsten, 226 zinc, 245 Regenerative firing, chromium, 219 , steel, 203 Reverberatory furnaces, as fore- hearths, copper, 81 for bismuth ores, 114 for copper smelting, 89 for copper matte smelting, 84 for copper ores, roasting of, 65 for lead ores, 125 for tin ores, 136 with moving hearths, 69, 126 Rhenish zinc smelting, 239 Rich tin slags, smelting of, 140 Rio Tin to leaching plant, copper ores, 93 Roasting antimony ores, 150 copper ores, 61 furnaces for copper ores, Allen, 67 , Argall, 70, 71 , Brown, 67 , Bruckner, 70 , Herreshoff, 66, 67 , Hixon, 67 , Hocking-Oxland, 70 , Holthoff-Wethey, 67 , Humboldt, 67 , Kaufmann, 67 , Keller, 69 , MacDougall, 66, 67 , O'Hara, 67 , O'Hara-Brown, 68 , Parkes, 66 , Pearce, 67 , Spence, 69 , Wethey, 69 INDEX 269 Roasting furnaces for copper ores, White-Howell, 70 , for gold ores, Pearce, 3 - lead ores, 119, 125 , processes, Brittany, 121 , , Carinthian, 121 , , English, 121 , , Hearth, 121 , , Tarnowitz, 121 - iron ores, 178, 180 mercury ores, 56 with desulphurizing fluxes, 59 muffles for chloridizing burnt pyrite, 72 nickel ores, 161, 164 matte or speiss, 162 Roast-reaction smelting, nickel, 170 Roast-reduction process, copper, 99 , lead, 122 , nickel, 169 , zinc, 232 Roberts-Austen, constitution of iron- carbon alloys, 188, 192 Rochling-Rodenhauser electric fur- nace for iron smelting, 212, 213 Rontgen, nature of copper matte, 89 Roozeboom, constitution of iron- carbon alloys, 192 Rose, aluminium, 256 Rossler's sulphurizing, fusion, silver, 51 Russell process for leaching silver ores, 49 S Savelsberg converter for lead ores, 127 Scheelite, reducing fusion of, 225 Schenck, reaction smelting for lead, 119 Schrubko, zinc furnace, 237, 238 Separation, electrolytic, of zinc, 246 , iron from chromium, 220 , tungsten from iron, manganese, and calcium, 226 Shaft furnaces for copper and lead, 66, 73, 127 , Allis-Chalmers, 76, 128 Shaft furnaces, American water- jacket, 77, 80, 129 , Colorado Iron Works, 76, 77 , early type, 75 , Johnson, 80 , Mathewson, 83 , Oker, 75 , Pilz, 128 , Sticht, 78, 79 , Upper Harz, 128 Siemens open-hearth furnace, iron, 206 Siemens-Martin process, 203 Silesian method of zinc extraction, 235 Silver, 28 , amalgamation, 36 , apparatus, 47 , Caso or caldron process, 46 , Krohnke process, 41 , Patio process, 40 with chemicals, 40 without chemicals, 36 amalgam catchers and auxiliary amalgamators, 38 amalgam filter, 45 American amalgamating pan, 45 arrastra, 40 , Augustin process, 48 barrel amalgamation, 47 chloridizing fusion, 51 concentration of ores, 28 extraction, 28 , English cupellation, 33 , cupelling furnace, 38 electrolysis, 51 electrolytic refining, 53 , Freiberg vitriolization process, 51 , German cupellation, 33 , cupelling furnace, 37 , Harz vitriolization process, 53 , Kiss process, 48 , Krohnke furnace for distilling amalgam, 46 , Krupp furnace, for distilling amalgam, 46 , Laszlo amalgamator, 39 270 INDEX Silver, Luce-Rozan process, 30 mortar amalgamation, 39 , nitric acid parting, 49 , Oker-Longmaid-Henderson proc- ess, 48 - ores, 28 pan amalgamation, 40, 47 , Parkes process, 30 , Patera process, 49 , Pattinson process, 30 properties, 54 , refining of the base bullion, 29 , Rossler's sulphurizing fusion, 51 , Russell process, 49 skimmings from base bullion, 29 sluice amalgamation, 38 solution in metals, 28 , solution of gold in, 2 stamp mill amalgamation, 38 , sulphuric acid parting, 51 , vat amalgamation, 47 , Washoe process, 46 , Ziervogel process, 47 Sinter roasting of lead ores, 125 Slag and matte spouts, copper, 81 calculation, copper, 73 in pig iron, 193 roasting of lead ores, 125 Sluice amalgamation, silver, 38 Smelting, electric, for iron, 207 , , of nickel matte, 174 for copper matte, 71, 87 for zinc in the Rhenish provinces, 239 of bottoms, nickel, 165 of chrome iron ore, 218 of low-grade silver ore, 29 of oxidized tin ores, 135 of poor tin slags, 141 of rich tin slags, 140 of tops, nickel, 165 , reaction, for lead, 119 , reducing of iron ores, 180 , reduction, for lead, 123 roast-reaction, for nickel, 170 Solution and precipitation of gold, 21 , chemical, of gold, 2 Solution, chemical, of platinum, 25, 26 Solution, chlorinating, of gold, 21 , electrolytic, of tin, 141 of gold in metals, 1 of silver in chemical solvents, 47 , in metals, 28 of platinum, 25 Sorbite, 193 Spence roasting furnace, 69 Spiegeleisen, 196, 251 Stadtberge works, leaching of copper ores at, 91 Stall roasting of copper ores, 63 Stalmann converter, 96, 98, 99 Stamp-mill amalgamation, 38 Starring antimony, 155 Steel, 197 , alloying in the bath, 215 , case hardening, 216 cementation, 216 , electric smelting, 207 , Girod furnace, 209, 210 , Heroult furnace, 208 induction furnaces, Colby, 211 , de Ferranti, 211 , Kjellin, 211 , Rochling-Rodenhauser, 212, 213 , production of, 201 , welding, 215 Stetefeldt furnace, 47, 65 Sticht, copper converters used by, 96, 98 , shaft furnace, 78, 79 Stoves for heating blast, 182 Sublimation of antimony, 150 Sulphur, melting bismuth with, 116 , removal from tin ores, 134 Sulphuric acid parting of silver and gold, 51 Sulphurizing fusion, antimony refin- ing, 155 , Rossler's, 51 Talbot process, iron, 203 INDEX 271 Tarnowitz process, lead, 121 Temper carbon, 193 Thiel process, iron, 203 Thomas-Gilchrist process, iron, 202 Tin, 133 , blast furnace process, 136 , chemical concentration, 133 , Chinese furnace for smelting ore, 137 concentration of ores, 133 crystallization, 144 , electric furnace, Borchers and Mattonet, 138 , electrolytic solution and depo- sition, 141 liquation, 143 , mechanical concentration, 133 ores, 133 , oxidizing fusion, 144 properties, 145 , reducing fusion, 135 , removal of sulphur and arsenic, 134 , of tungstates, 134 , reverberatory furnace for smelt- ing, 136, 139 , smelting of poor slags, 141 , of rich slags, 140 , operations for oxidized ores, 135 Tin-plate waste, working up, 141 Tops, smelting of, for nickel, 165 Troostite, 193 Tungstates, removal from tin ores, 134 Tungsten, 225 , lixiviation, 226 ores, 225 , properties, 227 , reducing fusion, 225, 227 , separation of iron, manganese, and calcium from, 226 Tungstic acid, precipitation of, 226 , reduction of, 227 W Walker casting machine, copper, 103 Warlimont's process, nickel, 168 Washing with acids, mercury, 59 Washoe process, silver, 46 Water-jacket furnace, lead, 129 Weidtmann, constitution of leap matte, 124 Welding iron, 215 Wethey furnaces, copper, 67, 69 Wet-grinding amalgamation, silver, 38 Wet-mill amalgamation, 40 Wet tube mill, gold, 13 White iron, 190 White-Howell furnace, copper, 70 Wolframite, 225 Wrought iron, production of, 199 Wiist, constitution of iron-carbon alloys, 188, 192 , malleablizing, 198 U Upper Harz lead furnace, 128 Vacuum filter for gold slimes, 19 Vat with filter bottom and Butters' distributer, 10 Vitriolization process, Freiberg, 51 , Harz, 53 Ziervogel process, silver, 47 Zinc, 231 , Belgian process, 236 deposition, electrolytic, 246 , desilverization, 31 distillation, 245 , electrolytic deposition, 246 furnaces, Broken Hill, 247 , Faber du Faur, 247 , Schrubko, 237, 238 , liquating, 245 , lixiviation, Hopfner, 246 muffle system, 235 ores, 231 , precipitation of gold by, 16 properties, 248 retort press, 243 , Rhenish process, 239 , roast-reduction work, 232 , Silesian muffles, 235 SHORT-TITLE CATALOGUE OF THE PUBLICATIONS OF JOHN WILEY & SONS NEW YORK LONDON: CHAPMAN & HALL, LIMITED ARRANGED UNDER SUBJECTS Descriptive circulars sent on application. 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